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Form 8-K

sec.gov

8-K — i-80 Gold Corp.

Accession: 0001193125-26-397337

Filed: 2026-09-22

Period: 2026-09-21

CIK: 0001853962

SIC: 1040 (GOLD & SILVER ORES)

Item: Other Events

Item: Financial Statements and Exhibits

Documents

8-K — d29391d8k.htm (Primary)

EX-23.1 (d29391dex231.htm)

EX-23.2 (d29391dex232.htm)

EX-23.3 (d29391dex233.htm)

EX-23.4 (d29391dex234.htm)

EX-23.5 (d29391dex235.htm)

EX-23.6 (d29391dex236.htm)

EX-23.7 (d29391dex237.htm)

EX-99.1 (d29391dex991.htm)

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XML — IDEA: XBRL DOCUMENT (R1.htm)

8-K

8-K (Primary)

Filename: d29391d8k.htm · Sequence: 1

8-K

00-0000000 0001853962 false 0001853962 2026-09-21 2026-09-21 0001853962 us-gaap:CommonStockMember 2026-09-21 2026-09-21 0001853962 us-gaap:WarrantMember 2026-09-21 2026-09-21

UNITED STATES

SECURITIES AND EXCHANGE COMMISSION

Washington, D.C. 20549

FORM 8-K

CURRENT REPORT

Pursuant to Section 13 or15(d)

of the Securities Exchange Act of 1934

Date of Report (Date of earliest event reported): September 21, 2026

i-80 GOLD CORP.

(Exact name of registrant as specified in its charter)

British Columbia

001-41382

N/A

(State or other jurisdiction of

incorporation or organization)

(Commission

File Number)

(I.R.S. Employer

Identification No.)

150 York Street, Suite 1802, Toronto, Ontario

M5H 3S5

(Address of principal executive offices)

(Zip Code)

Registrant’s telephone number, including area code: (775) 525-6450

(Former name or former address, if changed since last report)

Check the appropriate box below if the Form 8-K filing is intended to simultaneously satisfy the filing obligation of the registrant under any of the following provisions:

☐

Written communications pursuant to Rule 425 under the Securities Act (17 CFR 230.425)

☐

Soliciting material pursuant to Rule 14a-12 under the Exchange Act (17 CFR 240.14a-12)

☐

Pre-commencement communications pursuant to Rule 14d-2(b) under the Exchange Act (17 CFR 240.14d-2(b))

☐

Pre-commencement communications pursuant to Rule 13e-4(c) under the Exchange Act (17 CFR 240.13e-4(c))

Securities registered pursuant to Section 12(b) of the Act:

Title of each class

Trading

Symbol(s)

Name of each exchange

on which registered

Common Shares

IAUX

NYSE American

Warrants to Purchase Common Shares

IAUX.WS

NYSE American

Common Shares

IAU

TSX

Warrants to Purchase Common Shares

IAU.WT.U

TSX

Indicate by check mark whether the registrant is an emerging growth company as defined in Rule 405 of the Securities Act of 1933 (§230.405 of this chapter) or Rule 12b-2 of the Securities Exchange Act of 1934 (§240.12b-2 of this chapter).

Emerging growth company ☒

If an emerging growth company, indicate by check mark if the registrant has elected not to use the extended transition period for complying with any new or revised financial accounting standards provided pursuant to Section 13(a) of the Exchange Act. ☐

Item 8.01

Other Events.

Updated Pre-Feasibility Study and Technical Report Summary for the Granite Creek Underground Project

On September 21, 2026, the Company provided the SEC Technical Report Summary, Pre-Feasibility Study, Granite Creek, Humboldt County, NV for its Granite Creek Underground Project, dated September 21, 2026 with an effective date of March 31, 2026, ( the “PFS”) prepared in accordance with the Securities and Exchange Commission S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for the Company by the following third-party qualified persons: SRK Consulting (U.S.), Inc., Global Resource Engineering, Practical Mining LLC, TR Raponi Consulting Ltd., Hatch Associates Consultants, Inc., and WSP USA Inc. The PFS was also prepared by Kenji Umeno, and employee of the Company. None of the third-party qualified persons are affiliated with the Company or any other entity that has an ownership, royalty, or other interest in the property.

Mineral resources and mineral reserves are reported using the definitions in Subpart 229.1300 – Disclosure by Registrants Engaged in Mining Operations in Regulation S-K 1300 (“S-K 1300”). All capital and operating cost estimates meet the requirements of S-K 1300. Unless otherwise indicated, all financial values are reported in United States dollars. Totals may not sum correctly due to rounding.

The information below is based on, or extracted from, the SEC Technical Report Summary, Pre-Feasibility Study, Granite Creek, Humboldt County, NV, dated September 21, 2026 with an effective date of March 31, 2026.

Table 1: Granite Creek Underground PFS Key Operating and Financial Metrics

Project Economics

Unit

Gold Price

$/oz

$

2,750

Pre-Tax NPV(5%)(3)

$M

$

138

After-Tax NPV(5%)(3)

$M

$

118

After-Tax Cash Flow(3)

$M

$

153

Production Profile

Mine Life

years

8.5

Total Ore Mined

000s tonnes

2,172

Mined Ore Grade

g/t Au

7.88

Total Stockpile Processed

000s tonnes

26

Total Ore Processed

000s tonnes

2,198

Gold Grade Processed

g/t Au

7.87

Total Gold Contained

000s oz

556

Average Gold Recovery

%

87

%

Total Gold Recovered

000s oz

485

Average Annual Gold Production (2028-2032)

000s oz

75.1

Average Annual Gold Production (LOM)

000s oz

53.9

Unit Operating Costs

LOM Operating Cost

Mining (ore and waste)

$/t mined

$

182

Processing (refractory and oxide)

$/t milled

$

122

Transportation & Power Costs

$/t milled

$

31

Surface Op Costs

$/t mined

$

54

G&A, Taxes and Royalties(7)

$/t milled

$

71

Total Cash Costs - LOM(4)

$/oz

$

2,076

Total Cash Costs - (2028-2032)(4)

$/oz

$

1,827

All-in Sustaining Costs - LOM(4,5)

(Excl. Lone Tree Refurbishment Capital)

$/oz

$

2,273

All-in Sustaining Costs - (2028-2032)(4,5)

(Excl. Lone Tree Refurbishment Capital)

$/oz

$

1,915

Total Capital Costs

Sustaining Capital

$M

$

82.9

Lone Tree Autoclave Capital Allocation

$M

$

49.3

Closure Costs

$M

$

12.7

Total Capital & Closure Costs

$M

$

144.9

For detailed metrics presented on an annual basis, see the Granite Creek Underground Detailed Cash Flow Model in Table 1 in the Appendix.

Initial Mineral Reserve Estimate and Updated Mineral Resource Estimate

The updated underground mineral resource estimate included in the PFS incorporates approximately 36,470 meters of drilling from 164 core holes conducted between 2023 through the end of 2025, relative to the resource estimate that formed the basis of the PEA published in March 2025.(1) The PEA has been superseded by this PFS. Numerous step-out drill holes have resulted in a larger mineralized envelope, while infill drilling increased definition and confidence in the mineral resource and supported conversion of a majority of the resource to mineral reserves.

The resource estimate was prepared using stope optimization software which is in line with industry recommended best practices for the assessment of the in-situ mineralization which satisfies the reporting criteria of the Reasonable Prospects for Eventual Economic Extraction (RPEEE). Additionally, a more conservative approach was applied to establish the creation of grade shell domains to match current mining methods, resulting in lower grades than in the previously published PEA. These updated grades are in line with the average grade currently mined at Granite Creek and are believed to more accurately reflect the expected grades of the orebody.

For more information regarding the estimation parameters and assumptions of the mineral resources and mineral reserves, please see Sections 11 and 12 of the PFS, respectively.

Exploration

The ongoing 2026 exploration program includes a combination of infill drilling to support continued reserve development for optimized ongoing mine operations and step-out drilling to expand the current and known mineralization within the South Pacific Zone, which remains open to the north and at depth, and the Rangefront, Otto, and Ogee zones.

Additional near-mine exploration targets include the CX Zone beneath the historical CX pit, which is not currently included in the underground mineral resource. Previous drilling in the CX Zone includes high-grade intercepts from legacy holes, including 18.0 g/t Au over 9.1 m in hole HPR 078 and 15.3 g/t Au over 9.1 m in RHC-1763, and i-80 hole of 36.4 g/t Au over 9.1 m in iGM21-06.(8) A successful drill program in the CX Zone has the potential to add high-grade ounces to the mineral resource and near-term mine plan. The Company intends to evaluate the high-grade portion of the CX Zone as an underground target as part of an anticipated extensive multi-year drill program planned to commence in 2027. Multiple near-mine targets are still to be evaluated over a multi-year drill program with the potential to extend the mine life.

Table 2: Granite Creek Underground Mineral Reserve Estimate as of March 31, 2026(2)

Mineral Reserves

Classification

Domain

Tonnage

(000s tonnes)

Au Grade

(g/t)

Au Contained

(000s oz)

Proven

Ogee

71

9.34

21.2

Otto

171

7.04

38.7

South Pacific

40

11.21

14.5

Stockpile

26

7.33

6.2

Total

308

8.13

80.6

Probable

Ogee

88

7.95

22.4

Otto

366

6.70

78.8

South Pacific

1,437

8.11

374.7

Total

1,890

7.83

475.9

Proven and Probable

Ogee

158

8.57

43.6

Otto

537

6.81

117.6

South Pacific

1,477

8.19

389.2

Stockpile

26

7.33

6.2

Total

2,198

7.87

556.5

Notes:

1.

All figures are rounded to reflect the relative accuracy of the estimates. Totals may not sum due to rounding. Mineral Reserves have been stated on the basis of a mine design, mine plan, and economic model.

2.

Mineral Reserves are reported using a stope center of gravity (“CoG”) range of 4.42 g/t (0.13 oz/st) Au – 6.41 g/t (0.19 oz/st) Au and an incremental processing CoG range of 1.51 g/t (0.04 oz/st) Au – 2.91 g/t (0.09 oz/st) Au depending on the processing method. The CoG calculations assume a US$2,500/oz Au price, and 60% to 92% metallurgical recovery depending on the processing method applied.

3.

Operating costs include mining (US$190.18/tonne), processing (US$59.94/tonne to US$131.13/tonne), G&A (US$22.29/tonne), shipping costs (US$16.04/tonne). Additionally, a 6% royalty and 0.75% excise tax are applied, along with treatment/refinery charges of US$1.85/oz Au.

4.

The reserves are mined using the underhand drift and fill method with cemented waste rock backfill. Mining dilution of 10% is included in the reserve at zero grade.

5.

Stockpile mineral reserves reflect ore mined and placed on surface stockpiles prior to the effective date of the PFS and not yet processed.

6.

The Mineral Reserves were estimated by SRK Consulting, U.S. (Inc.), a Qualified Person under S-K 1300.

All reported underground mineral reserves are fully attributable to the Company. The point of reference for the reporting of mineral reserves is the in-situ diluted tonnage and grades contained within the Mine Design, which accounts for the modifying factors.

Table 3: Exclusive Mineral Resource Statement for Granite Creek Underground – Effective Date March 31, 2026

Mineral Resource Estimate

Classification

Zone

Mass

(000s

tonnes)

Mass

Attributable

(000s

tonnes)

Au Grade

(g/t)

Material

Content Au

(000s oz)

Attributable

Material Content

Au (000s oz)

Measured

Ogee

175

175

6.82

38.4

38.3

Otto

166

166

5.58

29.7

29.7

SPZ

31

31

6.44

6.5

6.5

Total

372

372

6.23

74.6

74.5

Indicated

Ogee

241

241

6.52

50.6

50.6

Otto

425

425

5.23

71.5

71.5

SPZ

771

771

4.70

116.6

116.6

Total

1,438

1,438

5.16

238.6

238.6

Measured & Indicated

Ogee

417

417

6.64

89.0

89.9

Otto

591

591

5.32

101.2

101.2

SPZ

803

803

4.77

123.1

123.1

Total

1,810

1,810

5.38

313.3

313.1

Inferred

Ogee

38

38

6.71

8.3

8.3

Otto

417

417

7.62

102.0

102.0

SPZ

415

405

6.49

86.6

83.9

Total

871

860

7.04

197.0

194.2

Notes:

1.

Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, market or other relevant issues. The deposit has been classified as Measured, Indicated and Inferred based on confidence in the geological model and drill spacing. The quantity and grade of reported Inferred resources are uncertain in nature, and there has not been sufficient work to define these Inferred Mineral Resources as Indicated or Measured resources. There is no certainty that any part of a mineral resource will ever be converted into mineral reserves.

2.

Mineral Resources are reported on an exclusive basis with units reported in in Metric Tonnes, grade is reported in grams/tonnes, and metal in 000’s troy ounces.

3.

Mineral Resources are reported within mineable stopes which have been defined based on metal price assumptions,* variable metallurgical recovery assumptions, mining costs, processing costs, shipping, G&A costs. Metal Pricing is based on Gold (US$3,000 /oz). Other key assumptions include:

*

Mining Costs (US$/t 190.18);

*

Autoclave Processing Cost (US$/t 131.13), Low-grade Oxide (US$/t 59.94)

*

Autoclave Recoveries ranging 77.6%-93.4% (average 87.6%) and Oxide recoveries ranging 60.0%-87.4% (average 68.9%);

*

Shipping Costs (US$/t 16.04) and G&A Costs (US$/t 22.29);

*

Totaling US$/t 281.45 – 359.64 for underground mining and processing;

*

Based on the variable recoveries an average cut-off grade of 3.75 - 9.81 g/t (0.286 oz/st) has been used reporting.

4.

Mineral Resources have been depleted for previous underground mining activity, and reported on an attributable basis.

5.

Mineral Resources are based on validated data, which have been subjected to quality analysis/quality control (QA/QC) analysis, using capped, composited samples at 2 m (5ft) intervals. Estimation has been completed using a combination of OK and IDW estimation methodologies and classified based on confidence in the underlying data and drill spacing. Mineral resource tonnages have been rounded to reflect the precision of the estimate.

6.

The mineral resources were estimated by SRK Consulting (U.S.), Inc.

7.

Rounding of some figures may lead to minor discrepancies in totals.

8.

The point of reference for the reporting of mineral resources is the in-situ diluted tonnage and grades contained within the defined mineable stope optimizer limits.

Table 4: Granite Creek Open Pit Mineral Resource Statement

Deposit

Cutoff

Grade

(ppm)

Mass

(000’s tonnes)

Mass

(000’s short tons)

Mass

Attributable

(000’s tonnes)

Au

Grade

(g/t)

Au

Grade

(opt)

Au

Contained

(‘000s tr oz)

Attributable

Material Content

Au (000’s tr oz)

Measured

Pit B

0.2

3,900

4,299

3,575

1.06

0.031

133.49

122.37

Pit A

0.2

1,098

1,211

1,007

0.71

0.021

25.08

22.99

CX

0.2

13,493

14,873

12,835

1.10

0.032

478.41

454.02

Mag

0.2

15,055

16,596

14,703

1.03

0.030

499.50

487.90

Total

33,546

36,978

32,121

1.05

0.031

1,136.47

1,087.28

Indicated

Pit B

0.2

823

907

754

0.64

0.019

16.86

15.45

Pit A

0.2

969

1,069

889

0.68

0.020

21.06

19.30

CX

0.2

4,423

4,876

4,189

1.00

0.029

141.96

133.03

Mag

0.2

11,509

12,687

11,415

0.74

0.022

275.03

272.53

Total

17,725

19,538

17,247

0.80

0.023

454.91

440.32

Measured + Indicated

Pit B

0.2

4,723

5,206

4,329

0.99

0.029

150.34

137.82

Pit A

0.2

2,068

2,280

1,896

0.69

0.020

46.13

42.29

CX

0.2

17,916

19,749

17,024

1.08

0.031

620.37

587.05

Mag

0.2

26,565

29,283

26,119

0.91

0.026

774.53

760.43

Total

51,271

56,517

49,368

0.97

0.028

1,591.38

1,527.60

Inferred

Pit B

0.2

50

55

46

0.52

0.015

0.84

0.77

Pit A

0.2

440

485

403

0.44

0.013

6.22

5.70

CX

0.2

1,997

2,201

1,921

0.94

0.028

60.66

58.24

Mag

0.2

1,339

1,476

1,320

0.75

0.022

32.12

31.69

Total

3,826

4,217

3,689

0.81

0.024

99.84

96.40

1)

The effective date of the Mineral Resources Estimate is March 31, 2026

2)

The Qualified Persons for the estimate is GRE.

3)

Mineral resources are not ore reserves and are not demonstrably economically recoverable.

4)

Mineral resources are reported at a 0.20 g/t cutoff, an assumed gold price of 3,000 $/tr. oz, using variable recovery, a slope angle of 41 degrees, 6% royalty, heap leach processing cost $9.04 per tonne (includes admin), CIL processing cost of $17.22 per tonne (includes admin). The point of reference for the reporting of open pit Mineral Resources is the tonnage and grades contained within the Whittle pit shell limits, reported at a 0.20 g/t cut-off grade using variable metallurgical recove

5)

Attributable resources are those resources excluding certain thrid party claims in the Pinson #1A-18A unpatented mining claims and in the Section 28 Fee lands.

Figure 3: Granite Creek Underground Project Plan View

Figure 4: Granite Creek Underground Longitudinal Section

Project Overview

Granite Creek Underground is a fully permitted, constructed and operating gold mine currently ramping up towards steady-state production. It is the first brownfield project to be redeveloped under the Company’s development plan. The broader Granite Creek property (the “Property”) also hosts the Granite Creek open pit oxide deposit adjacent to the underground mine, which is currently advancing through early-stage permitting and technical work. Situated in the Potosi mining district, the Property lies approximately 43 kilometers northeast of Winnemucca, within Humboldt County, Nevada at the intersection of the highly prolific Battle Mountain-Eureka and Getchell gold trends.

Since 1980, the Property has produced approximately one million ounces of gold from both underground and open pit mining. The current underground minerals resources are primarily hosted adjacent to, but independent of, the past-producing open pits on the Property. Additionally, the Property hosts the Mag and CX oxide open pit resources.

Figure 5: Regional Map of i-80 Gold Assets in Northern Nevada

Geology and Mineralization

Mineralization at Granite Creek Underground is Carlin-type, with gold hosted in fine-grained arsenian pyrite similar to the nearby deposit at the Turquoise Ridge Complex. The primary host rocks at the Project are interbedded shale, siltstone, and limestone of the Ordovician Comus Formation. Higher-grade mineralization is found underground, proximal to the Cretaceous Osgood Mountains stock where the Comus Formation has been metamorphosed to marble and hornfels. Mineralization is strongly structurally controlled, typically by inverted thrust faults trending north to northeast. The Company has conducted significant exploration since acquiring the property in 2021 which led to the discovery of the South Pacific Zone, a northeastern extension of the existing underground deposit.

Mining

Granite Creek Underground is currently ramping up production and is expected to achieve commercial production in late 2026. The mine currently produces approximately 540 tonnes of ore per day and is operated by a local mining contractor with more than three decades of underground mining experience in northern Nevada. The mining rate is expected to rise to approximately 900 tonnes per day at full production supporting higher gold production. The PFS outlines an updated mine life of approximately 8.5 years with annual steady state gold production of approximately 75,000 ounces. The PFS incorporates the Project’s current operating costs, outlined in Table 6 and the Appendix.

Underground access is through portals located in the north wall of the CX pit. The main decline provides personnel and equipment access to all areas of the mine and can accommodate 30-ton haul trucks. Mining will be conducted using underhand drift and fill methods with cemented waste rockfill optimized for the Project’s ground conditions. Production drifts are 15 feet x 15 feet with a 14% decline. Mining of the Ogee and Otto zones are between the 4,790-foot level and 4,150-foot level and the decline to the South Pacific Zone is developed to the 3,950-foot level. Fresh air comes in from a fresh air portal at the bottom of the CX pit and is connected to the various workings through raises. The haulage ramp serves as an exhaust drift.

Production drift widths will be maintained at 15 feet or less to minimize dilution and enhance recovery. Underhand drift and fill mining is well suited for the ground conditions at the Project and allows for a high degree of selectivity, productivity, and safety for mine personnel. This mining method also results in backfill with superior geotechnical quality compared to the in-situ rock.

Processing

Metallurgical testing has demonstrated that the Ogee, Otto and South Pacific zones at Granite Creek Underground contain predominantly single refractory material and require an oxidation process to increase gold extraction using cyanidation. While samples responded to both alkaline and acid pressure oxidation, significant metallurgical testing has demonstrated that an acidic autoclave environment achieved higher sulfide oxidation and superior gold recovery rates. These results confirm that the Project’s ore can be processed via the Lone Tree Plant’s acid pressure oxidation (“POX”) conditions once the Plant is recommissioned.

The PFS incorporates an updated processing strategy that assumes third-party toll milling arrangements and associated over-the-road trucking costs through to mid-2027. To maximize the proportion of Granite Creek Underground sulfide material processed at the Company’s Lone Tree Plant and reduce reliance on third-party processing, the mine plan assumes the stockpiling of sulfide refractory material beginning approximately six months prior to the planned commissioning of the Company’s wholly owned Lone Tree Plant in late 2027. Beginning in 2028, the PFS incorporates the anticipated lower processing costs associated with hauling refractory ore and processing it through the Lone Tree Plant.(6) The acidic environment has higher operating costs versus alkaline, however, the higher costs are mostly offset by higher recoveries. The PEA envisaged a combination of acidic and alkaline operating environment.

Oxide mineralization is present in all zones within all deposits. Mine operations to date have encountered a higher proportion of oxide ore than what has been defined in the resource estimate and PFS. The oxide mineralization not accounted for in the PFS presents an opportunity to process this material in the Lone Tree Plant throughout the LOM with lower overall operating costs. Work is underway to determine the cost benefit. The PFS only incorporates a small component of oxide mineralization in the Ogee Zone. Oxide ore will be sold to a third party under an existing sales agreement until mid-2027 and then stockpiled similarly to the refractory ore. The oxide ore will be processed directly by the carbon-in-leach (“CIL”) circuit at the Lone Tree Plant, bypassing the POX circuit.

Ore at Granite Creek Underground is screened before shipping to remove oversize material, reducing the tonnes shipped and processed while increasing the grade delivered for processing. The potential economic benefits of this practice continue to be evaluated and were not incorporated into the PFS, representing a potential opportunity for future optimization.

A LOM processing schedule is illustrated in Figure 6.

Table 6: LOM Mining and Processing Unit Costs

LOM Mining and Processing Unit Costs

Mining (ore and waste)

$

/t mined

$

182

Surface Op Costs

$

/t mined

$

54

Processing - Lone Tree Autoclave

$

/t milled

$

105

Processing - Lone Tree Oxide

$

/t milled

$

34

Transportation

$

/t milled

$

16

Power Costs

$

/t milled

$

15

G&A

$

/t milled

$

23

Taxes and Royalties(7)

$

/t milled

$

49

Total

$

/t milled

$

458

Note to table above:

Numbers may not add up due to rounding and differences in $/t mined vs $/t milled.

Figure 6: LOM Annual Processing Schedule

Hydrology Update

Water inflow volumes to the mine have stabilized and continue to be managed well using the current underground pumping system, which presently operates near capacity. Work on an enhanced pumping system, that includes expanded sumps at lower levels and higher-capacity pumps, continues to advance. Pump installation will continue throughout the remainder of 2026 to increase overall water discharge capacity as the mine progresses at depth. The current dewatering model estimates residual passive inflow into Granite Creek will exceed 2,500 to 2,700 gallons per minute until the additional dewatering wells advance below the current workings. The second water treatment plant has been constructed and is currently in commissioning, which is expected to increase surface water treatment capacity by approximately 3,500 gallons per minute to support the Company’s long-term groundwater management objectives.

Capital Costs

LOM sustaining capital is estimated at $82.9 million, including all surface and underground development. A Lone Tree Plant refurbishment capital allocation of $49.3 million, as well as a closure and reclamation costs of $12.7 million are also included in the total PFS LOM capital cost (see Table 7). The Lone Tree autoclave refurbishment capital allocation of $49.3 million has been derived using a $25.97/tonne charge multiplied by the estimated total potential tonnage expected to be processed at the Lone Tree Plant (1.9 Mt) in the future from Granite Creek Underground.

Table 7: LOM Capital Cost Summary

LOM Capital

($Millions)

Underground Mine Capital

$

60.7

Surface Mine Capital*

$

22.2

Sustaining Capital

$

82.9

Lone Tree Plant Refurbishment Capital Allocation

$

49.3

Closure and Reclamation Costs

$

12.7

Total Capital

$

144.9

*

Surface mine capital includes approximately $5.6 million of capitalized costs in connection with underground mine development.

Operating Costs

During the five years of steady-state production following the commissioning of the Company’s wholly-owned Lone Tree Plant (2028 to 2032), the PFS estimates cash costs(4) of $1,827/oz Au and all-in sustaining costs(4,5) of $1,915/oz Au (see Table 8).

During the LOM, the PFS estimates cash costs(1) of $2,076/oz Au and all-in sustaining costs(4,5) of $2,273/oz Au, which includes the earlier period of third-party toll milling prior to Lone Tree’s commissioning and the final two years when production is currently expected to wind down (see Table 8).

Table 8 outlines the operating costs during the first five years of steady-state production following the commencement of processing at the Lone Tree Plant, excluding the earlier years of third-party toll processing and the final two years of the LOM.

The annual cash waterfall above in Figures 7 and 8 demonstrates the importance of the planned refurbishment of the Company’s Lone Tree Plant and the ending of third-party toll milling, which is expected to increase production and cash flow beginning in 2028 once commissioned. Figure 9 illustrates operating costs over the Project’s estimated production profile.

Table 8: Total and Unit Operating Cost Summary: Five-Year Steady-State Period Following Lone Tree Plant Commissioning

2028 - 2032

Total Costs

Unit Cost

Cost per Ounce

Cost Analysis

(Base Case Scenario of $2,750/oz Au)

($Millions)

($/t)

($/oz Au)

Mining (Ore and Waste)

$

271

$172/t mined

$

721

Processing (Refractory and Oxide)

$

177

$104/t milled

$

470

Transportation & Power Costs

$

48

$29/t milled

$

129

Surface Op Costs

$

77

$49/t mined

$

206

G&A, Taxes and Royalties(7)

$

113

$67/t milled

$

301

Total Operating Cost/Cash Costs(4)

$

686

$405/t milled

$

1,827

Sustaining Capital

$

33

$20/t milled

$

89

Lone Tree Plant Refurbishment Capital Allocation

$

41

$26/t milled*

$

108

Closure Costs

$

0

$0/t milled

$

0

All-in Sustaining Costs(4)

$

760

$449/t milled

$

2,023

All-in Sustaining Costs(4,5)

(Excluding Lone Tree Plant Refurbishment Capital)

$

720

$425/t milled

$

1,915

Notes to table above:

Numbers may not add up due to rounding and differences in $/t mined vs $/t milled.

*

Based solely on tonnes milled through the Lone Tree autoclave.

Table 9: Total and Unit Operating Cost Summary: LOM

LOM

Total Costs

Unit Cost

Cost per Ounce

Cost Analysis

(Base Case Scenario of $2,750/oz Au)

($Millions)

($/t )

($/oz Au)

Mining (Ore and Waste)

$

396

$182/t mined

$

816

Processing (Refractory and Oxide)

$

268

$122/t milled

$

552

Transportation & Power Costs

$

69

$31/t milled

$

142

Surface Op Costs

$

118

$54/t mined

$

243

G&A, Taxes and Royalties(7)

$

157

$71/t milled

$

323

Total Operating Cost/Cash Costs(4)

$

1,008

$458/t milled

$

2,076

Sustaining Capital

$

83

$38/t milled

$

171

Lone Tree Plant Refurbishment Allocation

$

49

$26/t milled*

$

102

Closure Costs

$

13

$6/t milled

$

26

All-in Sustaining Costs(4)

$

1,153

$524/t milled

$

2,375

All-in Sustaining Costs(4,5)

(Excluding Lone Tree Plant Refurbishment Capital)

$

1,103

$502/t milled

$

2,273

Notes to table above:

Numbers may not add up due to rounding and differences in $/t mined vs $/t milled.

*

Based solely on tonnes milled through the Lone Tree autoclave.

Figure 9: LOM Production and Cost Profile (Excluding Lone Tree Plant Refurbishment Capital Allocation)

Permitting

The Granite Creek Underground mine is fully permitted. The primary focus for the operations remains compliance and reporting requirements associated with existing site permits. Other than potential minor modifications to existing site permits for operational purposes, no other major permitting actions are expected in the foreseeable future.

Technical Disclosure and Qualified Persons

The PFS focuses solely on pre-feasibility-level technical work for Granite Creek Underground. The PFS does not update the work in the March 2025 PEA in respect of the Granite Creek open pit other than an updated mineral resource estimate, which remains subject to separate ongoing technical studies and permitting activities. The updated Granite Creek open pit mineral resource estimate included in the PFS accounted for an increase in gold price resulting in a larger optimized pit shell, equating to an increase in the overall open pit mineral resource inventory.

The PFS will contain detailed descriptions of data verification, sampling and assay procedures, quality assurance and quality control measures, key assumptions, parameters, risks and other factors supporting the mineral resource and mineral reserve estimates and the economic analysis. Inferred mineral resources are too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves. Mineral resources that are not mineral reserves do not have demonstrated economic viability, and there is no certainty that the results of the PFS will be realized.

Endnotes

(1)

A Preliminary Economic Assessment/Initial Assessment (“PEA”) on the Granite Creek Property was previously prepared by the Company in accordance with S-K 1300, filed in March 2025. The PFS focuses solely on feasibility-level technical work for Granite Creek Underground and does not update the previous work in the PEA in respect of the Granite Creek open pit, which remains subject to separate ongoing technical studies and permitting activities by the Company.

(2)

All reported mineral reserves for Granite Creek underground are fully attributable to the Company. All reported underground and open pit mineral resources are 100% attributable to the Company with the exception of a fee parcel and certain unpatented mining claims, which resources are 91.67% attributable to the Company.

(3)

The Mineral Reserve estimate is based on the Mineral Resource estimate depleted through March 31, 2026. The associated reserve mine plan therefore begins April 1, 2026. For purposes of the economic analysis, the reserve schedule was shifted one month such that the economic model begins May 1, 2026. This timing adjustment does not change the Mineral Reserve inventory, mine plan, or scheduled quantities, and does not result in double counting of production.

(4)

This is a non-GAAP measure. Please see the section titled “Non-GAAP Financial Performance Measures” below.

(5)

All-in sustaining costs presented exclude the allocation of Lone Tree Plant autoclave refurbishment capital. The Lone Tree autoclave capital allocation of $49.3M has been derived using a $25.97/tonne charge multiplied by the total tonnage processed at Lone Tree (1.90 million tonne).

(6)

Pending the successful completion of the refurbishment and commissioning of the Company’s Lone Tree autoclave and carbon-in-leach Plant.

(7)

Calculated based on a Base Case scenario of $2,750/oz; includes net profit interest (NPI) and refining. Royalties include a 10% NPI to Gold Royalty Corp. The illustrative NPI interests stated herein do not include the impacts of any future open pit projects on the Property, to which the NPI interest similarly applies. Payments made under the NPI are only made after recovery by the Company of all “recoverable costs” as further described under Chapter 3 of the PFS.

(8)

These historical drill results were generated by Homestake Mining Company in 1997 and Pinson Mining Company in 1995 and were not generated by i-80 Gold. The results have been reviewed by Tyler Hill, VP Geology for the Company. No QA/QC data is available for these results, however these holes were found to be acceptable for use in the 2026 Granite Creek open pit resource estimate as conducted by Global Resource Engineering (GRE).

Cautionary Statement Regarding Forward-Looking Information

Certain statements in this report constitute “forward-looking statements” or “forward-looking information” within the meaning of applicable securities laws, including but not limited to, statements regarding the updated results of the Technical Studies , such as future estimates of internal rates of return, net present value, future production, estimates of cash cost, proposed mining plans and methods, mine life estimates, cash flow forecasts, metal recoveries, estimates of capital and operating costs, timing for permitting and environmental assessments, timing, completion and results of feasibility studies, and the size and timing of phased development of the Project. Furthermore, forward-looking statements are necessarily based upon a number of estimates and assumptions that, while considered reasonable by the Company as of the date of such statements, are inherently subject to significant business, economic and competitive uncertainties and contingencies. With respect to this specific forward-looking information concerning the development of the Project, the Company has based its assumptions and analysis on certain factors that are inherently uncertain. Uncertainties include: (i) the adequacy of infrastructure; (ii) geological characteristics; (iii) metallurgical characteristics of the mineralization; (iv) the ability to develop adequate processing capacity; (v) the price of gold, silver and other commodities; (vi) the availability of equipment and facilities necessary to complete development; (vii) the cost of consumables and mining and processing equipment; (viii) unforeseen technological and engineering problems; (ix) natural disasters and/or accidents; (x) currency fluctuations; (xi) changes in regulations; (xii) the compliance by and/or key suppliers with terms of agreements; (xiii) the availability and productivity of skilled labour; (xiv) the regulation of the mining industry by various governmental agencies, including permitting and environmental assessments; (xv) the ability to raise sufficient capital to develop such projects; (xiv) changes in project scope or design; and (xv) political factors.

All statements other than statements of historical fact are forward-looking statements. Often, but not always, forward-looking statements can be identified by the use of words such as “plans”, “expects”, “is expected”, “budget”, “scheduled”, “estimates”, “continues”, “forecasts”, “projects”, “predicts”, “intends”, “anticipates” or “believes”, or variations of, or the negatives of, such words and phrases, or state that certain actions, events or results “may”, “could”, “would”, “should”, “might” or “will” be taken, occur or be achieved. Readers are cautioned that the assumptions used in the preparation of information, although considered reasonable at the time of preparation, may prove to be inaccurate and, as such, reliance should not be placed on forward-looking statements. The Company’s actual results, performance or achievement could differ materially from those expressed in, or implied by, these forward-looking statements and, accordingly, no assurance can be given that any of the events anticipated by the forward-looking statements will transpire or occur, or if any of them do so, what benefits, if any, that the Company will derive therefrom. By their nature, forward looking statements are subject to numerous risks and uncertainties, some of which are beyond the Company’s control, including general economic and industry conditions, volatility of commodity prices, title risks and uncertainties, uncertainty in geological, metallurgical and geotechnical studies and opinions, and ability to access sufficient capital from internal and external sources such as selling assets, restructuring debt or obtaining additional equity capital on terms that may be onerous or highly dilutive.

This report also contains references to estimates of mineral resources. The estimation of mineral resources is inherently uncertain and involves subjective judgments about many relevant factors. Mineral resources that are not mineral reserves do not have demonstrated economic viability. The accuracy of any such estimates is a function of the quantity and quality of available data, and of the assumptions made and judgments used in engineering and geological interpretation (including estimated future production from the Project, the anticipated tonnages and grades that will be mined and the estimated level of recovery that will be realized), which may prove to be unreliable and depend, to a certain extent, upon the analysis of drilling results and statistical inferences that may ultimately prove to be inaccurate. Mineral resource estimates may have to be re-estimated based on: (i) fluctuations in commodities prices; (ii) results of drilling, (iii) metallurgical testing and other studies; (iv) proposed mining operations, including dilution; (v) the evaluation of mine plans subsequent to the date of any estimates; and (vi) the possible failure to receive required permits, approvals and licenses or changes to existing mining licenses.

Forward-looking statements and information involve significant known and unknown risks and uncertainties, should not be read as guarantees of future performance or results and will not necessarily be accurate indicators of whether or not such results will be achieved. A number of factors could cause actual results to differ materially from the results expressed or implied by such forward-looking statements or information, including, but not limited to: the Company’s ability to finance the development of its mineral properties; assumptions and discount rates being appropriately applied to the PFS, uncertainty as to whether there will ever be production at the Company’s mineral exploration and development properties; risks related to the Company’s ability to commence production at the Project and generate material revenues or obtain adequate financing for its planned exploration and development activities; uncertainties

relating to the assumptions underlying resource and reserve estimates; mining and development risks, including risks related to infrastructure, accidents, equipment breakdowns, labor disputes, bad weather, non-compliance with environmental and permit requirements or other unanticipated difficulties with or interruptions in development, construction or production; the geology, grade and continuity of the Company’s mineral deposits; the uncertainties involving success of exploration, development and mining activities; permitting timelines; government regulation of mining operations; environmental risks; unanticipated reclamation expenses; prices for energy inputs, labor, materials, supplies and services; uncertainties involved in the interpretation of drilling results and geological tests and the estimation of reserves and resources; unexpected cost increases in estimated capital and operating costs; the need to obtain permits and government approvals; material adverse changes, unexpected changes in laws, rules or regulations, or their enforcement by applicable authorities; the failure of parties to contracts with the Company to perform as agreed; social or labor unrest; changes in commodity prices; and the failure of exploration programs or studies to deliver anticipated results or results that would justify and support continued exploration, studies, development or operations. For a more detailed discussion of such risks and other factors that could cause actual results to differ materially from those expressed or implied by such forward-looking statements, refer to i-80 Gold’s filings with Securities and Exchange Commission at www.sec.gov, including the most recent Annual Report on Form 10-K.

Non-GAAP Financial Performance Measures

The Company has included certain terms or performance measures in this report commonly used in the gold mining industry that are not defined under United States Generally Accepted Accounting Principles (“US GAAP”). This includes: all-in sustaining costs per ounce and cash cost per ounce – including and excluding the allocated Lone Tree Plant refurbishment capital. Non-IFRS/Non-GAAP financial performance measures do not have any standardized meaning prescribed under US GAAP, and therefore, they may not be comparable to similar measures employed by other companies. The data presented is intended to provide additional information and should not be considered in isolation or as a substitute for measures prepared in accordance with US GAAP and should be read in conjunction with the Company’s financial statements. Because the Company has provided these measures on a forward-looking basis, it is unable to present a quantitative reconciliation to the most directly comparable financial measure calculated and presented in accordance with US GAAP without unreasonable efforts. This is due to the inherent difficulty of forecasting the timing or amount of various reconciling items that would impact the most directly comparable forward-looking US GAAP measure that have not yet occurred, are outside of the Company’s control and/or cannot be reasonably predicted. The cost build-ups in the accompanying tables show the principal components used to calculate the disclosed measures.

Definitions

“All-in sustaining costs” is a non-GAAP financial measure calculated based on guidance published by the World Gold Council (“WGC”). The WGC is a market development organization for the gold industry and is an association whose membership comprises leading gold mining companies. Although the WGC is not a mining industry regulatory organization, it worked closely with its member companies to develop these metrics. Adoption of the all-in sustaining cost metric is voluntary and not necessarily standard, and therefore, this measure presented by the Company may not be comparable to similar measures presented by other issuers. The Company believes that the all-in sustaining cost measure complements existing measures and ratios reported by the Company. All-in sustaining cost includes both operating and capital costs required to sustain gold production on an ongoing basis. Sustaining operating costs represent expenditures expected to be incurred at the Project that are considered necessary to maintain production. Sustaining capital represents expected capital expenditures comprising mine development costs, including capitalized waste, and ongoing replacement of mine equipment and other capital facilities, and does not include expected capital expenditures for major growth projects or enhancement capital for significant infrastructure improvements. The Company’s presentation of AISC excluding the allocated Lone Tree Plant refurbishment capital is an additional non-GAAP measure and should be read together with the accompanying cost build-up. “Cash cost per gold ounce” is a common financial performance measure in the gold mining industry but has no standard meaning under US GAAP. The Company believes that, in addition to conventional measures prepared in accordance with US GAAP, certain investors use this information to evaluate the Company’s performance and ability to generate cash flow. Cash cost figures are calculated in accordance with a standard developed by The Gold Institute. The Gold Institute ceased operations in 2002, but the standard is considered the accepted standard of reporting cash cost of production in North America. Adoption of the standard is voluntary, and the cost measures presented may not be comparable to other similarly titled measures of other companies.

For a more detailed breakdown on how these measures were calculated, please see the Tables 7 and 8.

APPENDIX

Table 1: Granite Creek Underground Project Detailed Cash Flow Model

Table is continued on next page

APPENDIX Table 1: Granite Creek Underground Project Detailed Cash Flow Model Granite Creek Underground UNITS TOTAL LOM 2028 - 2032 (2) 2026E 2027E 2028E 2029E 2030E 2031E 2032E 2033E 2034E MINING Mine Life Years 8.3 5 Refractory Ore Mined k tonnes 2,139 1,560 161.8 192.0 330.3 329.9 289.0 336.7 273.6 175.9 50.2 Refractory Ore Mined Grade g/t 7.85 7.86 7.52 7.42 7.67 7.71 8.06 7.88 8.04 8.09 9.20 Oxide Ore Mined k tonnes 33 13 7.7 11.7 4.7 3.3 2.5 2.6 0.3 - - Oxide Ore Mined Grade g/t 10.02 10.43 11.8 8.4 12.1 9.1 10.8 9.7 3.1 - - Total Ore Mined k tonnes 2,172 1,573 169.6 203.7 335.0 333.2 291.5 339.3 274.0 175.9 50.2 Daily Mining Rate tpd 724 874 465 558 918 913 799 930 751 482 137 Stockpile Used in Processing k tonnes 26 - 26.2 Stockpile Grade g/t 7.33 7.3 PROCESSING Total Material for Processing k tonnes 2,198 1,692 162 118 454 333 292 339 274 176 50 Gold Average Grade g/t Au 7.87 7.9 7.77 7.07 7.75 7.73 8.08 7.89 8.04 8.09 9.20 Contained Gold ‘000 oz Au 556.5 429 40.5 26.8 113.2 82.8 75.7 86.1 70.8 45.8 14.8 Toll Mill Processing (3rd party processing) Total Tonnes Processed k tonnes 266 - 156 110 Gold Grade g/t Au 7.33 0.00 7.59 6.96 Contained Gold ‘000 oz Au 63 - 38 25 Gold Average Recovery % 90% 0% 90% 90% Recovered Gold ‘000 oz Au 56 - 34.4 22.1 Lone Tree Autoclave Processing Total Tonnes Processed k tonnes 1,899 1,673 444 330 289 337 274 176 50 Gold Grade g/t Au 7.91 7.85 7.70 7.71 8.06 7.88 8.04 8.09 9.20 Contained Gold ‘000 oz Au 483 423 110 82 75 85 71 46 15 Gold Average Recovery % 87% 87% 87% 87% 87% 87% 87% 87% 87% Recovered Gold ‘000 oz Au 422 371 98.2 71.8 65.8 74.2 61.2 38.2 12.4 Oxide Processing (Third-party ore purchase agreement until 2028) Total Tonnes Processed k tonnes 33 19 5.6 8.1 10.4 3.3 2.5 2.6 0.3 - - Gold Grade g/t Au 10.02 9.85 12.93 8.43 10.12 9.08 10.77 9.71 3.14 - - Silver Grade g/t Ag 0.00 0.00 - - - - - - - - Contained Gold ‘000 oz Au 10.6 6.1 2.3 2.2 3.4 1.0 0.9 0.8 0.0 - - Gold Payability / Recovery % 69% 76% 60% 60% 76% 75% 74% 75% 66% - - Recovered Gold ‘000 oz Au 7.3 4.6 1.4 1.3 2.6 0.7 0.6 0.6 0.0 - - Total Tonnes Processed k tonnes 2,198 1,692 162.1 118.0 454.4 333.2 291.5 339.3 274.0 175.9 50.2 Total Gold Production(3) ‘000 oz Au 485.4 375.7 35.7 23.5 100.8 72.5 66.4 74.8 61.2 38.2 12.4 Recoverable Ag ounces remaining (BOP) REVENUE Gold Price US$/oz Au $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 $2,750 Revenues (@99.75% payability) US$M $1,331 $1,031 $98 $64 $276 $199 $182 $205 $168 $105 $34 OPERATING COSTS Mining Costs (Mineralized Material incl. Backfill) Surface Op Costs Processing - Refractory / Autoclave Processing - Refractory / Third party Processing - Oxide (2028 Onwards) Transportation Power Costs G&A Total Operating Cost US$M US$M US$M US$M US$M US$M US$M US$M US$M $396 $118 $200 $68 $0.65 $36 $33 $50 $271 $77 $176 $0 $0.65 $28 $20 $30 $33.2 $40.7 $52.6 $56.9 $55.3 $57.8 $48.3 $35.4 $15.9 $7.0 $13.5 $14.9 $15.6 $15.7 $16.0 $15.2 $12.0 $7.9 - - $46.7 $34.7 $30.4 $35.4 $28.8 $18.5 $5.3 $39.7 $27.9 - - - - - - - - - $0.4 $0.1 $0.1 $0.1 $0.0 - - $1.3 $2.8 $6.0 $6.0 $5.2 $6.1 $4.9 $3.1 $0.9 $1.8 $3.3 $3.7 $4.1 $4.2 $4.2 $4.2 $4.2 $3.1 $3.2 $5.8 $5.9 $6.3 $6.2 $6.0 $6.1 $5.8 $4.5 $86.3 $93.9 $130.2 $123.6 $117.1 $125.5 $107.4 $79.0 $37.7 $901 $604 Refining Royalties & State Taxes(4) NPI4 US$M US$M US$M $2.6 $104.5 $0.0 $2.0 $80.4 $0.0 $0.2 $0.1 $0.5 $0.4 $0.4 $0.4 $0.3 $0.2 $0.1 $7.3 $4.6 $20.0 $14.4 $13.3 $18.0 $14.7 $9.2 $3.0 - - - - - - - - - Unit Costs ($/t) Mining Costs (Ore & Waste ) Surface Op Costs Processing - Refractory / Autoclave Processing - Refractory / Third party Processing - Oxide (2028 Onwards) Transportation Power Costs G&A Total US$/t mined US$/t mined US$/t milled US$/t milled US$/t milled US$/t milled US$/t milled US$/t milled US$/t milled $182 $54 $105 $254 $34.0 $16.5 $172 $49 $105 $0 $34.0 $16.6 $196 $200 $157 $171 $190 $170 $176 $201 $318 $41 $66 $45 $47 $54 $47 $55 $68 $158 - - $105 $105 $105 $105 $105 $105 $105 $254 $254 - - - - - - - - - $34 $34 $34 $34 $34 - - $8 $24 $13 $18 $18 $18 $18 $18 $18 $11 $28 $8 $12 $14 $12 $15 $24 $62 $20 $49 $13 $19 $21 $18 $22 $33 $90 $15 $23 $12 $18 $410 $357 $532 $796 $286 $371 $402 $370 $392 $449 $751 Table is continued on next page Notes to table above: This is a non-GAAP measure. Please see the section titled “Non-GAAP Financial Performance Measures” above. The Mineral Reserve estimate is based on the block model depleted through March 31, 2026. The associated reserve mine plan therefore begins April 1, 2026. For purposes of the economic analysis, the reserve schedule was shifted one month such that the economic model begins May 1, 2026. This timing adjustment does not change the Mineral Reserve inventory, mine plan, or scheduled quantities and does not result in double counting of production. The PFS is based on a mineral reserve and resource estimate derived from drilling data 2023 through to the end of 2025 and should not be relied upon for production forecasts in 2026. The Company expects recovered gold ounces of between 30,000 to 40,000 ounces in 2026 from Granite Creek Underground. Royalties include a 10% net profit interest (NPI) to Gold Royalty Corp. and refining. The illustrative NPI interests stated herein do not include the impacts of any future open pit projects on the Property, to which the NPI interest similarly applies. Payments made under the NPI are only made after recovery by the Company of all “recoverable costs” as further described under Chapter 4 of the NI-43-101 Study. 5. All-in Sustaining Costs presented excludes the allocation of Lone Tree autoclave refurbishment capital.

Granite Creek underground UNITS TOTAL LOM 2028-2032 2026(2)E 2027E 2028E 2029E 2030E 2031E 2032E 2033E 2034E CAPITAL EXPENDTURES Lone Tree Capital Allocation US$M $49.3 $40.5 - $3.0 $8.6 $8.6 $7.5 $8.7 $7.1 $4.6 $1.3 Mine Capex US$M $60.7 $30.1 $16.6 $11.8 $10.3 $6.7 $4.4 $4.4 $4.4 $2.3 - Surface Capitalized Costs US$M $22.2 $3.2 $15.6 $3.4 $1.7 $0.6 $0.5 $0.3 $0.0 - - Closure Costs US$M $12.7 $0.0 - - - - - - - - $12.7Total Capital US$M $144.9 $73.8 $32.2 $18.1 $20.6 $15.9 $12.4 $13.5 $11.5 $6.8 $14.0 CASH COSTS & AISC Total Cash Costs (Inc. Royalties)(1) US$/oz $2,076 $1,827 $2,622 $4,208 $1,496 $1,1910 $1,968 $1,922 $2,002 $2,317 $3,297 All-in Sustaining Costs(1)(5) US$/oz $2,375 $2,023 $3,524 $4,980 $1,701 $2,129 $2,154 $2,102 $2,190 $2,496 $4,428All-in Sustaining Costs, Excluding Lone Tree US$/oz $2,273 $1,915 $3,524 $4,854 $1,615 $2,011 $2,041 $1,985 $2,073 $2,376 $4,322 Capital Allocation(1)(5) CASH FLOW ANALYSIS Revenue net of Refining US$M $1,329 $1,028 $98 $64 $276 $198 $182 $205 $167 $104 $34 Royalty and NPI(4) US$M ($104) ($80) ($7) ($5) ($20) ($14) ($13) ($18) ($15) ($9) ($3) Operating Costs US$M ($901) ($604) ($86) ($94) ($130) ($124) ($117) ($125) ($107) ($79) ($38) Depreciation US$M ($145) ($57) ($1) ($5) ($8) ($10) ($12) ($13) ($15) ($16) ($66) Net Operating Income US$M $179 $287 $3 ($40) $118 $51 $40 $48 $31 $0 ($73) Income Taxes US$M ($25) ($25) ($0) ($0) ($5) ($3) ($5) ($7) ($5) ($0) ($0) Net Income US$M $153 $263 $3 ($40) $113 $47 $35 $42 $25 $0 ($73) Depreciation US$M $145 $57 $1 $5 $8 $10 $12 $13 $15 $16 $66 Working Capital US$M $0 ($7) ($5) ($11) ($8) ($1) ($7) ($2) $10 ($1) $3 Operating Cash Flow US$M $298 $312 ($1) ($24) $113 $56 $40 $53 $50 $15 ($4)Capital Expenditures (incl. closure costs) US$M ($145) ($74) ($32) ($18)($21) ($16) ($12) ($13) ($11) ($7) ($14) NET CASH FLOW(2) US$M $153 $238 ($33) ($42) $92 $40 $28 $40 $38 $8 ($18) After-tax NPV 5% of discounting(2) US$M $118

Notes to table above:

1.

This is a non-GAAP measure. Please see the section titled “Non-GAAP Financial Performance Measures” above.

2.

The Mineral Reserve estimate is based on the block model depleted through March 31, 2026. The associated reserve mine plan therefore begins April 1, 2026. For purposes of the economic analysis, the reserve schedule was shifted one month such that the economic model begins May 1, 2026. This timing adjustment does not change the Mineral Reserve inventory, mine plan, or scheduled quantities and does not result in double counting of production.

3.

The PFS is based on a mineral reserve and resource estimate derived from drilling data 2023 through to the end of 2025 and should not be relied upon for production forecasts in 2026. The Company expects recovered gold ounces of between 30,000 to 40,000 ounces in 2026 from Granite Creek Underground.

4.

Royalties include a 10% net profit interest (NPI) to Gold Royalty Corp. and refining. The illustrative NPI interests stated herein do not include the impacts of any future open pit projects on the Property, to which the NPI interest similarly applies. Payments made under the NPI are only made after recovery by the Company of all “recoverable costs” as further described under Chapter 4 of the NI-43-101 Study.

Item 9.01

Financial Statements and Exhibits.

(d) Exhibits.

Exhibit

No.

Description

23.1

Consent of SRK Consulting (U.S.), Inc.

23.2

Consent of Global Resource Engineering

23.3

Consent of TR Raponi Consulting Ltd.

23.4

Consent of WSP USA Inc.

23.5

Consent of Practical Mining LLC

23.6

Consent of Hatch Associates Consultants, Inc.

23.7

Consent of Kenji Umeno

99.1

SEC Technical Report Summary, Pre-Feasibility Study, Granite Creek, Humboldt County, NV dated September 21, 2026 with an effective date of March 31, 2026

104

Cover Page Interactive Data File (embedded with the inline XBRL document)

SIGNATURES

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned hereunto duly authorized.

i-80 GOLD CORP.

Date: September 22, 2026

By:

/s/ Ryan Snow

Ryan Snow

Chief Financial Officer

EX-23.1

EX-23.1

Filename: d29391dex231.htm · Sequence: 2

EX-23.1

Exhibit 23.1

CONSENT OF SRK CONSULTING (U.S.), INC.

In connection with the i-80 Gold Corp. Current Report on Form 8-K dated

September 21, 2026 (the “Form 8-K”), the undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

SRK Consulting (U.S.), Inc.

By:

/s/ Benjamin Parsons

Name:

Benjamin Parsons

Title:

SRK Consulting (U.S.), Inc. President

EX-23.2

EX-23.2

Filename: d29391dex232.htm · Sequence: 3

EX-23.2

Exhibit 23.2

CONSENT OF GLOBAL RESOURCE ENGINEERING

In connection with the i-80 Gold Corp. Current Report on Form 8-K dated

September 21, 2026 (the “Form 8-K”), the undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

Global Resource Engineering

By:

/s/ Terre Lane

Name:

Terre Lane

Title:

Principal Mining Engineer

EX-23.3

EX-23.3

Filename: d29391dex233.htm · Sequence: 4

EX-23.3

Exhibit 23.3

CONSENT OF TR RAPONI CONSULTING LTD.

In

connection with the i-80 Gold Corp. Current Report on Form 8-K dated September 21, 2026 (the “Form 8-K”), the

undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

TR Raponi Consulting Ltd.

By:

/s/ Tommaso Roberto Raponi

Name:

Tommaso Roberto Raponi

Title:

President

EX-23.4

EX-23.4

Filename: d29391dex234.htm · Sequence: 5

EX-23.4

Exhibit 23.4

CONSENT OF WSP USA INC.

WSP USA Inc.

(“WSP”) states that it is responsible for preparing or supervising the preparation of part(s) of the Technical Report Summary, titled “SEC Technical Report Summary Pre-Feasibility Study

Granite Creek, Humboldt County, NV” (the “Technical Report Summary”) dated September 21, 2026, with an effective date of March 31, 2026.

Furthermore, WSP states that:

I.

WSP consents to the public filing of the Technical Report Summary by

i-80 Gold Corp. (the “Company”);

II.

The Technical Report Summary was prepared in accordance with Subpart 1300 of Regulation S-K promulgated by the U.S. Securities and Exchange Commission and supports the Current Report on Form 8-K dated September 21, 2026 (the “Form 8-K”);

III.

WSP consents, as a company whose principal business is providing engineering or geoscientific services, and

whose business gives authority to a statement made by the company, to the use of WSP’s name in the Form 8-k, to any quotation from or summarization in the Form 8-K

of the parts of the Technical Report Summary for which WSP is responsible, to the filing of the Technical Report Summary as an exhibit to the Form 8-K, and to the incorporation by reference in the

Company’s Registration Statements on Form S-8 (File No. 333-280725) (the “Form S-8 Registration

Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531 and

333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”);

IV.

WSP confirms that it has read the portions of the Form 8-K relating to

the parts of the Technical Report Summary for which WSP is responsible, and that such portions of the 8-K fairly and accurately reflect such information.

Dated September 21, 2026

WSP USA, Inc.

Authorized Signatory

/s/ Mathew Oommen

Name: Mathew Oommen

Title: Senior Vice President

EX-23.5

EX-23.5

Filename: d29391dex235.htm · Sequence: 6

EX-23.5

Exhibit 23.5

CONSENT OF PRACTICAL MINING LLC

In

connection with the i-80 Gold Corp. Current Report on Form 8-K dated September 21, 2026 (the “Form 8-K”), the

undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

Practical Mining LLC

By:

/s/ Dagny G. Odell

Name:

Dagny G. Odell

Title:

Manager

EX-23.6

EX-23.6

Filename: d29391dex236.htm · Sequence: 7

EX-23.6

Exhibit 23.6

CONSENT OF HATCH

In connection with the i-80 Gold Corp. Current Report on Form 8-K dated September 21, 2026 (the “Form 8-K”), the undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

Hatch

By:

/s/ William McCombe

Name:

William McCombe, P.Eng

Title:

Global Director, Gold

EX-23.7

EX-23.7

Filename: d29391dex237.htm · Sequence: 8

EX-23.7

Exhibit 23.7

CONSENT OF KENJI UMENO, P.ENG

In

connection with the i-80 Gold Corp. Current Report on Form 8-K dated September 21, 2026 (the “Form 8-K”), the

undersigned consents to:

i.

the filing and use of the technical report summary titled “SEC Technical Report Summary Pre-Feasibility Study Granite Creek, Humboldt County, NV” dated September 21, 2026, with an effective date of March 31, 2026 (the “Granite Creek Mine 1300 Report”) as referenced in the

Form 8-K;

ii.

the use of and reference to their name, including their status as an expert or “qualified person”,

(as described in Subpart 1300 of Regulation S-K promulgated by the Securities and Exchange Commission), in connection with the Granite Creek Mine 1300 Report, the Form

8-K, the Registration Statement on Form S-8 (File No. 333-280725) (the “Form

S-8 Registration Statement”) and the Registration Statements on Form S-3, as amended and supplemented (File Nos. 333-286531

and 333-287243) (together with the Form S-8 Registration Statement, “Registration Statements”); and

iii.

any extracts or summaries of the Granite Creek Mine 1300 Report included or incorporated by reference in the

Form 8-K or the Registration Statements and the use of any information derived, summarized, quoted or referenced from the Granite Creek Mine 1300 Report, or portions thereof, that was prepared by them, that

they supervised the preparation of, and/or that was reviewed and approved by them, that is included or incorporated by reference in the Form 8-K and which is incorporated by reference in the Registration

Statements.

Date: September 21, 2026

By:

/s/ Kenji Umeno

Name:

Kenji Umeno, P.Eng

EX-99.1

EX-99.1

Filename: d29391dex991.htm · Sequence: 9

EX-99.1

Exhibit 99.1

SEC Technical Report Summary

Pre-Feasibility Study

Granite Creek

Humboldt County, NV

Effective Date: March 31, 2026

Report Date:

September 21, 2026

Report Prepared for

i-80 Gold Corp.

5190 Neil Road Suite 460

Reno, Nevada 89502

Report Prepared by

SRK Consulting (U.S.), Inc.

999 Seventeenth Street, Suite 400

Denver, CO 80202

SRK Project Number: USPR002507

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page ii

Table of Contents

1  Executive Summary

1

1.1  Property Description (Including Mineral Rights) and Ownership

1

1.2  Geology and Mineralization

1

1.3  Status of Exploration, Development and Operations

2

1.4  Mineral Resource and Mineral Reserve Estimates

3

1.4.1 Underground

3

1.4.2 Open Pit

5

1.5  Metallurgical Testing and Recovery Methods

6

1.6  Mineral Reserve

7

1.7  Summary Capital and Operating Cost Estimates

9

1.8  Economic Analysis

10

1.9  Permitting Requirements

12

1.10  Conclusions and Recommendations

14

1.10.1 Drilling

14

1.10.2 Mining

14

1.10.3 Metallurgical Testing

15

1.10.4 Permitting / Environmental

15

1.10.5 Economics

15

2  Introduction

16

2.1  Registrant for Whom the Technical Report Summary was Prepared

16

2.2  Terms of Reference and Purpose of the Report

16

2.3  Sources of Information

16

2.4  Details of Inspection

16

2.5  Report Version Update

17

3  Property Description

18

3.1  Property Description

18

3.2  Mineral Tenure and Surface Rights

18

3.2.1 Owned Unpatented Mining Claims

22

3.2.2 Leased Unpatented Mining Claims

22

3.2.3 Fee Lands

22

3.3  Royalties and Net Profit Interests

22

3.3.1 1996 Royal Gold Royalty – Granite Creek Property (Royal Gold, Inc. –

Current Holder)

23

3.3.2 2026 Franco- Nevada Royalty – Granite Creek Property – 1.5% to 3%

NSR

24

3.3.3 Pinson Private Royalty – Pinson Claims – 3.125% NSR (Noceto, Phillips,

Murphy - Current Holders) & 0.17% NSR (FN USA & S&G Pinson – Current Holders)

24

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page iii

3.3.4 Pinson Royalty – Section 21 Bee Dee 21-56 Claims – 0.5% NSR (Pinson

Mining Company – Current Holder)

24

3.3.5 2001 Royal Gold Royalty – Section 21, 29 & Pacific Claims – 3-5%

NSR (Royal Gold Inc. - Current Holder)

24

3.3.6 Goldfield Royalty – Section 33 – 2% NSR (FN USA - Current

Holder)

25

3.3.7 Bee Dee Reserved Royalty - Bee Dee Claims - 2% NSR Royalty (FN USA (50%) and

S&G Pinson, LLC (50%) – Current Holders)

25

3.3.8 Section 28 Private Royalty – Section 28 Fee lands - 2% NSR (Murphy, Noceto,

Phillips, D. and J. Christison – Current Holders)

25

3.3.9 Mineral Production Royalty Agreement – Granite Creek Property - 10% Net

Profit Interest (NGM - Current Holder)

25

3.4  History of Ownership and Prior Mining

26

3.5  Environmental Liabilities

26

3.6  Permits and Authorizations

26

4  Accessibility, Climate, Local Resources, Infrastructure and

Physiography

27

4.1  Topography, Elevation and Vegetation

27

4.2  Means of Access

27

4.3  Climate and Length of Operating Season

27

4.4  Infrastructure Availability and Sources

27

4.4.1 Water

28

4.4.2 Electricity

28

4.4.3 Personnel

28

4.4.4 Supplies

28

5  History

29

5.1  Previous Operations

29

5.2  Exploration and Development of Previous Owners or Operators

29

6  Geological Setting, Mineralization, and Deposit

31

6.1  Regional, Local and Property Geology

31

6.1.1 Regional Geology

31

6.1.2 Local Geology

33

6.1.3 Property Geology

36

6.2  Mineral Deposit

42

6.3  Stratigraphic Column and Local Geology Cross-Section

42

7  Exploration

45

7.1  Exploration Work (Other Than Drilling)

45

7.1.1 Procedures and Parameters Relating to the Surveys and Investigations

45

7.1.2 Sampling Methods and Sample Quality

49

7.1.3 Information About the Area Covered

49

7.1.4 Significant Results and Interpretation

49

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Page iv

7.2  Exploration Drilling

57

7.2.1 Drilling Type and Extent

57

7.2.2 Drilling, Sampling, or Recovery Factors

67

7.2.3 Drilling Results and Interpretation

68

7.3  Hydrogeology

78

7.4  Geotechnical Data, Testing and Analysis

85

7.5  Property Plan View

85

7.6  Exploration Target

85

8  Sample Preparation, Analysis and Security

86

8.1  Sample Preparation Methods and Quality Control Measures

86

8.2  Sample Preparation, Assaying and Analytical Procedures

86

8.3  Quality Control Procedures/Quality Assurance

87

8.4  Opinion on Adequacy

102

8.5  Non-Conventional Industry Practice

103

9  Data Verification

104

9.1  Data Verification Procedures

104

9.2  WSP Verification

119

9.3  Limitations

119

9.4  Opinion on Data Adequacy

119

10  Mineral Processing and Metallurgical Testing

121

10.1  Metallurgical Testing Programs

121

10.2  Sample Representativeness

136

10.3  Laboratories

149

10.4  Relevant Results

150

10.5  Adequacy of Data and Non-Conventional Industry Practice

170

11  Mineral Resource Estimate

172

11.1  Open Pit

172

11.1.1 Drillhole Database

172

11.1.2 Topography

174

11.1.3 Geologic Model

175

11.1.4 Estimation Domains

176

11.1.5 Compositing

183

11.1.6 Evaluation of Outliers

186

11.1.7 Density

187

11.1.8 Variography

187

11.1.9 Block Model Parameters

188

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Page v

11.1.10 Estimation Domains

188

11.1.11 Estimation Parameters

189

11.1.12 Geometallurgical Modeling

190

11.1.13 Block Model Validation

191

11.1.14 Mineral Resource Classification

199

11.1.15 Open Pit Mineral Resource Statement

201

11.1.16 Mineral Resource Sensitivity by Domain

203

11.1.17 Factors that Could Affect Open Pit Mineral Resources

206

11.2  Underground Mineral Resources

206

11.2.1 Drillhole Database

206

11.2.2 Topography and Mining Depletion

208

11.2.3 Geological Model

210

11.2.4 Estimation Domain Analysis

210

11.2.5 Compositing

223

11.2.6 Assay Capping

226

11.2.7 Variogram Analysis and Modeling

230

11.2.8 Block Model Parameters

233

11.2.9 Estimation Methodology

234

11.2.10 Geo-Metallurgical Parameters

241

11.2.11 Density

251

11.2.12 Model Validation

251

11.2.13 Classification

264

11.2.14 Underground Mineral Resource Statement

268

11.2.15 Mineral Resource Sensitivity

271

11.2.16 Sources of Uncertainty

273

11.2.17 Reconciliation

274

11.2.18 Relevant Factors

275

12  Mineral Reserve Estimates

282

12.1  Key Assumptions, Parameters, and Methods Used

282

12.2  Cut-Off Grades Estimates

285

12.3  Reserves Classification and Criteria

286

12.4  Multiple Commodity Reserve

286

12.5  Impact of Changes of Modifying Factors to the Mineral Reserve

Estimates

286

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SEC

Technical Report Summary – Granite Creek

Page vi

13  Mining Methods

287

13.1  Parameters Relevant to Mine Designs and Plans

287

13.1.1 Geotechnical

287

13.1.2 Hydrological

293

13.1.3 Ventilation

295

13.2  Production Rates, Expected Mine Life, Mining Unit Dimensions, and Mining

Dilution and Recovery Factors

301

13.3  Requirements for Stripping, Underground Development, and Backfilling

308

13.4  Required Mining Equipment Fleet and Machinery, and Personnel

309

13.5  Final Mine Outline Map

311

14  Processing and Recovery Methods

313

14.1  Lone Tree Facility Historical Processing

313

14.2  Lone Tree Facility Process Description

313

14.2.1 Key Process Design Criteria

316

14.2.2 Ore Reclaim

316

14.2.3 Grinding

317

14.2.4 Thickening and Acidulation

318

14.2.5 Pressure Oxidation

318

14.2.6 POX Off-Gas Treatment

319

14.2.7 Neutralization

320

14.2.8 Carbon-in-Leach

320

14.2.9 Carbon Elution and Regeneration

321

14.2.10 Cyanide Destruction

322

14.2.11 Tailings Management

322

14.2.12 Water Distributions

323

14.2.13 Reagents

324

14.2.14 Oxygen Plant

324

14.2.15 Compressed Air

325

14.2.16 Utilities Consumption

325

14.2.17 Water Consumption

325

14.3  Third-Party Refractory Processing

326

15  Infrastructure

329

15.1  Summary

329

15.2  Access and Local Resources

329

15.3  Administrative Facilities

330

15.4  Dewatering

331

15.4.1 Dewatering Wells

331

15.4.2 Monitoring Wells and VWPs

334

15.4.3 Underground Contact Water Management

338

15.4.4 Water Treatment and Rapid Infiltration Basins

339

15.4.5 Fresh Water Supply

343

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Technical Report Summary – Granite Creek

Page vii

15.5  Electrical

343

15.6  Portal Facilities

344

15.7  Backfill Aggregate Crusher

346

15.8  Screen Plant

346

15.9  Waste Rock Disposal (WRD)

347

15.10 Waste Disposal

349

15.11 Communications

350

15.12 Equipment

350

16  Market Studies

351

16.1  Market Information

351

16.2  Contracts and Status

353

17  Environmental Studies, Permitting, and Plans, Negotiations, or Agreements

with Local Individuals or Groups

355

17.1  Environmental Study Results

355

17.2  Requirements and Plans for Waste and Tailings Disposal, Site Monitoring, and

Water Management During Operations and After Mine Closure

359

17.3  Project Permitting

362

17.4  Local Individuals and Groups

364

17.5  Mine Closure

365

17.6  Adequacy of Plans

367

17.7  Commitments to Ensure Local Procurement and Hiring

367

18  Capital and Operating Costs

368

18.1  Underground

368

18.1.1 Expected Accuracy of Cost Estimates

368

18.1.2 Capital Cost Estimates

368

18.1.3 Operating Cost Estimates

368

18.1.4 Closure and Reclamation

370

18.1.5 Infrastructure

370

18.2  Surface Operating Costs

370

18.2.1 Autoclave Capital Cost Allocation (Processing Fee)

372

18.2.2 Operating Cost Estimates

375

18.2.3 General and Administrative Costs

379

19  Economic Analysis

380

19.1  General Description

380

19.2  Results

389

19.3  Sensitivity Analysis

408

September 2026

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SEC

Technical Report Summary – Granite Creek

Page viii

20  Adjacent Properties

410

21  Other Relevant Data and Information

412

22  Interpretation and Conclusions

413

23  Recommendations

422

23.1  Recommended Work Programs

422

23.2  Recommended Work Program Costs

425

24  References

426

25  Reliance on Information Provided by the Registrant

432

26  Date and Signature Page

434

Appendices

1

Appendix A: Consent Letter

2

List of Tables

Table 1-1: Inclusive Mineral Resource Statement for Granite Creek Underground - Effective Date March 31, 2026

4

Table 1-2: Exclusive Mineral Resource Statement for Granite Creek Underground - Effective Date March 31, 2026

5

Table 1-3: Granite Creek Open Pit Mineral Resource Statement

6

Table 1-4: Underground Cut-off Grade Parameters

8

Table 1-5: Mineral Reserve Estimate for Granite Creek as of March 31, 2026 - SRK Consulting (U.S.), Inc.

9

Table 1-6: Life of Mine Capital Costs

10

Table 1-7: Life of Mine Operating Costs

10

Table 1-8: Indicative Economic Result

11

Table 2-1: Site Visit Participants

17

Table 3-1: Summary of Mineral and Surface Title Holders, Granite Creek Project

20

Table 3-2: Holding Costs, Granite Creek Project

21

Table 3-3: Owned Unpatented Claims, Granite Creek Project

21

Table 3-4: Leased Unpatented Claims, Granite Creek Project

21

Table 7-1: Salient Results of the Ogee Zone Channel Sample Assays

57

Table 7-2: Summary of Historical Drilling (pre i-80) on the Granite Creek Property Since 1970

59

Table 7-3: PMC Drilling 1970 to 1996

59

Table 7-4: PMC – Homestake Drilling 1997 to 2000

59

Table 7-5: PMC – Barrick Drilling 2003

59

Table 7-6: Atna Drilling 2004

60

Table 7-7: Atna Drilling 2005 to 2006

60

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Technical Report Summary – Granite Creek

Page ix

Table 7-8: PMC (Barrick) Drilling 2007

60

Table 7-9: PMC (Barrick) 2008 Drilling

61

Table 7-10: 2012 Atna Mag Pit Core Drilling

61

Table 7-11: 2013 to 2015 Atna Underground Development RC Drilling

61

Table 7-12: Drillholes Within the Current Property Boundary by Type and Operator

63

Table 8-1: Historical (pre-i-80) Operators, Periods, and Analytical Laboratories at Granite Creek

87

Table 8-2: Historical (pre-i-80) CRMs at Granite Creek — Certified Au Values and Insertion Counts

88

Table 8-3: i-80 Gold CRM Catalog Used at Granite Creek, 2021 to 2026 (year-to-date) — Certified Au Values and Insertion Counts

91

Table 8-4: Lower Limit of Detection (LLD) and 5xLLD Blank Acceptance Threshold by Gold Analytical Method

95

Table 8-5: Duplicate Tolerance Envelopes Applied to the i-80 Gold Drillhole Assay Database

96

Table 8-6: Annual QA/QC Insertion Summary, i-80 Gold Drillhole Sampling at Granite Creek, 2021 to 2026

100

Table 9-1: i-80 Standard Operating Procedure List

112

Table 10-1: Metallurgical Test Work Programs

121

Table 10-2: Underground Samples Batch Pressure Oxidation Conditions from FLS Program

123

Table 10-3: Granite Creek Underground Metallurgical Sample Blends for Additional Testing

124

Table 10-4: Granite Creek Underground Metallurgical Testing Program Follow-Up BTAC Test Conditions

124

Table 10-5: Granite Creek Underground Metallurgical Testing Program Continuous POX Run Test Conditions

124

Table 10-6: Underground Blend Samples Cyanide Detox Conditions from FLS Program

126

Table 10-7: Proportions of Continuous Pressure Oxidation Feed from i-80 Sources

127

Table 10-8: Particle Size Analysis of Continuous Pressure Oxidation Blend and Components

128

Table 10-9: Continuous Pressure Oxidation Test Conditions

128

Table 10-10: Continuous Pressure Oxidation Testing, Cyanide Detoxification Conditions

133

Table 10-11: Malvern Laser Diffraction Particle Size Analysis of Continuous Pressure Oxidation Samples

134

Table 10-12: Required Solids Feed Concentration Ranges

134

Table 10-13: Continuous Fill Tests on Continuous Pressure Oxidation Samples

134

Table 10-14: Slurry Rheology on Continuous Pressure Oxidation Samples

134

Table 10-15: Pressure Filtration on Cyanide Detoxification Samples

135

Table 10-16: Bond Ball Mill Grindability Test Results

135

Table 10-17: Underground Samples Head Assays from FLS Program

137

Table 10-18: Composition of Continuous Pressure Oxidation Feed Blend

138

Table 10-19: South Pacific Test Program Sample List

138

Table 10-20: Projected Feed from Granite Creek Underground

149

Table 10-21: Summary of Feed Values for Granite Creek Test Programs

149

Table 10-22: Autoclave Pre-Treatment Tests from Dawson Test Work Program

150

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Technical Report Summary – Granite Creek

Page x

Table 10-23: Underground Samples Baseline and Batch Pressure Oxidation CIL Results from FLS Program

153

Table 10-24: Granite Creek Underground Metallurgical Testing Program Follow-Up BTAC Sulfide

Oxidation Results Compared to Predicted Results

154

Table 10-25: Granite Creek Underground Metallurgical Testing Program Follow-Up BTAC Gold Recovery

Results Compared to Predicted Results

155

Table 10-26: Underground Samples (OAPC) Continuous Autoclave Tests from FLS Program

156

Table 10-27: Underground Cyanide Detox Reagent Consumption from FLS Program

157

Table 10-28: Underground Cyanide Detox WAD from FLS Program

157

Table 10-29: Pressure Oxidation Continuous Test, Sulfide Sulfur Oxidation and Carbonate Decomposition

158

Table 10-30: Bench Top Autoclave Tests on Continuous Pressure Oxidation Feed Blend

158

Table 10-31: Continuous Pressure Oxidation, CIL Au Recovery

159

Table 10-32: Bench Top Autoclave, CIL Au Recovery

159

Table 10-33: POX Results

161

Table 10-34: Baseline CIL Results

162

Table 10-35: POX Residue CIL Results

163

Table 10-36: Gold Recovery Results from Toll Milling BTAC Tests

165

Table 10-37: BTAC Gold Recovery Results on Otto Zone Samples

168

Table 11-1: Negative Values in Drillhole Database

174

Table 11-2: Open Pit Estimation Zone and Pit Name

177

Table 11-3: Open Pit Numeric Indicator Model Parameters

177

Table 11-4: Open Pit Compositing Interval Statistics

185

Table 11-5: Open Pit Compositing Comparison, 6 m (20 ft) Intervals

186

Table 11-6: Open Pit Upper Clipping (Capping) Au ppm Values by Domain

187

Table 11-7: Open Pit Domain Density Summary

187

Table 11-8: Open Pit Variogram Parameters

188

Table 11-9: Open Pit Block Model Parameters

188

Table 11-10: Open Pit ID2 Estimation Parameters

190

Table 11-11: Open Pit Combined Estimator Hierarchy

190

Table 11-12: Open Pit Comparison of Composite Values to Grade Estimation Methods

195

Table 11-13: Open Pit Mineral Resource Classification Parameters

200

Table 11-14: Open Pit Parameters for Resource Class Numeric Indicator Model

200

Table 11-15: Granite Creek Resource Parameters for Open Pit Optimization

202

Table 11-16: Granite Creek Open Pit Mineral Resource Statement

203

Table 11-17: Granite Creek Mineral Resource Sensitivity to Cut-off Grade – Reported on 100% basis

204

Table 11-18: Summary of Drilling and Sampling in Target Area

207

Table 11-19: Summary of Drilling and Sampling Information Available within Leapfrog

207

Table 11-20: Summary of Domains and Criteria Used

221

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Technical Report Summary – Granite Creek

Page xi

Table 11-21: Summary of Raw vs. Capped Composite Statistics per Domain

230

Table 11-22: Summary of Gold Variogram Parameters

231

Table 11-23: Summary of Rotated Prototype Parameters Used in Granite Creek Model

234

Table 11-24: Summary of Faults and Key Structures Used to Define Variable Orientations for Estimation

235

Table 11-25: Summary of Estimation Parameters for Gold Estimates

240

Table 11-26: Summary Statistics and Coverage of Raw Sampling per Metallurgical Parameter vs. Gold Coverage

242

Table 11-27: Summary Statistics of Capped Composites (5 ft), Showing Capping (Maximum) Levels per Metallurgical Elements

243

Table 11-28: Summary of Variogram Parameters Used for Metallurgical Elements

245

Table 11-29: Summary of Search Parameters Used for AuCN

246

Table 11-30: Summary Estimation Parameters (CO3 %)

247

Table 11-31: Summary Estimation Parameters (Preg Rob %)

248

Table 11-32: Summary Estimation Parameters (Total Organic Carbon, TOC %)

249

Table 11-33: Summary Estimation Parameters (Sulfide Sulfur %)

250

Table 11-34: Density Values Used in the Underground Model

251

Table 11-35: Summary Statistics of Raw and Declustered Composite Grades vs. Block Estimates and NN-Assigned Values (0 g/t Cut-off)

259

Table 11-36: Summary of Mineral Resource Cut-off and MSO Assumptions

269

Table 11-37: Inclusive Mineral Resource Statement for Granite Creek Underground - Effective Date March 31, 2026

270

Table 11-38: Exclusive Mineral Resource Statement for Granite Creek Underground - Effective Date March 31, 2026

271

Table 11-39: Grade Sensitivity (Combined Measured and Indicated), inside MSO on a 100% Project basis

272

Table 11-40: Grade Sensitivity Table (Inferred), inside MSO on a 100% Project Basis

273

Table 11-41: Open Pit Mineral Resource Sensitivity by Domain

277

Table 11-42: Underground Mineral Resource Sensitivity (1 of 2)

280

Table 11-43: Underground Mineral Resource Sensitivity (2 of 2)

281

Table 12-1: Mineral Reserve Estimate for Granite Creek Underground – Effective Date March 31, 2026

284

Table 12-2: Underground Cut-off Grade Parameters

285

Table 13-1: Geotechnical Domains

289

Table 13-2: Geotechnical Parameters for the Three Geotechnical Domains

289

Table 13-3: RMR Summary for Each Geotechnical Domain

290

Table 13-4: Major Joint Sets Defined by Televiewer Data Analysis in Dips Software

290

Table 13-5: Stress Analysis from Assumed Depth and k Value of 0.75

291

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Technical Report Summary – Granite Creek

Page xii

Table 13-6: Ground Support Recommendations for Granite Creek SPZ Development and Mining

291

Table 13-7: Recommended Operating Practices for Ground Support Installation

292

Table 13-8: Site Visit Temperature and Humidity Data

295

Table 13-9: MSHA Standards

297

Table 13-10: Capital Cost Estimate

300

Table 13-11: Annual Operating Cost Estimate

300

Table 13-12: Productivity Rates

305

Table 13-13: Schedule Parameters for Underground Mining

306

Table 13-14: Granite Creek Annual Production Schedule

307

Table 13-15: Mobile Equipment Fleet

310

Table 13-16: Granite Creek Manpower

311

Table 14-1: Summary of Key Process Statistics

316

Table 14-2: Design Scenarios for Lone Tree Autoclave Feed

316

Table 14-3: Lone Tree Facility Water Consumption by Type

325

Table 14-4: Lone Tree Facility Energy Usage by Area

326

Table 15-1: Granite Creek Active Dewatering Wells

333

Table 15-2: Summary of Locations, Construction Information and Water Levels for Dewatering Wells, Monitoring Wells, and Piezometers

336

Table 15-3: Projected Electrical Demand by Area

343

Table 15-4: WRD Capacity

348

Table 17-1: MWMP Results of Rock Placed in CX Pit 2005 to 2022

356

Table 17-2: MWMP Results of Rock Placed in CX Pit 2023 to 2025

357

Table 17-3: Water Quality April 2023 to January 2025

357

Table 17-4: Granite Creek Mine Project Permits

363

Table 18-1: Estimated Mining Capital Cost

368

Table 18-2: Mining Operating Costs

369

Table 18-3: Infrastructure Capital Cost Estimates1

370

Table 18-4: Infrastructure Unit Operating Cost Estimates

371

Table 18-5: Annual Infrastructure Operating Costs (US$000’s)

371

Table 18-6: Unit Costs for Over-the-Road Ore Haulage

372

Table 18-7:Capital Cost Summary

373

Table 18-8: Sustaining CAPEX Assumptions for Lone Tree

374

Table 18-9: Lone Tree Processing Facility Operating Cost Summary

375

Table 18-10: Lone Tree Pressure Oxidation-CIL Power Costs

375

Table 18-11: Lone Tree Pressure Oxidation-CIL Labor Costs

377

Table 18-12: Lone Tree Pressure Oxidation-CIL Maintenance Costs.

377

Table 18-13: Lone Tree Pressure Oxidation-CIL Consumable Costs

377

Table 18-14: Key Consumables Costs for Lone Tree Pressure Oxidation-CIL

378

Table 18-15: Lone Tree Pressure Oxidation-CIL Facility Costs

378

Table 18-16: Lone Tree Processing Facility Operating Cost Summary

378

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SEC

Technical Report Summary – Granite Creek

Page xiii

Table 18-17: Average Annual G&A Spend

379

Table 19-1: Basic Model Parameters

380

Table 19-2: Net Proceeds of Mineral Tax Sliding Scale

381

Table 19-3: Life-of-Mine Production Summary

383

Table 19-4: Granite Creek Processing Summary

384

Table 19-5: Granite Creek Mining Cost Summary

386

Table 19-6: Granite Creek Processing Costs

387

Table 19-7: Granite Creek Surface Costs

387

Table 19-8: Granite Creek Power Costs

387

Table 19-9: Granite Creek G&A Costs

387

Table 19-10: Modeled Sustaining Capital

388

Table 19-11: Indicative Economic Results

389

Table 19-12: Economic Results – Tabular Data

390

Table 23-1: Summary of Costs for Open Pit and Underground Recommended Work Programs

425

Table 25-1: Reliance on Information Provided by the Registrant

433

List of Figures

Figure 1-1: Annual Cash Flow Summary

12

Figure 3-1: Granite Creek Project Location

18

Figure 3-2: Land Position Map, Granite Creek Project

19

Figure 6-1: Regional Geology Map

33

Figure 6-2: Local Geology Map

35

Figure 6-3: Property Geology Map

40

Figure 6-4: Alteration of the Mag Pit

41

Figure 6-5: Granite Creek Stratigraphic Column

43

Figure 6-6: Cross-section A-A’ looking Northeast showing Structure, Lithology and Mineralization

44

Figure 7-1: Gravity Survey, 2,587 Stations, Magee Geophysical Services, 2006

46

Figure 7-2: Location of the MT Survey Lines on the Geology and Pit Locations (Left) and on the Residual Gravity (Right)

48

Figure 7-3: Pinson Local Gravity Interpretation

50

Figure 7-4: MT Resistivity Depth Inversion for Line 6090

51

Figure 7-5: MT Resistivity Depth Inversion for Line 12300

52

Figure 7-6: MT Resistivity Depth Inversion for Line 13860

53

Figure 7-7: MT Resistivity Depth Inversion for Line 15300

54

Figure 7-8: MT Resistivity Depth Inversion for Line 17160

55

Figure 7-9: MT Resistivity Depth Inversion for Line 19230

56

Figure 7-10: Granite Creek Project Drill Plan by Operator

58

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Technical Report Summary – Granite Creek

Page xiv

Figure 7-11: Plan View Section Lines of Granite Creek Mine Project

62

Figure 7-12: Drilling by Hole Type Completed by PMC

64

Figure 7-13: Drilling by Type Completed by PMC with Barrick as Operator

65

Figure 7-14: Drilling Completed by Atna

66

Figure 7-15: Drilling Completed by i-80

67

Figure 7-16: Vertical Section A-A’ of the Mag Pit Area

69

Figure 7-17: Vertical Section B-B1 of the Pit CX and C Area

70

Figure 7-18: Vertical Section C-C1 of the Pit A Area

71

Figure 7-19: Vertical Section D-D1 of the Pit B Area

72

Figure 7-20: Vertical Section E-E1 of the Underground Resource Area

73

Figure 7-21: Plan View Showing Section Locations through the Underground Resource Area

74

Figure 7-22: Section A-A’ Showing Drilling in the CX Zone, 30.5 m (100 ft) thick, looking North

75

Figure 7-23: Section B-B’ Showing Drilling in the Otto and Ogee Zones, looking North

76

Figure 7-24: Section C-C Showing Drilling in the SPZ, 130 m (100 ft) thick, looking North

77

Figure 7-25: Well Locations

79

Figure 7-26: Timeline for Hydrogeologic Characterization with Relationship to Mining

80

Figure 7-27: Predicted Passive Inflow and Dewatering Well Pumping

82

Figure 7-28: Predicted Groundwater Elevation in the Vicinity of Ore Bodies: Otto, Ogee, and South Pacific

83

Figure 7-29: Predictive and Passive Inflows from Scenarios One and Two

84

Figure 8-1: Historical Rocklabs CRM Control Chart (OxD57, Representative)

89

Figure 8-2: Historical Blank Control Chart, 2005 to 2018

89

Figure 8-3: Historical Combined Duplicate Scatter, 2005 to 2018

90

Figure 8-4: CDN-GS-7J Control Chart, i-80 Era

92

Figure 8-5: CDN-GS-30C Control Chart, i-80 Era

92

Figure 8-6: CDN-GS-P6E Control Chart, i-80 Era

93

Figure 8-7: OREAS-277 Control Chart, i-80 Era

93

Figure 8-8: OREAS-279 Control Chart, i-80 Era

94

Figure 8-9: OREAS-282 Control Chart, i-80 Era

94

Figure 8-10: i-80 Era Blank Control Chart, 2021 to Q1 2026

95

Figure 8-11: Field (Quarter-core) Duplicate Scatter, i-80 Era

97

Figure 8-12: Preparation (Coarse Reject) Duplicate Scatter, i-80 Era

98

Figure 8-13: Lab (Pulp / Analytical) Duplicate Scatter, i-80 Era

99

Figure 8-14: Check Sample Report Primary vs. Check Assay Scatter, Au, 2023 to 2026

102

Figure 9-1: Underground Working Heading at Granite Creek Showing Mineralized Structure with Sample

Outline and Drill Pattern Marks Observed During the QP Site Visit

106

Figure 9-2: Underground Working Heading at Granite Creek Showing Exposed Mineralized Face Observed During the QP Site Visit

107

Figure 9-3: Drill Core from Hole iGS22-07 (1,499 to 1,482 m (581 ft)) with Printed Assay Sheet

Used by the QP to Cross-Check Mineralization Against Database-Recorded Au and Ag Grades

109

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Figure 9-4: Drill Core from Hole iGS23-05 (1,498 to 1,491 m (610 ft)) with Printed Assay Sheet

Used by QP to Cross-Check Mineralization Against Database Au and Ag Grades

110

Figure 9-5: Diamond Core Saw Station at the Lone Tree Laboratory Facility

113

Figure 9-6: Palleted and Tarped Coarse Reject and Core Storage Yard at the Lone Tree Facility

114

Figure 9-7: Pulp Sample Storage at the Lone Tree Facility

115

Figure 9-8: Specific Gravity (Water Immersion) Measurement Station at the Lone Tree Laboratory Facility

117

Figure 10-1: Granite Creek POX Pilot Plant Sulfide Oxidation Profile

125

Figure 10-2: Continuous Pressure Oxidation Program Flow Sheet

127

Figure 10-3: Continuous Pressure Oxidation Autoclave at FLSmidth

129

Figure 10-4: Continuous Pressure Oxidation Slurry Temperature Profile

130

Figure 10-5: Continuous Pressure Oxidation Vessel Pressure Profile

130

Figure 10-6: Continuous Pressure Oxidation Vent Oxygen Purity

131

Figure 10-7: Continuous Pressure Oxidation Calculated Feed Flow Rate

131

Figure 10-8: Continuous Pressure Oxidation Sulfide Oxidation Profile

132

Figure 10-9: CNWAD Concentration in Cyanide Detoxification

133

Figure 10-10: SGS Metallurgical Program Head Assays

139

Figure 10-11: Bulk Modal (QEMSCAN) Results

141

Figure 10-12: Pyrite/Marcasite Liberation by Mass

142

Figure 10-13: Arsenopyrite Liberation by Mass

142

Figure 10-14: Silicate Liberation by Mass

143

Figure 10-15: Carbonate Liberation by Mass

144

Figure 10-16: Locations of Metallurgical Samples (Section View)

145

Figure 10-17: Locations of Metallurgical Samples (Plan View)

146

Figure 10-18: Distribution of Gold Grade in Metallurgical Samples

147

Figure 10-19: Distribution of Organic Carbon in Metallurgical Samples

147

Figure 10-20: Distribution of Carbonate in Metallurgical Samples

148

Figure 10-21: Distribution of Sulfide Sulfur in Metallurgical Samples

148

Figure 10-22: Gold Cyanide Solubility and Sulfide Influence – Ogee Samples

151

Figure 10-23: CIL Gold Recovery as a Function of Sulfide Sulfur Oxidation – Underground Samples

154

Figure 10-24: Neutralization Lime Usage in Continuous Pressure Oxidation Discharge

159

Figure 10-25: South Pacific Gold Recovery as a Function of Sulfide Oxidation (SGS 2026)

164

Figure 10-26: CIL Recovery vs Cyanide Soluble Gold Percent

166

Figure 10-27: CIL Recovery vs Feed Sulfide Sulfur Content

166

Figure 10-28: OG Zone Gold Recovery vs Gold Head Grade

167

Figure 10-29: OG Zone Gold Recovery vs Organic Carbon Content

168

Figure 10-30: Sulfide Oxidation vs Carbonate Content, 45-Minute Retention Time, Acidic Conditions

169

Figure 10-31: Mass Loss versus Carbonate Content, 45-Minute Retention Time, Acidic Conditions

169

Figure 10-32: CIL Gold Recovery vs CIL Feed Sulfide Sulfur, All Conditions

170

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Figure 11-1: Drillhole Used Plan View on Topography

173

Figure 11-2: Current Topography Used for Open Pit Resource Estimation

175

Figure 11-3: Open Pit Geologic Model, Oblique View

176

Figure 11-4: Open Pit Estimation Zones

177

Figure 11-5: Example of Numeric Indicator High-Grade Trend Analysis, Mag Pit

178

Figure 11-6: Open Pit Zone 3 Sub-Domains

179

Figure 11-7: High-Grade and Low-Grade Open Pit Domains in the CX Fault

180

Figure 11-8: Box and Whisker Plot of Open Pit Estimation Domains

181

Figure 11-9: Histograms of the High-Grade and Low-Grade Distributions in Zones 1 and 2

182

Figure 11-10: Histograms of the High-Grade and Low-Grade Distributions in Zones 3 and 4

183

Figure 11-11: Open Pit Sample Interval Length Statistics of Au ppm Assays

184

Figure 11-12: Open Pit Compositing Comparison, 6 m (20 ft) Intervals

185

Figure 11-13: Example of Open Pit Cumulative Log Probability Plot, Zone 1 HG

186

Figure 11-14: Open Pit Numeric Indicator Models

189

Figure 11-15: Open Pit Zone 1 Visual Comparison Composite to Block Model Grade, Plan View

191

Figure 11-16: Open Pit Zone 2 Visual Comparison Composite to Block Model Grade, Plan View

192

Figure 11-17: Open Pit Zone 3 Visual Comparison Composite to Block Model Grade Plan View

192

Figure 11-18: Open Pit Zone 4 Visual Comparison Composite to Block Model Grade Plan View

193

Figure 11-19: Open Pit Zone 1 Section Composites and Block Model Cross Sections

193

Figure 11-20: Open Pit Zone 2 Section Composites and Block Model Cross Section

194

Figure 11-21: Open Pit Zone 3 Section Composites and Block Model Cross Section

194

Figure 11-22: Open Pit Zone 4 Section Composites and Block Model Cross Section

195

Figure 11-23: Cumulative Frequency of Composite and Estimated Block Grades

196

Figure 11-24: X Axis, Zone 1 High-Grade Domain

197

Figure 11-25: Open Pit Swath Plot, Y Axis, Zone 1 High-Grade Domain X axis, Zone 1 High-Grade Domain

198

Figure 11-26: Open Pit Swath Plot Z Axis, Zone 1 High-Grade Domain

199

Figure 11-27: Open Pit Constrained Resource Class All Areas Plan View

200

Figure 11-28: Plan Showing Collar Locations and Borehole Traces within the MRE-Focused Dataset

208

Figure 11-29: Isometric View Showing Open Pit Depletion (Surface) and Underground Workings (Grey)

209

Figure 11-30: Plan Showing Main Defined Mineralization Domain Boundary Areas at the Granite Creek Underground Operation

212

Figure 11-31: Plan Showing Main Defined Mineralization Areas at the Granite Creek Underground Operation

213

Figure 11-32: Long Section (Looking Northwest) Showing Main Defined Mineralization Areas at the Granite Creek Underground Operation

214

Figure 11-33: Example of EDA Analysis in X-10 to Identify Potential Changes in Grade Populations within the Ogee Domain

215

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Figure 11-34: Example of Log-Probability Plots Showing Polynomial Model Fitting (Pink) and Potential Trend Breaks (Yellow)

216

Figure 11-35: Example of Sensitivity Analysis on ISO Values Completed by SRK for the Ogee Domain

218

Figure 11-36: Level Plan 4230 Showing Selected Ogee and Otto Domain Boundaries vs. Sampling (Including Grade Control Muck Sampling)

219

Figure 11-37: SRK Domain Model Used for Grade Estimation

222

Figure 11-38: Box-Whisker Plot Showing Gold Grade Distributions within the SRK Domained Models, showing mean grades.

223

Figure 11-39: Log Probability Plot Showing Sampling Length in Diamond and RC Sampling at Granite Creek

224

Figure 11-40: Summary of Sampling Lengths Inside the High-Grade Domains within Diamond and RC Drillholes

225

Figure 11-41: Examples of X-10 Capping Analysis Showing Capped Values in Yellow

227

Figure 11-42: Capping Analysis Showing Breaks in Data Population (Ogee HG Domain)

228

Figure 11-43: Capping Analysis Showing Breaks in Data Population (Otto HG Domain)

228

Figure 11-44: Capping Analysis Showing Breaks in Data Population (Ogee SPZ Domain)

229

Figure 11-45: Modeled Normal Score Downhole Variograms and Directional Correlograms for OgeeHG (Gold)

232

Figure 11-46: Modeled Normal Score Downhole and Directional Variograms for OttoHG (Gold)

232

Figure 11-47: Modeled Normal Score Downhole and Directional Variograms for SPZHG (Gold)

233

Figure 11-48: Block Size Analysis, Ogee HG Domain

236

Figure 11-49: KNA Analysis on Number of Samples vs. Slope of Regression (Top) and Impact of Negative Weights (Bottom) — Ogee HG Domain

237

Figure 11-50: KNA Analysis on Number of Samples vs. Slope of Regression (Top) and Impact of Negative Weights (Bottom) — SPZ HG Domain

238

Figure 11-51: Plan Section Showing Block Grades vs. Composites (Ogee)

252

Figure 11-52: Cross Section Showing Block Grades vs. Composites (Ogee)

253

Figure 11-53: Level Plan (Map) Showing Block Grades vs. Composites (Otto)

254

Figure 11-54: Cross Section Showing Block Grades vs. Composites (Otto)

255

Figure 11-55: Plan Showing Block Grades vs. Composites (SPZ)

256

Figure 11-56: Cross Section Showing Block Grades vs. Composites (SPZ)

257

Figure 11-57: Swath Plot Analysis of Au (g/t), Ogee HG Domain

261

Figure 11-58: Swath Plot Analysis of Au (g/t), Otto HG Domain

262

Figure 11-59: Swath Plot Analysis of Au (g/t), SPZ HG Domain

263

Figure 11-60: Long Section Showing Final Measured (red/brown) and Indicated Shapes (green) Selected for Granite Creek

265

Figure 11-61: Cross Section Showing Final Measured and Indicated Shapes Selected for Granite Creek (Ogee and Otto Domains)

266

Figure 11-62: Cross Section Showing Final Measured and Indicated Shapes Selected for Granite Creek (SPZ Domains)

267

Figure 11-63: Final Classification Shown Against Filtered Mineral Resource Blocks

268

Figure 12-1: General Locations of the Granite Creek Deposit

282

Figure 13-1: Contact Water Measurements

294

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Figure 13-2: Future Mine Airflow Configuration

296

Figure 13-3: Surface Fan Installation General Arrangement

299

Figure 13-4: Production Design with MSO Guidance

301

Figure 13-5: Schematic of Cut Mining Sequence

302

Figure 13-6: Production Mining Sequence

303

Figure 13-7: Underground Infrastructure Location

304

Figure 13-8: Granite Creek Mine Design Looking Northwest

305

Figure 13-9: Granite Creek Annual Production Schedule

307

Figure 13-10: Final Mine Outline Map

312

Figure 14-1: Overall Process Flow Diagram for the Lone Tree Facility

315

Figure 14-2: Ore Stockpile Area Layout Diagram

317

Figure 14-3: Third-Party Pressure Oxidation Facility Flow Diagram

327

Figure 15-1: Granite Creek Overview

330

Figure 15-2: Granite Creek Administration Area

331

Figure 15-3: Map of Dewatering Wells, Monitoring Wells, Piezometers at Granite Creek

335

Figure 15-4: Granite Creek Weekly Average Passive Infiltration

338

Figure 15-5: Underground Contact Water Collection and Discharge System

339

Figure 15-6: Map of Current and Planned RIBs at Granite Creek

340

Figure 15-7: Existing Water Treatment Plant (800 gpm capacity)

341

Figure 15-8: WTP-2 Process Flow

342

Figure 15-9: Granite Creek Electrical Distribution and Dewatering System

344

Figure 15-10: Granite Creek Portal Facilities

345

Figure 15-11: Screen Plant Area General Arrangement (OMC 2026)

347

Figure 15-12: C (Blue) and CX (Green) Pit WRD Plan

349

Figure 16-1: Average Annual Demand for Gold – by Sector

352

Figure 16-2: Daily Spot Gold Price

352

Figure 18-1: Yearly Mine Operating Cost Profile

369

Figure 18-2: Organizational Chart for Lone Tree Pressure Oxidation-CIL

376

Figure 19-1: Granite Creek Mining Profile (imperial)

382

Figure 19-2: Granite Creek Mining Profile (metric)

382

Figure 19-3: Granite Creek Processing Profile

384

Figure 19-4: LoM Operating Cost Summary

385

Figure 19-5: LoM Operating Cost Contributions

385

Figure 19-6: Granite Creek Sustaining Capital Profile

388

Figure 19-7: Annual Cash Flow Summary

392

Figure 19-8: Sensitivity Analysis

392

Figure 20-1: Summary of Gold Production at Turquoise Ridge (as reported by Barrick 2024)

393

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Figure 20-2: Barrick Mineral Resource Statement for Turquoise Ridge (effective date December 31, 2023) – reported on 100% basis

394

Figure 20-3: Barrick Mineral Reserve Statement for Turquoise Ridge (effective date December 31, 2023) – reported on 100% basis

394

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List of Abbreviations

The metric system has been used throughout this report. Tonnes are metric of 1,000 kg, or 2,204.6 lb. All currency is in U.S. dollars (US$)

unless otherwise stated.

Abbreviation

Unit or Term

A

ampere

AA

atomic absorption

A/m2

amperes per square meter

ANFO

ammonium nitrate fuel oil

Ag

silver

Au

gold

AuEq

gold equivalent grade

°C

degrees Centigrade

CCD

counter-current decantation

CIL

carbon-in-leach

CoG

cut-off grade

cm

centimeter

cm2

square centimeter

cm3

cubic centimeter

cfm

cubic feet per minute

ConfC

confidence code

CRec

core recovery

CSS

closed-side setting

CTW

calculated true width

°

degree (degrees)

dia.

diameter

EIS

Environmental Impact Statement

EMP

Environmental Management Plan

FA

fire assay

ft

foot (feet)

ft2

square foot (feet)

ft3

cubic foot (feet)

g

gram

gal

gallon

g/L

gram per liter

g-mol

gram-mole

gpm

gallons per minute

g/t

grams per tonne

ha

hectares

HDPE

Height Density Polyethylene

hp

horsepower

HTW

horizontal true width

ICP

induced couple plasma

ID2

inverse-distance squared

ID3

inverse-distance cubed

IFC

International Finance Corporation

ILS

Intermediate Leach Solution

kA

kiloamperes

kg

kilograms

km

kilometer

km2

square kilometer

koz

thousand troy ounce

kt

thousand tonnes

kt/d

thousand tonnes per day

kt/y

thousand tonnes per year

kV

kilovolt

kW

kilowatt

kWh

kilowatt-hour

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Abbreviation

Unit or Term

kWh/t

kilowatt-hour per metric tonne

L

liter

L/sec

liters per second

L/sec/m

liters per second per meter

lb

pound

LHD

Long-Haul Dump truck

LLDDP

Linear Low Density Polyethylene Plastic

LOI

Loss On Ignition

LoM

Life-of-Mine

m

meter

m2

square meter

m3

cubic meter

masl

meters above sea level

MARN

Ministry of the Environment and Natural Resources

MDA

Mine Development Associates

mg/L

milligrams/liter

mm

millimeter

mm2

square millimeter

mm3

cubic millimeter

MME

Mine & Mill Engineering

Moz

million troy ounces

Mt

million tonnes

MTW

measured true width

MW

million watts

m.y.

million years

NGO

non-governmental organization

NI 43-101

Canadian National Instrument 43-101

OSC

Ontario Securities Commission

oz

troy ounce

%

percent

PLC

Programmable Logic Controller

PLS

Pregnant Leach Solution

PMF

probable maximum flood

ppb

parts per billion

ppm

parts per million

QA/QC

Quality Assurance/Quality Control

RC

rotary circulation drilling

RoM

Run-of-Mine

RQD

Rock Quality Description

SEC

U.S. Securities & Exchange Commission

sec

second

SG

specific gravity

SPT

standard penetration testing

st

short ton (2,000 pounds)

t

tonne (metric ton) (2,204.6 pounds)

t/h

tonnes per hour

t/d

tonnes per day

t/y

tonnes per year

TSF

tailings storage facility

TSP

total suspended particulates

µm

micron or microns

V

volts

VFD

variable frequency drive

W

watt

XRD

x-ray diffraction

y

year

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1

Executive Summary

This report was prepared as a Pre-Feasibility Study-level Technical Report Summary in accordance with the Securities and Exchange Commission

(SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for i-80 Gold Corp. (“i-80” or the

“Registrant”) by SRK Consulting (U.S.), Inc. (SRK) on the Granite Creek Underground Project (the “ Project “).

1.1

Property Description (Including Mineral Rights) and Ownership

The Granite Creek Project is located in Humboldt County, Nevada, 45 kilometers (km) (28 miles (mi)) northeast of the town of Winnemucca, and it

is part of the historic Potosi mining district. It is centered at roughly 41° 8’ N latitude and 117° 15.5’ W longitude. It encompasses about 4,506 acres (1,823.5 hectares (ha)) including owned unpatented claims, leased unpatented

claims and owned surface fee land. i-80 Gold purchased the Granite Creek property from Waterton Global in June 2020.

1.2

Geology and Mineralization

The Property is located on the eastern flank of the Osgood Mountains within the Basin and Range tectonic province of northern Nevada. The

Granite Creek Mine occurs within a northeast-trending structural corridor known as the Getchell gold trend. This trend also encompasses a number of gold deposits located outside the Property including the Preble, Getchell, Turquoise Ridge, and Twin

Creeks. These deposits are hosted in Paleozoic marine sedimentary rocks. Gold mineralization at the Property is described as a Carlin-type, sedimentary-rock hosted system.

The Property geology comprises a sequence of Cambrian to Ordovician sedimentary rocks that form part of the Osgood Mountain Terrane and the

Osgood Mountains. Much of the Property comprises shales, hornfels sedimentary rocks and limestone interbeds of the Preble Formation, and an overlying (or juxtaposed), alternating sequence of limestone, shale, and dolomite with tuffaceous shale and

intraformational conglomerates belonging to the Comus Formation. The Preble and Comus Formations have been folded into a broad north-plunging anticline and have been intruded by a large Cretaceous granodiorite stock, resulting in irregular contact

metamorphism.

Gold mineralization at the Property is strongly structurally controlled, occurring at favorable sites within a fault network

occurring around the eastern edge of the Osgood granodiorite and predominantly within Comus Formation host rocks. Mineralization is commonly associated with the decarbonatization of carbonate rocks and the introduction of silica, fine grained

pyrite, arsenian pyrite, and remobilized carbon. Continuity of mineralization is highly variable, ranging from 40 to 4,500 feet (ft) (12 to 1,372 meters (m)) in strike extent, 250 to 1,800 ft (76 to 550 m (1,804 ft)) in down-dip extent and 5 to 400 ft (1.5 to 122 m (400 ft)) in thickness. The underground mineralization has a variable thickness between 5 and 30 ft (1.5 and 9 m (30 ft)).

Oxidation reaches depths of up to 550 m (1,800 ft) within fault zones. Oxide mineralization includes pervasive limonite, hematite, along with

other iron and arsenic oxides. Historical production from the open pits was focused on oxidized material.

Underground mineralization

displays pervasive argillization and decarbonatization of host lithologies, along with the formation of dissolution collapse breccias and intense shearing. Where the alteration is strongest, the altered zones consist of punky, spongy decarbonized

limestone in an argillically altered fine-grained, carbon-rich matrix (Gustavson, 2012). Silicification is minor and occurs as a broad overprint on the zone. Underground production includes both sulfide and oxide material.

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1.3

Status of Exploration, Development and Operations

Exploration at Granite Creek since the 1970’s has combined geologic mapping, geochemical sampling, and geophysical surveys (gravity,

magnetics, and IP/EM/MT/CSAMT), contributing to the discovery of several open pit gold deposits at the Property.

Since acquiring Granite

Creek in 2021, i-80 has not conducted new surface geophysical or geochemical programs but has focused the exploration on drilling known targets with the aim to develop Mineral Resources and reserves to support

mining operations. The Project is currently in operation and an underground exploration and development program targeting the South Pacific Zone (SPZ) at depth was completed during 2025, comprising approximately 174 m (570 ft) of development

drifting and 16,531 m (54,236 ft) of drilling.

Since 1970, a total of 2,083 drillholes totaling 291,312 m (955,748 ft) have been completed

at the Property by successive operators (Pinson Mining Company (PMC) and predecessors, Homestake, Barrick, and Atna).

Drilling completed

by i-80 has expanded substantially since the Granite Creek Initial Assessment (IA) Technical Report (March 2025). The 2021 and 2022 programs, approximately 69 and 154 holes, respectively, were reviewed in

whole or in part by the previous QP as part of that IA as they focused on the open pit potential for the Project.

i-80 increased the exploration drilling materially between 2023 – 2025 with approximately:

•

In 2023 a total of 278 holes were completed which included a deep SPZ infill program.

•

In 2024 approximately 185 holes were added to continue the SPZ infill campaign; and

•

In 2025 there was an increase in the number of holes to approximately 379 holes with the introduction of

underground production drilling alongside ongoing surface drilling. Drilling continued into Q1 2026 as the tail of the SPZ infill campaign and the start of new exploration drilling targeting the Ogee, Otto, Adam Peak, and Range Front fault systems.

In total, i-80 Gold has drilled 1,083 holes within the current property

boundary; 438 core holes for 93,552 m (306,929 ft) and 645 RC holes for 30,252 m (99,252 ft), for a combined total of 123,804 m (406,181 ft) combined. To date the drilling has been concentrated on the underground Ogee, Otto, and SPZ mineralization.

All holes have been drilled and logged using standard diamond core methods with downhole surveying and structural/geotechnical logging

captured directly in the acQuire database. Core handling follows a documented chain of custody from the drill site through sample preparation, with sample security procedures maintained throughout. Sample preparation and analysis for the i-80 Gold program has been performed by independent, accredited commercial laboratories (ALS Minerals (ALS), and MSA Labs (MSA) from 2025), with primary gold analysis transitioning from conventional fire assay to

Chrysos Photon Assay in 2025.

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1.4

Mineral Resource and Mineral Reserve Estimates

1.4.1

Underground

SRK completed the underground MRE using an updated geological and mineralization model built in Leapfrog® Edge, based on i-80’s drillhole database (cut-off date February 17, 2026) restricted to the area around

current underground operations. Three principal mineralized areas are recognized which include Ogee, Otto, and the SPZ, each split into high- and low-grade sub-domains

together with a separate South Pacific Zone hanging-wall domain (SPZhw), all controlled by a system of faults and structures including the CX West Fault, the Range Front Fault, and domain boundaries A, B, and C.

Grade estimation used 2 m (5 ft) composites, domain-specific high-grade capping (typically reducing mean grades by 2.5% to 5.0 %, with

more aggressive capping in the lower-grade Ogee domain), and variogram models developed for each domain. Grades were interpolated using Ordinary Kriging and Inverse Distance Weighting with variable search orientations tied to local fault geometry,

within a block model using 15 ft x 15 ft x 15 ft parent blocks.

The classification is based on standards as defined by the Society for

Mining, Metallurgy and Exploration (SME). The Mineral Resources at the Project have been classified as Measured, Indicated and Inferred based on drillhole spacing, composite and hole counts, and geostatistical confidence in each domain.

To determine the reasonable prospects for economic extraction (RPEE), SRK has used the following key assumptions for costing, and metallurgical

recoveries. Mining costs via underground cut and fill are estimated at US$/t 190.18, with processing costs estimated at US$/t 131.13 for the autoclave processing, and US$/t 59.94 for the oxide CIL processing. Shipping costs of US$16.04 and G&A

costs of US$/t 22.29 have been used for both material types.

Metallurgical recoveries are based on the outcomes of both the historical and

the most recent metallurgical test work completed by i-80. To define the limits for processing route selection and estimated recoveries, SRK estimated AuCn (ppm), TOC% (percent), PregRob% (percent), and

Sulfide Sulfur % (percent), which were reviewed and used by the mining team in relation to the metallurgical recovery. SRK worked with i-80 to review the findings of the metallurgical studies which indicated

that the metallurgical recoveries range between 77.6% to 93.4% (average 87.6%) for the autoclave and 60.0% to 87.4% (average 68.9%) for the oxide material.

SRK has used an MSO, using Deswik Mining Software to determine the limits for the underground RPEE and were reported within mineable shapes

generated using a mineable stope optimizer (MSO). For reporting purposes given the proximity and similar mineralization the SPZhw has been combined with the SPZ MSO domains for the final reporting.

The gold price selected for Mineral Resources has been based on review of market consensus forecast data through a database which SRK

subscribe, and review by the QP of industry standards, i-80 and peers considering a time period of 10 years. It is the QP’s opinion that this is a reasonable forecast for the time considered and the

current time of the mine being in operation (ramping up production), to cover short term pricing and future long term pricing. The point of reference for the reporting of mineral resources is the in-situ

diluted tonnage and grades contained with the defined MSO limits.

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The underground Mineral Resource Statement, effective March 31, 2026, is summarized

below (Measured & Indicated by domain, plus separately reported Inferred), with tonnage and metal presented on both a 100% Project basis and the Company Attributable basis.

Table 1-1: Inclusive Mineral Resource Statement for Granite Creek Underground - Effective Date

March 31, 2026

Classification

Domain

Mass

(000’s

tonnes)

Mass

Attributable

(000’s

tonnes)

Au

Grade

(g/t)

Material

Content Au

(000’s t. oz)

Attributable

Material

Content Au

(000’s t. oz)

Measured

Ogee

231

231

7.88

58.6

58.5

Otto

310

310

6.69

66.6

66.6

SPZ

65

65

9.85

20.7

20.7

Total

606

606

7.49

145.9

145.8

Indicated

Ogee

322

321

7.06

73.0

72.9

Otto

746

746

6.26

150.2

150.2

SPZ

2,055

2,055

7.42

490.4

490.4

Total

3,122

3,122

7.11

713.6

713.5

M&I

Ogee

553

552

7.40

131.6

131.4

Otto

1,056

1,056

6.39

216.8

216.8

SPZ

2,120

2,120

7.50

511.1

511.1

Total

3,729

3,728

7.17

859.5

859.3

Inferred

Ogee

39

38

6.70

8.3

8.3

Otto

425

425

7.63

104.4

104.4

SPZ

429

418

6.53

90.1

87.4

Total

893

882

7.06

202.8

200.1

Source: SRK, 2026

Notes:

1.

Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of

Mineral Resources may be materially affected by environmental, permitting, legal, title, market or other relevant issues. The deposit has been classified as Measured, Indicated and Inferred based on confidence in the geological model and drill

spacing. The quantity and grade of reported Inferred resources are uncertain in nature, and there has not been sufficient work to define these Inferred Mineral Resources as Indicated or Measured resources. There is no certainty that any part of a

mineral resource will ever be converted into reserves.

2.

Mineral Resources are reported on an inclusive basis with units reported in Metric Tonnes, grade is reported in

grams/tonnes, and metal in 000’s troy ounces.

3.

Mineral Resources are reported within Mineable Stopes which have been defined based on metal price

assumptions,* variable metallurgical recovery assumptions, mining costs, processing costs, shipping, G&A costs. Metal Pricing is based on Gold (US$3,000 /oz). Other key assumptions include:

*

Mining Costs (US$/t 190.18);

*

Autoclave Processing Cost (US$/t 131.13), Low-grade Oxide (US$/t 59.94)

*

Autoclave Recoveries ranging 77.6% - 93.4% (average 87.6%) and Oxide recoveries ranging 60.0% - 87.4% (average

68.9%);

*

Shipping Costs (US$/t 16.04) and G&A Costs (US$/t 22.29);

*

Totaling US$/t 281.45 – 359.64 for underground mining and Processing;

*

Based on the variable recoveries an average cut-off grade of 3.75 -

9.81 g/t (0.286 oz/st) has been used reporting.

4.

Mineral Resources have been depleted for previous underground mining activity, and reported on an attributable

basis

5.

Mineral Resources are based on validated data, which have been subjected to QA/QC analysis, using capped,

composited samples at 2 m (5ft) intervals. Estimation has been completed using a combination of OK and IDW estimation methodologies and classified based on confidence in the underlying data and drill spacing. Mineral resource tonnages have been

rounded to reflect the precision of the estimate.

6.

The mineral resources were estimated by SRK Consulting (U.S.), Inc.

7.

Rounding of some figures may lead to minor discrepancies in totals.

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Table 1-2: Exclusive Mineral Resource Statement

for Granite Creek Underground - Effective Date March 31, 2026

Classification

Domain

Mass

(000’s

tonnes)

Mass

Attributable

(000’s

tonnes)

Au

Grade

(g/t)

Material

Content Au

(000’s t. oz)

Attributable

Material

Content Au

(000’s t. oz)

Measured

Ogee

175

175

6.82

38.4

38.3

Otto

166

166

5.58

29.7

29.7

SPZ

31

31

6.44

6.5

6.5

Total

372

372

6.23

74.6

74.5

Indicated

Ogee

241

241

6.52

50.6

50.6

Otto

425

425

5.23

71.5

71.5

SPZ

771

771

4.70

116.6

116.6

Total

1,438

1,438

5.16

238.6

238.6

M&I

Ogee

417

416

6.64

89.0

88.9

Otto

591

591

5.32

101.2

101.2

SPZ

803

803

4.77

123.1

123.1

Total

1,810

1,810

5.38

313.3

313.1

Inferred

Ogee

38

38

6.71

8.3

8.3

Otto

417

417

7.62

102.0

102.0

SPZ

415

405

6.49

86.6

83.9

Total

871

860

7.04

197.0

194.2

Source: SRK, 2026

Notes:

1.

Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of

Mineral Resources may be materially affected by environmental, permitting, legal, title, market or other relevant issues. The deposit has been classified as Measured, Indicated and Inferred based on confidence in the geological model and drill

spacing. The quantity and grade of reported Inferred resources are uncertain in nature, and there has not been sufficient work to define these Inferred Mineral Resources as Indicated or Measured resources. There is no certainty that any part of a

mineral resource will ever be converted into reserves.

2.

Mineral Resources are reported on an Exclusive basis with units reported in in Metric Tonnes, grade is reported

in grams/tonnes, and metal in 000’s troy ounces.

3.

Mineral Resources are reported fully diluted within Mineable Stopes which have been defined based on metal

price assumptions,* variable metallurgical recovery assumptions, mining costs, processing costs, shipping, G&A costs. Metal Pricing is based on Gold (US$3,000 /oz). Other key assumptions include:

*

Mining Costs (US$/t 190.18);

*

Autoclave Processing Cost (US$/t 131.13), Low-grade Oxide (US$/t 59.94)

*

Autoclave Recoveries ranging 77.6% - 93.4% (average 87.6%) and Oxide recoveries ranging 60.0% - 87.4% (average

68.9%);

*

Shipping Costs (US$/t 16.04) and G&A Costs (US$/t 22.29);

*

Totaling US$/t 281.45 – 359.64 for underground mining and Processing;

*

Based on the variable recoveries an average cut-off grade of 3.75 -

9.81 g/t (0.286 oz/st) has been used reporting.

4.

Mineral Resources have been depleted for previous underground mining activity, and reported on an attributable

basis

5.

Mineral Resources are based on validated data, which have been subjected to QA/QC analysis, using capped,

composited samples at 2 m (5ft) intervals. Estimation has been completed using a combination of OK and IDW estimation methodologies and classified based on confidence in the underlying data and drill spacing. Mineral resource tonnages have been

rounded to reflect the precision of the estimate.

6.

The mineral resources were estimated by SRK Consulting (U.S.), Inc.

7.

Rounding of some figures may lead to minor discrepancies in totals.

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1.4.2

Open Pit

Table 1-3 shows the pit-constrained open pit Mineral Resource

at a gold grade cut-off of 0.20 g/t. (0.005 oz/st)

Table

1-3: Granite Creek Open Pit Mineral Resource Statement

Deposit

Cutoff

Grade

(ppm)

Mass

(000’s

tonnes)

Mass

(000’s

short

tons)

Mass

Attributable

(000’s tonnes)

Au

Grade

(g/t)

Au

Grade

(opt)

Au

Contained

(‘000s tr oz)

Attributable

Material

Content Au

(000’s tr oz)

Measured

Pit B

0.2

3,900

4,299

3,575

1.06

0.031

133.49

122.37

Pit A

0.2

1,098

1,211

1,007

0.71

0.021

25.08

22.99

CX

0.2

13,493

14,873

12,835

1.10

0.032

478.41

454.02

Mag

0.2

15,055

16,596

14,703

1.03

0.030

499.50

487.90

Total

33,546

36,978

32,121

1.05

0.031

1,136.47

1,087.28

Indicated

Pit B

0.2

823

907

754

0.64

0.019

16.86

15.45

Pit A

0.2

969

1,069

889

0.68

0.020

21.06

19.30

CX

0.2

4,423

4,876

4,189

1.00

0.029

141.96

133.03

Mag

0.2

11,509

12,687

11,415

0.74

0.022

275.03

272.53

Total

17,725

19,538

17,247

0.80

0.023

454.91

440.32

Measured + Indicated

Pit B

0.2

4,723

5,206

4,329

0.99

0.029

150.34

137.82

Pit A

0.2

2,068

2,280

1,896

0.69

0.020

46.13

42.29

CX

0.2

17,916

19,749

17,024

1.08

0.031

620.37

587.05

Mag

0.2

26,565

29,283

26,119

0.91

0.026

774.53

760.43

Total

51,271

56,517

49,368

0.97

0.028

1,591.38

1,527.60

Inferred

Pit B

0.2

50

55

46

0.52

0.015

0.84

0.77

Pit A

0.2

440

485

403

0.44

0.013

6.22

5.70

CX

0.2

1,997

2,201

1,921

0.94

0.028

60.66

58.24

Mag

0.2

1,339

1,476

1,320

0.75

0.022

32.12

31.69

Total

3,826

4,217

3,689

0.81

0.024

99.84

96.40

1)

The effective date of the Mineral Resources Estimate is March 31, 2026

2)

The Qualified Persons for the estimate is GRE.

3)

Mineral resources are not ore reserves and are not demonstrably economically recoverable.

4)

Mineral resources are reported at a 0.20 g/t cutoff, an assumed gold price of 3,000 $/tr. oz, using variable

recovery, a slope angle of 41 degrees, 6% royalty, heap leach processing cost $9.04 per tonne (includes admin), CIL processing cost of $17.22 per tonne (includes admin).

5)

Attributable resources are those resources excluding certain third party claims in the Pinson #1A-18A unpatented mining claims and in the Section 28 Fee lands.

Please note

that Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

1.5

Metallurgical Testing and Recovery Methods

Metallurgical test work to support the Granite Creek Project has been undertaken at several commercial laboratories including:

•

Dawson Metallurgical Laboratories

•

FLSmidth

•

McClelland Laboratories

•

SGS

Scope of testing has included bench top autoclave and CIL, cyanide destruction, and solid-liquid separation, along with continuous pressure

oxidation testing. The results of these tests were used to inform the projected gold recoveries for the Project.

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Granite Creek production will be hauled to the refurbished Lone Tree pressure oxidation

facility for processing. Lone Tree unit operations include:

•

Ore Reclaim, Grinding, and Thickening and Acidulation

•

Pressure Oxidation

•

POX Off-gas Treatment and Quench Water Loop

•

Neutralization, Carbon-in-Leach,

and Cyanide Destruction

•

Tailings Thickening and Filtration

•

Acid Wash, Carbon Stripping, and Carbon Regeneration

•

Electrowinning and Refinery

•

Plant and Instrument Air

•

Oxygen Plant

•

Reagent Preparation and Storage

•

Process and Plant Services Cooling Towers

•

Water Distribution

•

Steam Generating Plant and Propane Storage

Tailings thickening and filtration are new operations at Lone Tree required for filtered tailings disposal. The Lone Tree autoclave will

operate under acid conditions as this provides superior gold recoveries compared to alkaline pressure oxidation.

Granite Creek production

can also be shipped for third party processing.

1.6

Mineral Reserve

Underground Mineral Reserves were estimated by SRK. Estimates were prepared for the Ogee, Otto and SP zones located beneath and adjacent to the

historically mined open pit.

Potential mining blocks are generated from the resource block model based on the deposit geometry,

metallurgical process, and block values. Measured Mineral Resources were converted to Proven Mineral Reserves and Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the appropriate modifying factors, as described

herein. Inferred Mineral Resources were not converted to Mineral Reserves. However, where Inferred Mineral Resources are contained within the Mineral Reserves designs, they were assigned zero grade and are considered internal dilution.

Current estimated project costs and the calculated economic cut-off grade (CoG) are shown inTable 1-4. For reporting reserves within the design, a stope cut-off range of 4.42 g/t (0.13 oz/st) Au and 6.41 g/t (0.19 oz/st) Au was used for oxide material and range of 5.21 g/t

(0.15 oz/st) Au and 6.17 g/t (0.18 oz/st) Au was used for the sulfide material. Development headings generate additional material grading between 1.51 g/t (0.04 oz/st) Au and 2.19 g/t (0.06 oz/st) Au oxide material and2.47 g/t (0.07 oz/st) Au and

02.91 g/t (0.08 oz/st) Au for sulfide material. This material is economic to process on an incremental basis and is therefore included in the mineral reserve estimate.

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Table 1-4: Underground Cut-off Grade Parameters

Process

Method1

Parameter

Imperial Value

Imperial Unit

Metric Value

Metric Unit

Gold Price

2,500

US$/oz

80.3769

US$/g

TC/RC

1.85

US$/oz

0.0595

US$/g

NSR Royalty

6%

Nevada Excise Tax

0.75%

Oxides

Metallurgical Recovery

60.0% - 87.4%

Mining Cost

172.53

US$/st

190.18

US$/t

Process Cost

54.38

US$/st

59.94

US$/t

Shipping Cost

14.55

US$/st

16.04

US$/t

G&A Cost

20.22

US$/st

22.29

US$/t

Total Cost

261.68

US$/st

288.45

US$/t

Stope Cut-off Grade

0.129 – 0.187

oz/st Au

4.423–6.411

g/t Au

Processing Cut-off Grade2

0.044 – 0.064

oz/st Au

1.509–2.194

g/t Au

Sulfide

Metallurgical Recovery

77.6% - 92.2%

Mining Cost

172.53

US$/st

190.18

US$/t

Process Cost

118.96

US$/st

131.13

US$/t

Shipping Cost

14.55

US$/st

16.04

US$/t

G&A Cost

20.22

US$/st

22.29

US$/t

Total Cost

326.26

US$/st

359.64

US$/t

Cut-off Grade

0.152–0.180

oz/st Au

5.211–6.171

g/t Au

Processing Cut-off Grade2

0.072–0.085

oz/st Au

2.469–2.914

g/t Au

Source: SRK, 2026

1

Dynamic cut-off grade was used for the reserves design based on whether

the material is oxide or sulfide.

2

Processing Cut-off Grade considers the processing cost and G&A

cost.

Mineral Reserves were classified using the 2014 Canadian Institute of Mining Metallurgy and Petroleum Definition

Standards on Mineral Resources and Mineral Reserves (CIM Definition Standards). Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the appropriate modifying factors, as described herein, to potential mining shapes

created during the mine design process. In the same manner, Measured Mineral Resources were converted to Proven Mineral Reserves.

A 3D

design has been created representing the planned reserve mining areas. The underground mine design process resulted in 2.20 Mt (2.42 Mst) of Mineral Reserves at an average grade of 7.87 g/t (0.23 oz/st) Au. Table

1-5 presents the mineral reserve statement as of March 31, 2026.

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Table 1-5: Mineral Reserve Estimate for Granite

Creek as of March 31, 2026—SRK Consulting (U.S.), Inc.

Classification

Domain

Metric Units

Imperial Units

Tonnage

(000’ tonnes)

Au Grade

(g/t)

Tonnage

(000’s s. tons)

Au Grade

(oz/st)

Au Metal

Content

(000’s t. oz)

Proven

Ogee

71

9.34

78

0.27

21.2

Otto

171

7.04

189

0.21

38.7

SP

40

11.21

44

0.33

14.5

Stockpile

26

7.33

29

0.21

6.2

Total

308

8.13

340

0.24

80.6

Probable

Ogee

88

7.95

97

0.23

22.4

Otto

366

6.70

403

0.20

78.8

SP

1,437

8.11

1,584

0.24

374.7

Total

1,890

7.83

2,084

0.23

475.9

Proven + Probable

Ogee

158

8.57

174

0.25

43.6

Otto

537

6.81

592

0.20

117.6

SP

1,477

8.19

1,628

0.24

389.2

Stockpile

26

7.33

29

0.21

6.2

Total

2,198

7.87

2,423

0.23

556.5

Source: SRK, 2026

Notes:

•

All figures are rounded to reflect the relative accuracy of the estimates. Totals may not sum due to rounding.

Mineral Reserves have been stated on the basis of a mine design, mine plan, and economic model.

•

Reserves are reported using a stope CoG range of 4.42 g/t (0.13 oz/st) Au – 6.41 g/t (0.19 oz/st) Au and an

incremental processing CoG range of 1.51 g/t (0.04 oz/st) Au – 2.91 g/t (0.08 oz/st) Au depending on the processing method. The CoG calculations assume a US$2,500/oz Au price, and 60% to 92% metallurgical recovery depending on the processing

method applied.

•

Operating costs include mining (US$190.18/tonne), processing (US$59.94/tonne to US$131.13/tonne), G&A

(US$22.29/tonne), Shipping Costs (US$/t 16.04).. Additionally, a 6% royalty and 0.75% Excise Tax are applied, along with TC/RC charges of US$1.85/oz Au.

•

The reserves are mined using the underhand drift and fill method with cemented waste rock backfill. Mining

dilution of 10% is included in the reserve at zero grade.

•

Stockpile reserves reflect ore mined and placed on surface stockpiles prior to the effective date and not yet

processed.

•

The Mineral Reserves were estimated by SRK Consulting, U.S. (Inc.).

1.7

Summary Capital and Operating Cost Estimates

Costs for processing vary depending on ore type and processing method.

For processing refractory ore at Lone Tree, costs are as follows:

•

Sustaining OPEX of US$25.99/metric ton (processing fee). The processing fee has been presented by i-80 as an

Autoclave CAPEX Allocation as noted in Table 1-6

•

Processing cost of US$105.16/metric ton

For processing oxide ore at Lone Tree, costs are US$33.97/metric ton.

For third-party toll processing, costs are in accordance with the terms of the toll-milling contract and consist of costs incurred by the third

party for processing, a treatment charge that varies based on the gold price and any penalties for material that does not meet the specifications of the contract.

LoM capital costs are summarized in Table 1-6.

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Table 1-6: Life of Mine Capital Costs

Capital Costs

Unit

Value

Mine Capex

US$

M

60.74

Autoclave Capex Allocation

US$

M

49.33

Surface Capitalized Costs

US$

M

22.19

Closure

US$

M

12.66

Total

US$

M

144.92

Source: SRK, 2026

US$M: US$ million

LoM Operating

Costs are summarized in Table 1-7.

Table 1-7: Life of

Mine Operating Costs

Operating Costs

Unit

Value

Mining Cost

US$M

395.98

Oxide Cost

US$M

0.65

Autoclave Cost

US$M

267.30

Ore Transport Cost

US$M

36.27

Surface Ops Cost

US$M

117.85

Power Cost

US$M

32.77

G&A Cost

US$M

49.87

Total

US$M

900.68

Mining Cost

USD/ton ore

165.38

Oxide Cost

USD/ton ore processed

17.95

Autoclave Cost

USD/ton ore processed

111.97

Ore Transport Cost

USD/ton ore

15.15

Total Processing Cost

USD/ton ore processed

125.53

Surface Ops Cost

USD/ton ore

49.22

Power Cost

USD/ton ore

13.69

G&A Cost

USD/ton ore

20.58

Total

USD/ton ore processed

371.67

Mining Cost

USD/tonne ore

182.30

Oxide Cost

USD/tonne ore processed

19.78

Autoclave Cost

USD/tonne ore processed

123.43

Ore Transport Cost

USD/tonne ore

16.70

Total Processing Cost

USD/tonne ore processed

138.37

Surface Ops Cost

USD/tonne ore

54.25

Power Cost

USD/tonne ore

15.09

G&A Cost

USD/tonne ore

22.68

Total

USD/tonne ore processed

409.69

Source: SRK, 2026

US$M: US$ million

1.8

Economic Analysis

Economic analysis, including estimation of capital and operating costs, is inherently a forward-looking exercise. These estimates rely upon a

range of assumptions and forecasts that are subject to change depending upon macroeconomic conditions, operating strategy and new data collected through future study or operations and therefore actual economic outcomes often deviate significantly

from forecasts.

The Granite Creek operation consists of an underground mine with no onsite processing facility that mines and transports

gold ore to offsite processing facilities.

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The economic analysis metrics are prepared on annual

after-tax basis in US$. The results of the analysis are presented in Table 1-8. The results indicate that at a gold price of US$2,750/oz, the after tax NPV @ 5% is

US$118 million.

Table 1-8: Indicative Economic Result

LOM Cash Flow

Unit

Value

Total Revenue

US$M

1,328.81

Royalty

US$M

(104.47

)

Total Opex

US$M

(900.68

)

Operating Margin

US$M

323.65

Operating Margin Ratio

%

24

%

Taxes Paid

US$M

(25.35

)

Free Cashflow

US$M

153.38

Before-Tax

Free Cash Flow

US$M

178.73

NPV @ 5%

US$M

137.66

NPV @ 7.5%

US$M

120.98

NPV @ 10%

US$M

106.39

After-Tax

Free Cash Flow

US$M

153.38

NPV @ 5%

US$M

117.55

NPV @ 7.5%

US$M

102.97

NPV @ 10%

US$M

90.19

Source: SRK, 2026

US$M: US$ million

Within the

constraints of this analysis, the operation’s NPV appears to be most sensitive to metal prices, recovery assumptions and mined grades.

A summary of the cash flow on an annual basis if presented in Figure 1-1.

Source: SRK, 2026

Figure 1-1: Annual Cash Flow Summary

1.9

Permitting Requirements

The Granite Creek Underground Project occupies a site with mining history dating to the 1940’s. The Project supports a historic surface

mining operation and the current underground operation. The surface operation is permitted through Bureau of Land Management (BLM) and NDEP, while the underground operation is located on private property and is permitted through the NDEP. Limited

underground support facilities are permitted through the surface mine permits, under the jurisdiction of the BLM and NDEP.

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The surface operations have been the subject of environmental review since 1975, when the

first Environmental Assessment (EA) was prepared for the U.S. Department of the Interior – Bureau of Land Management (BLM) under the National Environmental Policy Act (NEPA). Two further assessments followed in 1992 and 2003. The surface

permits have been updated from time to time to accommodate closure and underground support facilities.

The majority of the current

underground project sits on private land and falls under the jurisdiction of the Nevada Division of Environmental Protection – Bureau of Mining Regulation and Reclamation (NDEP-BMRR) the underground operation permit has been updated from time-to-time to accommodate ongoing operations.

Geochemical

testing, beginning with a comprehensive 1998 study and continued through quarterly sampling from 2005 to 2025, consistently shows that waste rock placed in the CX Pit has strong acid-neutralizing capacity and low potential for acid rock drainage

(ARD). The main water-quality concern is metal leaching, particularly arsenic and antimony, which periodically exceed NDEP reference values in bedrock groundwater and underground mine water. These constituents are actively managed through the

site’s water treatment plant (WTP), which reliably brings effluent into compliance. A metals attenuation study is under development to support site specific discharge requirements that could reduce water treatment requirements for select

constituents and lower WTP operational costs. Surface water in Granite Creek and alluvial groundwater both meet applicable standards. The Mag Pit lake, monitored since 2015, shows elevated arsenic and manganese, but a 2025 SRK ecological risk

assessment (ERA) concluded these levels are not likely to cause harm to receptors. Modeling of the pit lake’s long-term, post-closure water quality indicates compliance with state standards over a

100-year horizon, with a proven contingency treatment (ferric sulfate dosing) available if arsenic levels were to rise.

SRK is not aware of any known environmental issue that would materially impede extraction of the project’s Mineral Resources or reserves.

One issue with potential cost implications was identified in 2025: petroleum-contaminated soil was found within CX Pit waste rock backfill. Because the exact location of the contaminated material is not known, as much as 1.17 million tons of

backfill may need to be re-handled and sampled. Osgood is working with NDEP-BMRR to develop a management plan and remediate the issue; this represents the project’s principal environmental cost

uncertainty rather than a permitting or resource-access risk.

The project operates under a full suite of state and federal environmental

management plans covering water management, waste rock handling, spill response, monitoring, and closure, most recently updated as part of permit amendments. No ore is processed on site; mined material is trucked to a permitted third-party facility.

Waste rock is placed in the CX Pit and C Pit under the terms of the site’s water pollution control permit and is sampled quarterly for acid-generation and metal-leaching potential. No tailings are generated by current operations, and the two

legacy tailings facilities on site have been reclaimed and released from financial assurance obligations. A second water treatment plant is planned to handle the project’s growing water management needs, and stormwater is actively diverted

around the site through an engineered pipe system under an active general permit. Monitoring spans water quality, waste rock chemistry, air quality, surface disturbance and reclamation, and stormwater controls, and the site’s permit-monitoring

obligations are all current.

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The project holds the permits needed for its current operations from Humboldt County, the

State of Nevada, and the BLM, with no permitting actions currently pending before any agency. Water rights total 3,746 ha (9,256 acre)-feet annually, of which 465 ha (1,149 acre)-feet is consumptive use, drawn primarily from certificated rights and

permits within a groundwater basin designated by the State Engineer for closer administration; this designation reflects committed water rights relative to the basin’s yield rather than evidence of depletion. Financial assurance bonding for

reclamation is in place and current, with the next required three-year bond review anticipated in 2028 to 2029.

Osgood Mining has no

formal agreements with local communities currently but maintains ongoing stakeholder engagement through periodic town hall meetings in Winnemucca, participation in regional economic development efforts, and support for local education initiatives.

The company follows i-80 Gold Corp’s broader corporate governance policies on local procurement and hiring.

Closure planning follows established state and federal reclamation requirements, with the current mine plan assuming the CX Pit is backfilled

(and the Mag Pit lake becomes a long-term groundwater sink posing no identified risk to regional groundwater quality. The current reclamation cost estimate, prepared using Nevada’s standardized cost model, is US$8.64 million; this figure

reflects a government-administered closure scenario and is expected to differ from the cost the operator would actually incur, and from the eventual cost at the 2032 anticipated closure date, as site conditions evolve. Overall, given the depth of

the existing regulatory framework and the documentation reviewed, SRK’s opinion is that the project’s current environmental, permitting, and community plans are adequate.

1.10

Conclusions and Recommendations

1.10.1

Drilling

The QP considers that the drilling and sampling information is sufficiently reliable to interpret the boundaries of the mineralized structures

and domain interpretations, and that the sample grade data are sufficiently reliable to support the MRE. The QP also notes that mineralization remains open at the deposit.

The 2027 to 2028 exploration program targets five zones at Granite Creek: the CX Fault depth / north extension, Rangefront and Adam Peak, the

Ogee Zone, SPZ infill and depth extension, and the SPZ / Rangefront / Mag structural intersection.

The combined drill program comprises

128 planned drillholes totaling 32,614 m (107,000 ft) plus associated CX Fault underground development and a gravity survey at the SPZ / Rangefront / Mag intersection. The proposed work program has an estimated cost of US$16 million based on

current conditions.

In the opinion of the QP, the proposed program is technically reasonable and appropriately sequenced, with Rangefront

and Ogee prioritized ahead of active and planned underground development to reduce the risk of mining ahead of drill data, while the CX Fault program advances the known

oxide-to-sulfide transition at depth.

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Upon completion of the above items, an update to the geologic model and MRE should be

conducted, along with updated metallurgical updated recovery assumptions.

1.10.2

Mining

•

Mine infrastructure has been completed. Production ramp up has reached approximately 400 tons per day. The mining

contractor is in place with the full complement of equipment and personnel.

•

Decline development has accessed 700 vertical ft of mineralization of the Otto and Ogee zones. Development has

reached the top of the SPZ allowing additional active production stopes.

•

The drill lateral drift over the SPZ has been completed.

•

Reconciliation of the model to mill indicates process head ounces exceed model by 19%. This appears to be from

mining in a larger low-grade halo around high grade core.

•

Processed grade is lower than the

life-of-mine planned grade due to extensive mining of marginal mineralization below the economic cut-off grade.

1.10.3

Metallurgical Testing

•

Conduct additional variability sampling using the most recent mine plan to select samples to provide additional

recovery information to further populate recovery models.

•

Testing should attempt to establish head grade and extraction relationships for use in more detailed resource

modeling.

•

Mineralogy impacts need to be established and geologic domains within each resource need to be determined.

•

Testing of additional samples to provide comminution data to assess hardness variability within the zones and any

potential impacts on throughput in the Lone Tree process plant.

•

The estimated cost for the suggested next phase metallurgical program is US$150,000 based on current market

pricing.

1.10.4

Permitting / Environmental

Dewatering

Complete the planned dewatering well and re-evaluate the ground water model and inflow into the

underground workings.

1.10.5

Economics

The Granite Creek operation consists of an underground mine with no onsite processing facility that mines and transports gold ore to offsite

processing facilities. The operational life as modeled for this analysis is approximately 9 years. Under the forward-looking assumptions modeled and documented in this report, the operation is forecast to generate positive cumulative cash flow.

However, the Project does not generate positive free cashflow in 2026 and 2027 largely due to the higher processing costs incurred because of third-party processing of ore and a six-month stockpiling period in

2027. Cashflow is forecast to be negative at the end of the operational life as the operation winds down, and closure costs are incurred. This estimated cash flow is inherently forward-looking and dependent upon numerous assumptions and forecasts,

such as macroeconomic conditions, mine plans and operating strategy, that are subject to change.

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The results indicate that at a gold price of US$2,750/oz the

after-tax NPV @ 5% is US$118 million.

The sensitivity analysis performed indicates that the

operation’s NPV is most sensitive to metal prices, recovery assumptions and mined grades.

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2

Introduction

2.1

Registrant for Whom the Technical Report Summary was Prepared

This Technical Report Summary was prepared in accordance with the Securities and Exchange Commission (SEC)

S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for i-80 Gold Corp. (“i-80” or the

“Registrant”) on the Granite Creek Underground Project.

This Technical Report Summary is filed as an exhibit to i-80 Gold Corp.’s Current Report onForm 8-K.

2.2

Terms of Reference and Purpose of the Report

The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in SRK’s

services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by i-80 subject to the terms and conditions of its contract with SRK and relevant securities legislation.

The purpose of this Technical Report Summary is to report mineral resources, mineral reserves and exploration results.

The effective date of this report is March 31, 2026.

2.3

Sources of Information

The sources of information include data and reports supplied by i-80 personnel as well as documents

cited throughout this report and listed in the References section. It is the QP’s opinion that the information and test work provided is fit for purpose for the current level of study.

2.4

Details of Inspection

Site visits conducted by the qualified persons are summarized in Table 2-1.

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Table 2-1: Site Visit Participants

Company

Expertise

Date(s) of Visit

Details of Inspection

SRK Consulting

Geology

February 4, 2026

UG mine tour and GC, drill shed and core review

WSP

Geology, Infrastructure

November 5-6, 2025

The purpose of the visit was to review geological, geotechnical, mining, and infrastructure conditions relevant to the SPZ and planned mine expansion. The site visit included inspection of drill core, underground workings, ground

support systems, cemented rockfill production facilities, and the proposed ventilation raise location to surface.

WSP

Ventilation

November 5, 2025

UG mine tour and ventilation assessment including inspection of primary, booster, and auxiliary ventilation systems; review of active and planned mine workings, haulage routes, escapeways, and proposed ventilation raise locations;

collection of temperature and humidity data for Ventsim™ model calibration.

Observations made during the site visits were generally consistent with the geological model,

geotechnical logging database, underground mapping, laboratory testing results, and operational information reviewed for this study.

2.5

Report Version Update

The user of this document should ensure that this is the most recent Technical Report Summary for the property.

This Technical Report Summary updates a previously disclosed Initial Assessment (IA) on the Granite Creek Open Pit Project (reported

March 26, 2025). Based on additional drilling and a shift in focus to the underground operation, the qualified person has determined that the IA is no longer current and has been superseded by this Technical Report Summary. This Technical

Report Summary is not valid if a further Technical Report Summary has since been issued.

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3

Property Description

3.1

Property Description

The Granite Creek Project is located in Humboldt County, Nevada, 45 km (28 m) northeast of the town of Winnemucca, and it is part of the

historic Potosi mining district. It is centered at roughly 41° 8’ N latitude and 117° 15.5’ W longitude. It encompasses about 1,823.5 ha (4,506 acres) including owned unpatented claims, leased unpatented claims and owned surface

fee land. The federal land is administered by the BLM. Figure 3-1 shows the location of the Granite Creek Project

Source: SRK, 2026

Figure 3-1: Granite Creek Project Location

3.2

Mineral Tenure and Surface Rights

Mineral title to the Granite Creek Project is held by Osgood Mining Company, LLC (Osgood) and Premier Gold Mines USA, Inc., with additional

unpatented claims leased from third parties. Surface and mineral rights on fee lands are owned by Osgood. Osgood holds the operating permits and reclamation bonds for the Project and is responsible for all current mining and exploration activities.

The Land Position Map along with the summary of the mineral and surface titleholders for the Project can be found below in Figure 3-2 and Table 3-1.

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Source: SRK, 2026

Figure 3-2: Land Position Map, Granite Creek Project

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Table 3-1: Summary of Mineral and Surface Title

Holders, Granite Creek Project

Legal Description

Identifying Numbers

Owner(s)

T38N R42E - Section 28 CX 1A-23A

(Unpatented Lode Mining Claims

BLM Serial No.: NMC319833 - NMC319855

Osgood Mining Company, LLC

T38N R42E - Section 28 PACIFIC 1A-7A

(Unpatented Lode Mining Claims)

BLM Serial No.: NMC319814 - NMC 319820

Osgood Mining Company, LLC

T38N R42E - Section 32 PINSON 1A-18A

(Unpatented Lode Mining Claims)

BLM Serial No.: NMC319856 - NMC319873

Osgood Mining Company, LLC (41.67%),

Premier

Gold Mines USA Inc. (50%), Michael C. Murphy (8.33%)

T38N R42E - Section 32

LEASED: BEE DEE 1A-18A, FRACTION 1 & 2

(Unpatented Lode Mining Claims)

BLM Serial No.: NMC282121 - NMC282122

Lessors: Franco-Nevada & S&G Pinson Lessee: Osgood Mining Company, LLC

T37N R42E - Section 6 LEASED: BEE DEE 21-56

(Unpatented Lode Mining Claims)

BLM Serial No.: NMC282123 - NMC282158

Lessors: Franco-Nevada & S&G Pinson Lessee: Osgood Mining Company, LLC

T38N R42E - Section 21

640 Acres

(FEE)

APN 07-0121-05

Osgood Mining Company, LLC

T38N R42E - Section 28

120 Acres

(FEE)

APN 07-0121-07

Osgood Mining Company, LLC (41.67%), Premier Gold Mines USA Inc. (50%), Michael C. Murphy (8.33%)

T38N R42E - Section 29

640 Acres

(FEE)

APN 07-0121-06

Osgood Mining Company, LLC

T38N R42E - Section 31

640 Acres

(FEE)

APN 07-0121-08

Premier Gold Mines USA Inc.

T38N R42E - Section 33

640 Acres

(FEE)

APN 07-0121-33

Osgood Mining Company, LLC

Source: i-80 Gold, 2026

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Unpatented claims have annual maintenance fees of US$200 per claim payable to the BLM and a

notice of intent to hold (NIH) in the amount of US$12 per claim plus US$12 document filing fee payable to Humboldt County. Claim maintenance fees are paid through September 1, 2026, with the BLM. The NIH fees to Humboldt County are paid through

November 1, 2026; payments are current at the time of this report. All mineral claims and fee lands comprising the Granite Creek Project are valid, in good standing, and free of any known disputes at the effective date of this report. Fee land

is subject to Nevada State real property tax, and certain mine infrastructure is subject to Nevada State personal property tax. Leased unpatented claims are subject to yearly lease fees. Holding costs for 2026 are listed in Table 3-2.

Table 3-2: Holding Costs, Granite Creek Project

Description

Payee

Quantity

Amount (US$)

Unpatented Claim Maintenance Fee

BLM

104

20,800.00

Notice of Intent to Hold Unpatented Claims

Humboldt County

104

1,260.00

Real Property Taxes

Humboldt County

5 parcels

7,234.63

Personal Property Taxes

Humboldt County

various infrastructure

94,594.15

Lease fees, annually adjusted by CPI

Lease Holders

56 unpatented claims

122,495.24

Total

$

246,384.02

Source: SRK, 2026

Ownership of the Granite Creek Project land position comprises various forms of title. i-80 Gold Corp

owns 48 unpatented lode claims covering about 363 ha (897 acres) (Table 3-3), and leases 56 unpatented lode claims covering about 468 ha (1,156 acres) (Table 3-4). i-80 Gold Corp also owns, through its subsidiaries, fee surface land parcels covering about 777 ha (1,920 acres).

Table 3-3: Owned Unpatented Claims, Granite Creek Project

Claim Name

Location

BLM Legacy Number

Claim Type

Number

of Claims

PACIFIC #1A - #7A

Section 28 of Township 38 North, Range 42 East

NMC319814 -NMC 319820

Lode

7

CX #1A - CX #23A

Section 28 of Township 38 North, Range 42 East

NMC319833 -NMC319855

Lode

23

PINSON #1A - #18A

Section 32 of Township 38 North, Range 42 East

NMC319856 -NMC319873

Lode

18

Total Owned Unpatented Claims

48

Source: SRK, 2026

Table 3-4: Leased Unpatented Claims, Granite Creek Project

Claim Name

Location

BLM Legacy Number

Claim Type

Number of Claims

NEW BEE DEE FRAC. #1, #2

Section 32 of Township 38 North, Range 42 East

NMC282121, NMC282122

Lode

2

BEE DEE #21 -BEE DEE #56

NMC282123 -NMC282158

Lode

36

BEE DEE #1A

NMC319892 -NMC319909

Lode

18

Total Leased Unpatented Claims

56

Source: SRK, 2026

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3.2.1

Owned Unpatented Mining Claims

i-80, through OMC, owns a 100% interest in the Pacific #1A-7A

as well as the CX1A-23A unpatented mining claims.

i-80,

through OMC and PGU, owns a 91.67% interest in the Pinson #1A-18A mining claims. The remaining 8.33% interest in these claims is owned by Michael Murphy and is not leased by OMC. The fact that OMC has not

leased the unowned 8.33% interest in these claims does not preclude OMC from mining the claims. By law, OMC, as the co-owner of an undivided interest in these claims, has the right to mine the claims without

permission or approval from (and even over any objections by) the other co-owner, subject, however, to an obligation on the part of OMC to account to the other co-owner

for their proportionate shares of mining revenues less their proportionate shares of mining expenses.

3.2.2

Leased Unpatented Mining Claims

Pursuant to a mining lease agreement from May 9, 1980, which has a term that extends until May 9, 2040, i-80, through OMC as current lessee thereunder, controls a 100% interest in the Bee Dee group of unpatented mining claims listed in Table 3-4 above. Following subsequent

assignments and corporate transactions, Franco-Nevada U.S. Corporation (50%) and S&G Pinson, LLC (50%) (“S&G Pinson”) are the current lessors under this lease (the “Bee Dee Lease Agreement”).

3.2.3

Fee Lands

“Fee land” refers to privately owned real property in which title has passed out of United States government ownership. Fee land at

the Granite Creek Project is identified by Humboldt County Assessor’s Parcel Number (APN) and, unless otherwise stated, comprises both the surface estate and the mineral estate. Fee land is distinguished from unpatented federal lode mining

claims, which are possessory interests in federal land administered by the BLM and remain subject to annual maintenance obligations.

i-80, through OMC, owns a 100% interest in Sections 29 and 33, Township 38 North, Range 42 East.

i-80, through OMC and PGU, also owns a 91.67% interest in the 120-acre (48.5-ha) parcel comprising the east 1⁄2 of the southwest 1⁄4 and southeast

1⁄4 of the southwest 1⁄4 of Section 28, Township 38 North, Range 42 East.

The remaining interest in this parcel is owned by Michael Murphy (8.33% undivided interest).

3.3

Royalties and Net Profit Interests

The Granite Creek Mine Project is subject to several net smelter royalty obligations, each with specific areas of applicability within the

Property and with rates that vary based on location or any pre-existing royalties on the land in question. In addition, there is one mineral production royalty agreement (the “MRPA”, defined

further below) which provides the counterparty with a 10% “net profits” interest (also as described further below) and is not contingent merely on production like a typical net smelter royalty agreement.

For the purposes of the economic analysis contained in this Report—which is restricted to the estimated mineral reserves for the

underground project at Granite Creek—the total combined applicable royalty is expected to be 7.5% until January 2031 and 9% thereafter. This total payable NSR and 10% “net profit” interest is based on the following facts and

qualifications:

•

The currently anticipated underground operations will impact areas under the Pacific Claims and portions of the

fee lands included in Section 29.

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•

The current 7.5% combined royalty applicable to the Pacific Claims and the Section 29 fee lands includes 1%

NSR and 3% NSR, respectively, owing under the 1996 Royal Gold Royalty:

•

0.17% of the royalty owing under the Section 29 fee lands is payable by the minority owner of this privately

held land and not OMC.

•

A Franco Nevada royalty (described further below) of 1.5% included within this 7.5% NSR estimate increases to a

3% NSR effective January 2031.

•

Applying the gold price assumptions used in this Report for the anticipated life of mine economic analysis

included in Chapter 22 (US$2,750 per gold oz), it is not anticipated that any “net profits” will be payable under the MRPA. However, if future gold prices are materially higher than as projected within this Report, then payments under

this MRPA may be triggered.

The following section describes the royalties present on the various land rights within the

Granite Creek Property.

3.3.1

1996 Royal Gold Royalty – Granite Creek Property (Royal Gold, Inc. – Current Holder)

In a NSR Royalty Agreement dated November 30, 1996, Pinson Mining Company (PMC) agreed to provide Rayrock

Mines, Inc. et al. an overriding NSR royalty that varied depending on the nature of the particular land holding and any underlying royalties existing on that land at the time of the transaction. Through a series of subsequent deeds, transfers and

corporate transactions this royalty interest is now held by Royal Gold, Inc. (which holds a 97.9% interest) and D.M. Duncan Inc. (which holds the remaining 2.1% interest), hereinafter referred to as the “1996 Royal Gold Royalty”.

The royalty under this agreement applies to the entire Granite Creek Property, but it is not payable until 200,000 troy ounces (oz) of gold

have been produced. Currently, based on historical production at Granite Creek post the execution of this NSR Royalty Agreement, the 200,000 oz threshold has now been reached, and this royalty is currently payable.

As noted, the royalty under this 1996 NSR Royalty Agreement is dependent on the land involved and, as applicable, any pre-existing royalty arrangements affixed to said lands. In practical terms it applies as follows:

•

Fee lands owned by PMC at November 30, 1996—which includes all current fee lands

now owned by OMC—the 1996 Royal Gold Royalty holders receive a 2.5% royalty on parcels not subject to an underlying or pre-existing royalty, and a 0.5% royalty on parcels subject to an underlying or pre-existing royalty, which increases to a 1% NSR royalty if the average gross value per ton of ore produced is greater than US$175/ton (which is currently the case at the time of this Report)

•

Fee lands leased by PMC at November 30, 1996 - the royalty varies from a minimum of

0.5% to a maximum of 5% depending on any underlying royalty. The royalty percentage is determined by the difference between a total royalty load of 6% less any underlying royalty; however, the royalty will never exceed 5% or be reduced to less than

0.5%. For example, if the underlying royalty is 4%, then the Royal Gold Royalty would be 6% less 4%, resulting in a 2% royalty payable to the holders of the Royal Gold Royalty. If the underlying royalty is 0.5%, the Royal Gold Royalty would be 6%

less 0.5% equaling 5.5%, which is greater than 5%, thus reducing the applicable royalty rate to 5%. If the underlying royalty is 6% or greater, the Royal Gold Royalty rate is limited to 0.5%

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•

Unpatented lode mining claims owned by PMC at November 30, 1996 – which includes

all current claims other than the BeeDee claims within the Property—the royalty is 2.5% with a provision to further reduce this royalty if the U.S. Federal Government were ever to impose a “federal royalty payment obligation” on

said claims, subject to a minimum 1.0% NSR in all instances

•

Unpatented lode mining claims leased by PMC at November 30, 1996 – which is

comprised only of the BeeDee claims – the royalty percentage is determined by the difference between a total royalty load of 6% less than the underlying royalty; however, the royalty will never exceed 5% or be reduced to less than 0.5% (as

further articulated under (ii) above)

3.3.2

2026 Franco- Nevada Royalty – Granite Creek Property – 1.5% to 3% NSR

On March 16, 2026, the Company entered into a net smelter return royalty financing arrangement with Franco-Nevada U.S. Corporation

(“FN USA”), granting a perpetual royalty on production from the Company’s Granite Creek project area (the “FN NSR Royalty”). The royalty rate is 1.5% of net smelter returns through December 31, 2030, increasing to

3.0% thereafter. The royalty is payable monthly in cash or in-kind as refined gold or silver at Franco-Nevada’s election.

3.3.3

Pinson Private Royalty – Pinson Claims – 3.125% NSR (Noceto, Phillips, Murphy—Current

Holders) & 0.17% NSR (FN USA & S&G Pinson – Current Holders)

Pursuant to a mining lease

agreement dated May 27, 1980, certain lands were leased to Cordilleran Explorations general partnership which included the Pinson unpatented mining claims described in Table 3-1 and Table 3-3 above (the “Pinson Lease”). Through subsequent transfers and conveyances, the Pinson claims are now 91.67% owned by i-80 (through its wholly owned subsidiaries

OMC and PG USA). As part of the acquisition of the Pinson claims, three individuals, specifically. Mrs. Barbara P. Noceto, Mrs. Patricia Phillips and Mrs. Kate Murphy, conveyed their interest by way of Deed in exchange for a

reserved NSR proportionate to their ownership interest in the Pinson claims at that time. Other conveyances from original counterparties under the Pinson Lease similarly conveyed their interests in exchange for a reserved 2% NSR proportionate to

their ownership interest which currently equates to a 0.17% NSR royalty interest now held by FN USA and S&G Pinson in equal shares.

3.3.4

Pinson Royalty – Section 21 Bee Dee 21-56 Claims –

0.5% NSR (Pinson Mining Company – Current Holder)

By way of a NSR Royalty Deed dated May 6, 2022, Nevada

Gold Mines LLC grants owned and leased lands – including those in Section 21 and the Bee Dee 21-56 unpatented mining claims – to Pinson Mining Company and in exchange reserves a 0.5% NSR

royalty interest in production emanating from these lands.

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3.3.5

2001 Royal Gold Royalty – Section 21, 29 & Pacific Claims – 3-5% NSR (Royal Gold Inc.—Current Holder)

By way of a Deed dated

September 14, 2001, Cordilleran Explorations partnership (“Cordilleran”), the original developer of the Property, agreed to deed (or convey) certain properties – including al the patented lands in Section 21 and 29 of

Township 38 North, Range 42 East, as well as the unpatented Pacific Claims (1A-7A – the “Pacific Claims”) to Pinson Mining Company (PMC), all within the current Property area. In exchange,

Cordilleran retained a 3% NSR on Section 21 and 29 fee lands and a 5% NSR over the Pacific Claims Similar to the 1996 Royal Gold Royalty, following a series of subsequent deeds and transfers, Royal Gold, Inc. is the current owner of the 2001

Cordilleran Royalty – hereinafter the “2001 Royal Gold Royalty”.

3.3.6

Goldfield Royalty – Section 33 – 2% NSR (FN USA—Current Holder)

By way of a Special Warranty Deed dated November 11, 1981, The Goldfield Corporation sold and conveyed certain

properties – including all of the fee lands within Section 33 of Township 38 North, Range 42 East – in exchange for reserving a 2% NSR royalty over any production from said lands (the “Goldfield Royalty”).

Section 33 is now owned by OMC and the Goldfield Royalty is now owed to Franco Nevada U.S. Corporation.

3.3.7

Bee Dee Reserved Royalty—Bee Dee Claims—2% NSR Royalty (FN USA (50%) and S&G Pinson, LLC

(50%) – Current Holders)

As described in 4.2.2 above, OMC controls a 100% interest in the 56 Bee Dee

unpatented federal lode mining claims by way of the Bee Dee Lease Agreement. The Bee Dee Lease Agreement imposes a 2% net smelter returns (NSR) royalty on the Bee Dee claims in favor of the current lessors.

3.3.8

Section 28 Private Royalty – Section 28 Fee lands—2% NSR (Murphy, Noceto, Phillips, D.

and J. Christison – Current Holders)

Pursuant to a mining lease agreement dated May 27, 1980, certain

lands were leased to Cordilleran Explorations general partnership from the original which included the fee lands within Section 28 of Township 38 North, Range 42 East as described in Table 3-1 above.

Through subsequent transfers and conveyances, the fee lands within Section 28 are now 91.67% owned by i-80 (through its wholly owned subsidiaries OMC and PG USA). As part of the acquisition of the

Section 28 fee lands, five individuals noted above, conveyed their interest by way of Deed in exchange for a reserved NSR proportionate to their ownership interest in the Pinson claims at that time. As a result, gold production from the

Section 28 fee lands is subject to a 2% NSR royalty interest to the private parties noted above

3.3.9

Mineral Production Royalty Agreement – Granite Creek Property—10% Net Profit Interest

(NGM—Current Holder)

Pursuant to a mineral production royalty agreement entered into on August 10, 2011,

the entire Property is subject to a 10% “net profits” interest royalty, payable to the current counterparty under this agreement, Gold Royalty U.S. Corp. (the “NPI”).

The NPI is triggered after the first 120,000 ounces of gold (and/or the gold-equivalent of other minerals) are produced from the Property and

OMC has derived “net profits” from the said operation. Under the terms of the NPI, “net profits” are only derived once all “recoverable costs” have been realized by OMC. Recoverable costs are defined broadly to

include all exploration, development and mining related costs – including mine closure and reclamation costs, financing costs and all third-party costs including to government entities and all taxes—incurred by OMC in exploring,

developing and operating the Property. The Property has produced approximately 73,500 ounces as of the date of this Report.

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3.4

History of Ownership and Prior Mining

The Property has been explored by several individuals and mining/exploration companies since the late 1930’s. The original discovery on

the Property was made by Clovis Pinson and Charles Ogee in the mid to late 1930’s, but production did not occur until after World War II, when ore from the original discovery was shipped to and processed at the Getchell mine mill. In 1949 and

1950, total production from the Granite Creek Mine amounted to approximately 9,071 t (10,000 st) grading approximately 0.14 ounces per ton (opt) (4.8 g/t).

3.5

Environmental Liabilities

The Granite Creek Project’s principal identified environmental liability is the 2025 discovery of petroleum-contaminated soil within CX

Pit waste rock backfill; because the affected material’s exact location is unknown, as much as 1.17 million tons of backfill may require re-handling and remediation. Osgood is developing a

management plan with NDEP-BMRR to address it. Beyond this, the project carries a standard reclamation and closure bonding obligation, currently estimated at US$8.64 million under Nevada’s standardized cost model, which will be revised at

the next required three-year bond review (2028 to 2029) and is likely to differ from the operator’s actual future closure cost given expected changes in site conditions before the 2032 anticipated closure date. Ongoing water treatment

represent continuing costs tied to naturally elevated arsenic and antimony in site dewatering water, and a proven contingency treatment (ferric sulfate dosing) is available should arsenic rise in the post-closure Mag pit lake, though water quality

modeling indicates this is unlikely to be needed. A metals attenuation study has been commissioned to support site specific discharge requirements that will be used to optimize operational water treatment costs. Notably, the site’s two legacy

tailings storage facilities have already been reclaimed and released from financial assurance requirements and are explicitly not a liability of the current operator.

3.6

Permits and Authorizations

All permits and authorizations necessary to conduct the currently proposed work at the Granite Creek Underground Project have been identified

and are already in hand. The project holds a Plan of Operations from the BLM, a Class II Air Quality Operating Permit and Mercury Operating Permit to Construct, two Water Pollution Control Permits (covering the Granite Creek Mine and the Rapid

Infiltration Basins), two Nevada Reclamation Permits, a Mining Stormwater General Permit, an Onsite Sewage Disposal System Permit, a Hazardous Materials Permit, and an approved Waters of the United States Jurisdictional Determination from the USACE

valid through November 2030. These permits, issued by the BLM, the Nevada Division of Environmental Protection (through its Bureaus of Mining Regulation and Reclamation, Air Pollution Control, and Water Pollution Control), and other applicable

agencies, are all currently in good standing. There are no permitting actions pending before either the NDEP or the BLM, and the site is up-to-date on all monitoring

obligations tied to these permits.

Based on the information provided, no significant factors or risks were identified that would affect

Osgood’s right or ability to perform the proposed work. All permits required for current operations are held and in good standing, with no permitting actions pending before the BLM or NDEP for the current underground operations, and the site

is meeting all associated monitoring obligations. The Property’s mixed public / private land tenure is already reflected in the jurisdictional split between the BLM and the state.

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4

Accessibility, Climate, Local Resources, Infrastructure and Physiography

4.1

Topography, Elevation and Vegetation

The Project lies in the Basin and Range Province, a structural and physiographic province comprised of generally north to north-northeast

trending, fault bounded mountain ranges separated by alluvial filled valleys. The Project is located on the eastern flank of the Osgood Mountains. Topography is gentle to moderate at the Project, ranging from an elevation of 1,475 m (4,840 ft) at

the mine offices to 1,676 m (5,500 ft) at the crest of the historic CX pit highwall, and rises steeply to the west over 1,158 m (3,800 ft) to Adam Peak at the top of the Granite Creek drainage. Vegetation is typical of the high desert with sagebrush

on the alluvial fans, and juniper on the mountain slopes. Climate, physiography, and local topography do not present any material constraints to underground mining, surface infrastructure development, or year-round operations at the Project. Water

rights held by OMC are sufficient for its anticipated underground operations.

4.2

Means of Access

The Project can be reached by traveling east on US Interstate Highway 80 from Winnemucca to the Golconda exit, about 24 km (15 miles), then

following Nevada State Route 789 and the Getchell Mine Road northeast about 33.0 km (20.5 miles) to the Granite Creek Mine access road. The last 7.2 km (4.5 miles) is unpaved, well maintained gravel road. Areas of Route 789 are designated open

range, and travelers must watch for cattle. The Granite Creek security facility is located about 0.3 km (0.2 miles) west of the intersection with Route 789. Traveling from the east, the Golconda exit lies west of Battle Mountain about 58.7 km (36.5

miles) on Interstate Hwy 80. The route from Battle Mountain passes i-80’s Lone Tree facility, which lies south of the interstate about 31 km (19 miles) west of Battle Mountain. Access to the Granite

Creek Project is available year-round, with only minor, short-duration disruptions possible during severe winter snowstorms or summer lightning events.

4.3

Climate and Length of Operating Season

The climate in Humboldt County is typical of the high-desert environment. Typical summer temperatures average roughly 24°C (75°F) with

occasional day / night extremes of 41°C (105°F) / 4°C (40°F). Winter temperatures average roughly -1°C (30°F) with occasional day / night extremes ranging 60°F (16°C) / -12°C (-10°F). Average annual precipitation is about 8 inches, the majority of which accumulates as snowfall during the winter months. Typical snow accumulation is

roughly 3 inches on average at lower elevations, although occasional large storms may accumulate significantly more for short durations. Mining operations are able to continue year-round with brief pauses for summer lightning storms or unusually

heavy winter snowstorms.

4.4

Infrastructure Availability and Sources

Existing infrastructure at the Project includes an office building, dry and warehouse facilities, and a lined stockpile area on the surface.

Over 2,743 m (9,000 ft) of underground workings have been completed. The mine is accessed through either of two portals, and dual egress has been established for most areas of the mine; where dual egress is not possible, rescue chambers have been

installed. Equipment is repaired in an underground mine shop. Air doors and a ventilation fan provide required air supply to the workings in compliance with Mine Safety and Health Administration (MSHA) standards. Landline telephone and digital

subscriber line service are available at the Project site; cellular phone service is also available, but is dependent on the strength of receiving antennas, topography, and lines of sight.

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4.4.1

Water

Six deep dewatering wells have been drilled and cased, four of which are currently in operation, with the remaining two scheduled to come

online in Q2 2026. Dewatering water not consumed by operations is discharged through pipelines to four permitted rapid infiltration basins (RIBs), two of which have been constructed to date; the existing RIBs and associated pipelines re-infiltrate water produced in mine dewatering into the valley aquifer.

4.4.2

Electricity

Electrical infrastructure suitable for mine operations is installed at the Project.

4.4.3

Personnel

The town of Winnemucca has a population of about 8,400. Some experienced and general labor is available locally and from other small towns in

the region. There are several mining operations in the region and, as such, there is always competition for employees.

4.4.4

Supplies

Basic services are available in Winnemucca. Local mining districts have been active since the 1980’s, and mining suppliers and

contractors are accustomed to working in the area.

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5

History

5.1

Previous Operations

The Property has been explored by several individuals and mining / exploration companies since the late 1930’s. The original discovery on

the Property was made by Clovis Pinson and Charles Ogee in the mid to late-1930’s, but production did not occur until after World War II, when ore from the original discovery was shipped to and processed at the Getchell mine mill. In 1949 and

1950, total production from the Granite Creek mine amounted to approximately 9,071 t (10,000 st) grading approximately4.8 g/t (0.14 oz/t) gold.

The Property has had a succession of operators since exploration resumed in 1970: the Cordex I Syndicate and its successor, Pinson Mining

Company (PMC), operated the Property from 1970 through the mine’s closure in 2000; Homestake and Barrick held a 50/50 interest in PMC and conducted exploration from 1996 to 2003; Atna Resources Ltd. earned into and then acquired PMC’s

joint venture interest between 2004 and 2011 and operated the underground mine from 2012 until it was placed on care and maintenance in June 2013; Osgood Mining Company LLC (OMC), a wholly owned subsidiary of Waterton Global Resources Management,

acquired the Project in 2016; and i-80 Gold Corp acquired OMC in April 2021 and has operated the Project since.

Historically, the Granite Creek Mine Project, with small additions from the nearby Preble and Kramer Hill mines, was credited with gold

production in excess of 1 million ounces and less than 100,000 oz of silver (Tingley, 1998). PMC independently compiled a record of production and credited the Granite Creek Mine Property with production of 986,000 oz of gold through 1999.

Historic production figures for the Granite Creek Mine are compiled from a combination of company records and published sources and may be incomplete or approximate; these figures are presented for historical context only and should not be relied

upon as an indication of current Mineral Resources or Mineral Reserves.

i-80 previously completed

an Initial Assessment (IA) on the Granite Creek Open Pit Project, reported March 26, 2025; the status of that prior study relative to this Technical Report Summary is addressed in the Report Version Update (Section 2.5).

5.2

Exploration and Development of Previous Owners or Operators

The Cordex I Syndicate leased the Property in 1970 and, following a 1971 surface mapping and sampling program, completed 17 reverse circulation

(RC) drillholes in and around the 1940s-era Granite Creek Mine pit, confirming low-grade gold values. An 18th step-out hole

intercepted 27.4 m (90 ft) of 5.8 g/t (0.17 oz/t) gold, forming the basis for delineation of the ‘A’ Zone. Cordex and its successor, PMC, explored the Property largely through mapping and geochemical sampling, including a regional

mapping program from Preble to Getchell in the late 1970s, a 1:6,000-scale mapping program of the Property in 1983, and 1:2,400-scale mapping of the pit areas through

the active life of the mine. PMC began developing the A Pit in 1980 and produced gold in 1981; the B Pit followed in 1982. Step-out drilling in 1982-1983 identified the C and CX Zones northeast of the A Zone,

and drilling in 1984 identified an independent fault system that became the core of the Mag deposit, which entered production in 1987. PMC produced from the CX, CX West, and Mag pits into the mid-to-late 1990s, until falling gold prices and erratic mill feed forced closure of the oxide mill in early 1998; all active mining ceased in January 1999 and the Project was officially closed in May 2000.

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Homestake and Barrick, having become 50/50 partners in PMC, conducted an exploration program

from 1996 to 2000, expending approximately US$12 million exploring the deeper feeder fault zones of the Property for a large, high-grade gold system. The program identified gold mineralization with underground grades but did not identify a

deposit of sufficient size to be of development interest, and the partners concluded the program. Barrick acquired Homestake in 2003 and drilled three additional exploration drillholes.

Atna Resources Ltd. acquired an option in August 2004 to earn a 70% joint venture interest in the Property from PMC, completing its earn-in in 2006 after expending the required US$12 million. PMC elected to back-in for an additional 40% interest in April 2006 and spent over US$30 million on

surface and underground diamond core and RC drilling, underground drifting, and surface infrastructure (including rapid infiltration basins, a mineralized material stockpile pad, and underground electrical upgrades) before completing its claw-back

in early 2009, at which point PMC held 70% and Atna 30% of the joint venture. In September 2011, Atna acquired PMC’s 70% interest outright, along with an evergreen processing agreement with Barrick for processing underground refractory ores at

Barrick’s Goldstrike facilities. Underground development commenced in early 2012, and mine ramp-up began in late 2012; a total of 1,832 m (6,011 ft) of primary and secondary development was completed

during 2012 and 2013, and approximately 27,216 t (30,000 st) of ore containing 7,900 oz of gold were mined and shipped to off-site processing facilities. The mine was placed on care and maintenance in June

2013, driven primarily by the steep decline in gold prices that year; in May 2014, the underground mine was changed to an intermittent production status to test revised stoping methods and prove mining economics at small production rates.

Osgood Mining Company LLC (OMC), a wholly owned subsidiary of Waterton Global Resources Management, acquired the Project in 2016 and completed

drillhole database compilation and verification campaigns, including migration of the Atna database to Maxwell Datashed software (2017) and further verification (2018). In 2016, OMC completed a project-scale structural geology study —

surface and underground mapping, historical data review, and cross-section interpretation — that formed the basis of an updated 3D litho-structural model used for the 2020 Mineral Resource estimate (AMC, 2020). From 2017 to 2018, OMC also

completed an extensive drill material inventory and salvage program securing available drill core and RC chips. OMC maintained environmental permit compliance throughout its ownership and continued to maintain and improve site infrastructure,

including a third-party hydrology and dewatering review that led to pump replacement in 2019 and upgraded dewatering well process controls. In April 2021, i-80 Gold Corp — created as a spinout of Premier

Gold Mines Limited’s Nevada assets — acquired OMC from Waterton Global Resources Management, and additional land was purchased the following month. Further land acquisitions followed in 2021 (Section 31 fee land and Christison interests)

and 2022 (Section 21 fee land and BEE DEE unpatented claims lessee interest).

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6

Geological Setting, Mineralization, and Deposit

6.1

Regional, Local and Property Geology

6.1.1

Regional Geology

The Property is located on the eastern flank of the Osgood Mountains within the Basin and Range tectonic province of northern Nevada (Figure 6-1). The Granite Creek Mine, together with the Preble, Getchell, Turquoise Ridge, and Twin Creeks mines, lies on the Getchell gold trend, which generally strikes northeast-southwest and has been cross-cut by secondary north-south and northwest-southeast-trending structures. Deposits along the trend are hosted in Paleozoic marine sedimentary rocks exposed in the Osgood Mountains, which have been complexly

thrust faulted and intruded by the Cretaceous-aged (92 Ma) Osgood Mountains granodiorite stock. Gold mineralization at Granite Creek is characteristic of a Carlin-type gold system, similar to other deposits along the Getchell Trend. These units are

unconformably overlain by Miocene volcanic rocks.

The Osgood Mountains Range is underlain by Cambrian Osgood Mountain Quartzite, Cambrian

Preble Formation, Ordovician “Comus” Formation and the “upper plate” Valmy Formation. These units are unconformably overlain by the Permian Etchart Formation (Antler Peak Equivalent) of the Roberts Mountains overlap

assemblage, and by the Triassic Golconda allochthon. These uppermost units form a belt of outcrops flanking the western and northern sides of the Osgood Range. These rocks have been intruded by the Cretaceous-aged Osgood Mountains granodiorite

stock, which forms the core of the Osgood Mountains. Stratigraphy throughout the Osgood Mountains plunges north (Chevillon, et al., 2000). A significant thermal metamorphic aureole surrounds the stock. At least four Paleozoic units, defined by

structure, lithology, and age comprise the Osgood Mountains (McLachlan, et al., 2000). These include the:

•

Autochthonous Cambrian Osgood Mountains Quartzite and Preble Formation and Cambrian to Ordovician Comus Formation

•

Allochthonous Ordovician Valmy Formation, part of the Roberts Mountains allochthon

•

Antler overlap sequence including the Mississippian Goughs Canyon Formation, Pennsylvanian Battle Formation, and

Pennsylvanian-Permian Etchart Limestone

•

Allochthonous Pennsylvanian-Permian Farrel Canyon Formation, part of the Golconda allochthon

The autochthonous Cambrian-Ordovician package has been described by Jones (1991, cited in McLachlan et al. 2000) and

comprises the Osgood Mountains Quartzite, Preble Formation, and Comus Formation. All these units have undergone regional metamorphism and intense, northwest-directed folding (McLachlan, et al., 2000). At the Getchell Project, these two units are

folded together to form the northwest-verging Pinson anticline. The Comus and Preble Formations show distinct facies changes across the district. These units at Turquoise Ridge and Twin Creeks contain tuffs, pillow basalts, and mafic sills, none of

which are present in the same units at Granite Creek.

The Roberts Mountains allochthon described by Stenger et al. (1998) is exposed

at the Turquoise Ridge and Twin Creeks mines where it has been mapped as the Valmy Formation. The Roberts Mountains allochthon is composed of a thick (299 m (>980 ft)) sequence of mid-ocean ridge

basalts and intercalated pelagic sediments that have been thrust over the Twin Creeks member of the Comus Formation (Stenger, et al., 1998). This sequence has not been identified at the Granite Creek Mine but was likely present and eroded prior to

the present day.

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The Antler overlap sequence in the Osgood Mountains consists of the Pennsylvanian Battle

Formation, and Pennsylvanian-Permian Etchart and Adam Peak formations (McLachlan, et al., 2000). The Battle Conglomerate consists of cobbles and quartzite pebbles. The Etchart lies conformably on the Battle and consists of calcareous sandstone

underlying fossiliferous limestone. South of the Getchell Project, the Battle and Etchart lie unconformably on the Preble Formation and Osgood Mountain Quartzite (McLachlan, et al., 2000). These units are not present at the Granite Creek Mine

Project.

The Golconda allochthon comprises the Mississippian Goughs Canyon and the Pennsylvanian Permian Farrel Canyon formations present

along the northwest flank of the Osgood Mountains. The thrust strikes north to northeast from the central part of the range to the Dry Hills in the north (McLachlan, et al., 2000). These units are not present at the Granite Creek Mine Project.

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Source: SRK, 2026

Figure 6-1: Regional Geology Map

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6.1.2

Local Geology

The geology throughout the Osgood Mountains is typified by folded

Cambrian-to-Ordovician sedimentary rocks intruded by Cretaceous stocks and cross-cut by later high-angle structural deformation

related to Basin and Range extension. The older rocks are overlain by Miocene andesitic basalt, with surrounding fault-bounded basins filled by Quaternary alluvial gravel. The Osgood Mountains trend generally northeast, although at a structural

hinge near the Granite Creek Mine the east flank of the range rotates and trends north toward the Getchell mine (Figure 6-2). Gold mineralization is primarily hosted by fine-grained marine sedimentary rocks

overlying a large stock of Cretaceous granodiorite. Cambrian-to-Ordovician siliciclastic and carbonate rocks have been intruded by the Cretaceous Osgood Mountains

granodiorite, forming large metamorphosed aureoles with several tungsten-bearing skarns, two of which occur on the Property. The Cambrian Osgood Mountains Quartzite is overlain by the Preble Formation and, in turn, the Ordovician Comus Formation;

both have been folded into a broad, north-plunging anticline — the Pinson Anticline — which is northeast-plunging and northwest-verging, extends approximately 4.8 km (3 miles) southwest from the Granite Creek Mine, and is cored by the

Preble Formation and flanked by the Comus Formation. Gold mineralization on the Property is primarily hosted in the Comus Formation and exhibits strong structural control, with the most important structural feature being the network of faults

bordering the escarpment marking the southern and eastern edge of the Osgood granodiorite. This fault system divides into three structural and stratigraphically mineralized zones, the Rangefront, CX, and Mag Zones, each defined by one or more major

structural elements.

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Source: i-80, 2026

Figure 6-2: Local Geology Map

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The geology throughout the Osgood Mountains is typified by folded Cambrian-to-Ordovician sedimentary rocks intruded by Cretaceous stocks and cross-cut by later high-angle structural deformation related

to Basin and Range extension. The older rocks are overlain by Miocene andesitic basalt, with surrounding fault-bounded basins filled by Quaternary alluvial gravel. The Osgood Mountains trend generally northeast, although at a structural hinge near

the Granite Creek Mine the east flank of the range rotates and trends north toward the Getchell mine. Gold mineralization is primarily hosted by fine-grained marine sedimentary rocks overlying a large stock of Cretaceous granodiorite. Cambrian-to-Ordovician siliciclastic and carbonate rocks have been intruded by the Cretaceous Osgood Mountains granodiorite, forming large metamorphosed aureoles with several

tungsten-bearing skarns, two of which occur on the Property. The Cambrian Osgood Mountains Quartzite is overlain by the Preble Formation and, in turn, the Ordovician Comus Formation; both have been folded into a broad, north-plunging anticline

— the Pinson Anticline — which is northeast-plunging and northwest-verging, extends approximately 4.8 km (3 miles) southwest from the Granite Creek Mine, and is cored by the Preble Formation and flanked by the Comus Formation. Gold

mineralization on the Property is primarily hosted in the Comus Formation and exhibits strong structural control, with the most important structural feature being the network of faults bordering the escarpment marking the southern and eastern edge

of the Osgood granodiorite. This fault system divides into three structural and stratigraphically mineralized zones — the Rangefront, CX, and Mag Zones — each defined by one or more major structural elements.

6.1.3

Property Geology

In 2022, i-80 Gold Corp created an updated comprehensive geologic model for the Granite Creek Mine

Project, incorporating surface and underground mapping, structural analyses, geologic interpretation, and a complete 3D geologic model built in Leapfrog Geo using all available drillhole, surface and underground mapping, televiewer, and structural

analysis data. Structure at the Project is highly complex, reflecting multiple deformation events; regional deformation events (the Antler, Sonoma, and Elko orogenies) are likely responsible for the numerous overprinted fabrics and compressional

structures observed, which appear to have been dissected and/or reactivated by subsequent Basin and Range extension. i-80 recognizes a west-northwest-verging imbricate thrust system comprising the Rangefront,

Adam Peak, Otto, and CX faults, interpreting most of the compressional, westward transport along these faults to have pre-dated emplacement of the Osgood stock. The main structural element on the Property is

the Rangefront fault (strike 010° to 045°, dip 60° near-surface shallowing with depth); the Adam Peak and Otto faults are hanging wall splays off its northern extent, and the CX fault (strike 050°, dip 55°) lies to the east. The

Mag fault system, on the eastern portion of the Property, trends 335° and is a younger fault system associated with Tertiary extension, with the Mag and Mag West faults forming a horst block (Figure 6-3).

Significant named structural features on the Property include:

Rangefront Zone (RFZ)

A northeast-trending fault zone forming a broad, persistent zone of shearing and brecciation along the Range Front Fault, involving the entire

stratigraphic sequence at the Property, including the Cambrian Preble, Ordovician Comus, and Cretaceous granodiorite.

Range Front

Fault (RFF)

A prominent 010° to 045° striking normal fault defining the eastern front of the Osgood Mountains, juxtaposing

the Comus Formation in the hanging wall against the Preble Formation in the footwall. Originated as a west-verging thrust fault and has been reactivated as a down-to-the-east normal fault of significant, unknown displacement, with a near-surface dip of 60° that shallows with depth.

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Adam Peak Fault

A 048° striking hanging wall splay off the Rangefront fault with a near-surface dip of 72° that shallows with depth, reactivated as a

normal fault of unknown displacement.

Otto Fault

A 040° striking hanging wall splay off the Rangefront fault with a near-surface dip of 80° that shallows with depth, defined by a zone

of discrete anastomosing splays; its on-strike and down-dip extent defines the SPZ fault system.

CX West Fault

A

younger offsetting normal fault (strike 245°, dip 70°) with approximately 46 m (150 ft) of displacement, offsetting stratigraphy as well as the Adam Peak and Otto faults in a down-to-the-northwest direction.

Ogee Fault

A 10 to 30 m (100 ft) wide fault zone (strike 070°, dip 85°) juxtaposing the Upper and Lower Comus formations, defined by an

anastomosing system of splays and interpreted as a long-lived accommodation zone with multiple phases of reactivation, including recent right-lateral oblique-normal motion of unknown displacement.

Linehole Fault

A

through-going southwest-trending normal fault dipping 85° to the northwest with approximately 30 m (100 ft) of displacement, acting as a structural and mineralization boundary on the Property.

LH Fault

A splay

off the Linehole fault (strike 030°, dip 80°) with approximately 15 m (50 ft) of normal-sense displacement; its intersection with the Ogee fault is one of the most important and prolific structural intersections on the Property.

CX Fault

A

complex zone of brittle fracturing juxtaposing Upper Comus argillite against limestone beds of the Lower Comus, striking approximately 035° to 045° and dipping 55° to 65° southeast. Originated as a west-verging thrust and has been

reactivated as a normal fault with an unknown amount of down-to-the-southeast displacement.

SPZ Fault System

Comprises the along-strike and down-dip extent of the Otto fault and its associated splays, trending

northeast with a dip of approximately 50° to the southeast. The Upper and Lower Comus formations are often juxtaposed along this suite of faults, with a transition from weakly metamorphosed rock in the hanging wall to more strongly metamorphosed

rock in the footwall.

Mag Fault

A younger, through-going normal fault (strike 340°, dip 75°) related to Basin and Range extension, with unknown but

interpreted-significant down-to-the-east displacement.

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Mag Fault System

A northwest-trending suite of brittle faults defining the Mag pit; its two main faults, the east-northeast dipping Mag fault and the

west-southwest dipping Mag West fault, form a horst block within which mineralization is concentrated.

Mineralization at Granite Creek is

structurally controlled, with faults the primary control especially in high-grade underground zones, while lithologic contacts, bedding, and folds play an important role in near-surface (open pit) mineralization. High-grade mineralized zones are

moderately continuous along faults, with the most prolific zones at structural intersections. Gold mineralization occurs within pyrite in two stages — an early

non-ore-pyrite stage and a gold-bearing arsenian pyrite stage — and correlates with arsenic, antimony, mercury, and thallium.

Gold mineralization is primarily hosted by the Upper and Lower Comus Formations (argillite and interbedded argillite/limestone, respectively):

the Upper Comus is the primary host in the Mag Zone and hosts most surface resources, and is also locally mineralized within the B, C, CX, CX West, and portions of the RFZ; the Lower Comus hosts most of the high-grade underground resources. Near the

Osgood Mountains stock, including in the underground resource areas, host rock has been metamorphosed (argillite to hornfels, limestone to garnet/pyroxene/wollastonite/marble), and higher gold grades typically occur in these metamorphosed rocks

along fault zones. The Preble Formation is a poor host for gold mineralization but contains localized gold concentrations where brecciated and adjacent to major fluid conduits. Oxide mineralization (limonite, hematite, and other iron/arsenic oxides)

is extensive in the Ogee Zone and CX Fault system to a depth of 457 m (1,500 ft), and more variable within the RFF system, ranging from less than 152 m (500 ft) to 549 m (1,800 ft).

Dimensions and characteristics of the significant mineralized zones on the Property are as follows:

Mag Pit

Hosted by

Upper Comus Formation argillite; the mineralized zone is north-northwest oriented, sub-parallel to the Mag Fault, dips east-northeast, and plunges south-southeast. The tabular body has a strike length of

approximately 1,219 m (4,000 ft), varies from 61 to 122 m (200 to 400 ft) in width, and has an average down-dip extent of 137 m (450 ft). Mineralization is more disseminated and lower grade than the

Rangefront, CX, and Ogee zones, and is spatially associated with decarbonatization, kaolinization, silicification, and quartz veinlets.

Otto-Adam Peak Zone

Defined by the Otto and Adam Peak faults and their associated splays; trends northeast, dips southeast, plunges east-northeast, and is offset down-dip by the CX West fault. Mineralization is moderately continuous, controlled by a network of discrete anastomosing faults and splays within the Lower Comus Formation, with high-grade mineralization at fault

intersections. Strike length approximately 152 m (500 ft), vertical extent 213 m (700 ft), average width 23 m (75 ft).

Ogee Zone

An east-northeast trending, near-vertical mineralized zone controlled by the Ogee Fault and associated splays; the upper portion

plunges east-northeast at 55° and is strongly oxidized, while the lower portion is near vertical, mostly oxidized, and contains more comingled sulfide. Strike length 122 m (400 ft), vertical extent 457 m (1,500 ft), average width 23 m (75 ft).

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South Pacific Zone (SPZ)

A northeast-trending, southeast-dipping zone of high-grade, fault-bound mineralization with a northeast plunge of 45°, controlled by the

along-strike and down-dip extent of the Otto fault; highest grades are concentrated along faults juxtaposing the Upper and Lower Comus Formations. Strike length 381 m (1,250 ft),

down-dip extent 274 m (900 ft), average width 7.6 m (25 ft).

CX Zone

Consists of both near-surface (open pit) and higher-grade underground mineralization. The near-surface portion is controlled by the

through-going CX fault and its splays, is discontinuous, and is associated with pervasive argillization and decarbonatization within the Lower Comus Formation (strike length 1,066 m (3,500 ft), down-dip extent

122 m (400 ft), average width 23 m (75 ft)). The higher-grade underground portion is more tightly structurally controlled along the down-dip section of the CX Fault (strike length 305 m (1,000 ft), down-dip extent 366 m (1,200 ft), average width 12 m (40 ft)).

Rangefront Zone (RFZ)

Pervasively argillized and decarbonatized with intense brecciation along the lower-bounding Range Front Fault; mineralization occurs as

discontinuous amorphous bodies within the Comus Formation, with high-grade zones concentrated in the Lower Comus and anomalous mineralization present in the Preble Formation proximal to the RFF. Strike length approximately 290 m (950 ft), down-dip extent 335 m (1,100 ft), average width 30 m (100 ft).

Multiple areas of high-grade gold

mineralization are amenable to underground mining methods, including the Otto-Adam Peak, Ogee, SPZ, CW, and Rangefront zones; the current underground operation mines the Otto-Adam Peak and Ogee zones, with 2025 exploration focused on the SPZ,

together forming the underground Mineral Resource at the Project.

Alteration assemblages at Granite Creek include silicification,

decarbonatization, pyrite, and remobilization of carbon. In the CX Zone (encompassing the A, B, C, CX, and CX West pits), alteration style and intensity vary gradationally: in the B Pit, gold mineralization occurs in fractured shale and silty

carbonate that is weakly silicified and clay-altered; in the A Pit, alteration consists of intense silicification of carbonate lithologies and gold-rich jasperoid formation along structures, with gold grains typically less than 5 microns hosted as

inclusions in arsenian pyrite; in the C Pit, high-grade material is hosted in decarbonatized carbonates cross-cut by small faults. In the CX Pit, mineralization consists of silica and pyrite replacing

carbonate along narrow structures; in the CX-West Pit, mineralization is hosted in strongly calc-silicate carbonates with strong argillic alteration. In the Mag Pit, mineralization is associated with

decarbonatization, kaolinization, silicification, and quartz veining, with silicification occurring as replacement of decalcified lithologies and healing of fault gouge and breccia. The RFF Zone displays pervasive argillization and decarbonatization

with dissolution collapse breccias and intense shearing, with minor silicification as a broad overprint and calcite veining along the RFF margins (Figure 6-4)

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Figure 6-3: Property Geology Map

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Source: Chadwick 2002

Figure 6-4: Alteration of the Mag Pit

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6.2

Mineral Deposit

The structural setting, alteration mineralogy, and mineralization characteristics of Granite Creek are consistent with Carlin-type deposits.

Carlin-type deposits formed in the mid-Tertiary after the onset of extension in an

east-west-trending, subduction-related magmatic belt. The deposits are located along long-lived, deep crustal structures inherited from Late Proterozoic rifting and the formation of a passive margin within Paleozoic carbonate sequences composed of

silty limestone to calcareous siltstone. The carbonate sequences are overlain by either structurally controlled siliciclastic sequences controlled by the Early Mississippian-aged Roberts Mountain allochthon or by stratigraphically controlled

siliciclastic sequences. The siliciclastic rocks are less permeable than the underlying carbonate rocks, which traps fluids along major structures, causing them to flow laterally into the permeable and reactive carbonate sequences.

Alteration of host carbonate sequences consists of decarbonatization, argillization, and selective silicification, forming jasperoid and

causing carbon flooding. Gangue minerals in Carlin-type deposits consist of calcite, siderite, and ferroan dolomites that can occur as geochemical fronts beyond the mineralized zones.

Gold deposition occurs in arsenian pyrite, is hosted within carbonaceous sequences near major high angle structural zones, and is concentrated

in structural traps and/or replacement horizons of reactive and permeable sedimentary beds.

Carlin-type deposits typically show enrichment

in antimony, arsenic, mercury, and thallium, caused by hydrothermal fluids with temperatures ranging from 180°C (356°F) to 230°C (446°F). The source of fluids is likely deep-seated magmas that release gold-bearing fluids at depths of

10 to 12 km (7 mi). These magmas formed during Eocene slab-rollback of the Farallon plate as upwelling asthenosphere impinged on a strongly metasomatized sub-continental lithospheric mantle. Tertiary dikes

associated with mineralization and radiometric age dates between 39 and 42 Ma, along with isotopic data, provide evidence toward this hypothesis. This genetic model represents the prevailing scientific interpretation for Carlin-type gold systems and

is based on regional observations and published studies; alternative interpretations may exist.

Structural pathways, reactive rocks, and

sources of heat, gold, sulfur, and iron are required for Carlin-type deposits to form. Large regional structures transecting reactive rocks create contacts, faults, and shears; these secondary structures create pathways and traps for hydrothermal

and metalliferous fluids. Mineralization at the Granite Creek Project exhibits the key geological, mineralogical, geochemical, and structural characteristics of Carlin-type gold deposits as described above.

6.3

Stratigraphic Column and Local Geology Cross-Section

The stratigraphy of the Osgood Mountains, from youngest to oldest, comprises Quaternary alluvium and basalt; Tertiary andesite/basalt flows,

rhyolitic tuffs, chert-shale-rhyolite clastics, and dacite/andesite dikes; the Cretaceous Osgood Mountains granodiorite/quartz diorite stock; Permian-Pennsylvanian Havallah Formation and Etchart Limestone; Pennsylvanian Battle Formation; and

Ordovician Valmy and Comus Formations overlying the Cambrian Preble Formation and Osgood Mountain Quartzite. Gold mineralization at the Property is primarily hosted in the Comus Formation, as shown in the stratigraphic column below (Figure 6-5) and the representative cross-section (Figure 6-6), which together illustrate the geometry of mineralization-controlling structural features, including faults and the

lithologic contact between the Upper and Lower Comus Formation.

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Figure 6-5: Granite Creek Stratigraphic Column

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Figure 6-6: Cross-section A-A’ looking Northeast

showing Structure, Lithology and Mineralization

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7

Exploration

Exploration techniques employed on the Property to define additional gold resources have consisted primarily of mapping, geochemical sampling,

geophysical surveys, and drilling. Use of these methods has resulted in the discovery of approximately one million ounces of gold in several open pit deposits. Atna became involved in Project planning in July 2004 and began drilling the Property in

August 2004 after execution of the earn-in agreement with PMC; i-80 Gold has not performed any new geophysical or surface geochemistry work. Exploration work other than

drilling is discussed in Section 7.1; drilling is discussed in Section 7.2; and hydrogeology and geotechnical data relevant to underground mine development are discussed in Sections 7.3 and 7.4.

7.1

Exploration Work (Other Than Drilling)

7.1.1

Procedures and Parameters Relating to the Surveys and Investigations

Cordex and its successor, PMC, explored the Property through geologic mapping and geochemical sampling, including a regional mapping program

from the Preble to the Getchell mines in the late 1970’s, a 1:6,000-scale mapping program of the Property in 1983, and a 1:2,400-scale mapping program of the

Pinson pit area through the active life of the mine. Bench mapping in the pits occurred during mining and was followed by detailed 1:1,200-scale mapping of the A, B, C, CX, Mag, CX West, and Blue Bell pits by

Tom Chadwick starting in 2000, completed under the Homestake / Barrick partnership agreement. In 2016, OMC contracted Robert Leonardson to complete a geological study focused on advancing the understanding of the structural framework and providing

exploration targeting guidance, including structural and geologic mapping of the open pits and underground exposures and construction of property-wide cross-sections. Numerous geophysical surveys have also been conducted on the Property, including

regional gravity and aeromagnetic surveys and detailed electromagnetic techniques — Induced Polarization (IP), Electromagnetics (EM), Magnetotellurics (MT), and Controlled Source Audio-frequency Magnetotellurics (CSAMT). A detailed gravity

survey was conducted by Magee Geophysical Services, LLC of Reno, Nevada in October 2006 (Figure 7-1), acquiring 2,587 gravity readings on a 100 m (328 ft) station spacing over approximately 27 square km (10

square miles); the results were interpreted by Fritz Geophysics in 2007.

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Source: Magee, 2006

Figure 7-1: Gravity Survey, 2,587 Stations, Magee Geophysical Services, 2006

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In 2002, Quantec Geoscience was contracted to acquire TITAN 24 MT data over the Pinson

property. Six east-west lines were collected along the Rangefront, spaced on average around 610 m (2,000 ft) apart, with a dipole spacing along line of 91 m (300 ft). Quantec ran regular two-dimensional

inversions on the MT data to create resistivity depth sections (Figure 7-2).

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Figure 7-2: Location of the MT Survey Lines on the Geology and Pit Locations (Left) and on the

Residual Gravity (Right)

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7.1.2

Sampling Methods and Sample Quality

Cordex collected 737 rock chip samples in conjunction with its mapping programs, assayed for gold, silver, arsenic, antimony, and mercury, with

select samples also analyzed for lead, zinc, copper, and manganese; the combined mapping/sampling programs were responsible for the discoveries of the Blue Bell and Felix Canyon deposits. PMC completed six float chip geochemical grids consisting of

8,756 samples covering the Mag deposit and along strike south of the A and B Pits. A biogeochemical sagebrush sampling program was conducted in the 1990s with inconclusive results. Under the Homestake/Barrick joint venture, an additional 312 rock

samples and 273 soil samples were collected on strike south of the existing pit areas and west of the A, B, C, and CX Pits. For underground channel sampling, Atna channel-sampled 14 ribs in the Ogee Zone and sent 74 rib and face samples out for

assay.

7.1.3

Information About the Area Covered

Historic mapping and geochemical sampling programs covered the Property from the Preble to the Getchell mines at regional scale, narrowing to 1:6,000-scale coverage of the Property (1983) and 1:2,400- to 1:1,200-scale coverage of individual pit areas through the active

mine life. The 2006 Magee gravity survey covered approximately 27 square km (10 square miles) on a 100 m (328 ft) station spacing (2,587 stations total). The 2002 Quantec TITAN 24 MT survey comprised six east-west lines along the Rangefront, spaced

on average around 610 m (2,000 ft) apart, with a 91 m (300 ft) dipole spacing along each line. Underground channel sampling in the Ogee Zone covered 14 ribs.

7.1.4

Significant Results and Interpretation

Mr. Leonardson’s 2016 structural study concluded that potential targets for additional gold mineralization occur at intersections of

the east-dipping, north-south faults (Rangefront/Mag) with the southeast-dipping CX-type faults, and at intersections of sub-vertical northwest-striking faults with CX-type faults; zones of limestone decarbonatization, such as those seen in the CX Pit, are also potential hosts for gold mineralization, indicating strong fluid/vapor flow through the rock mass. Identified areas

for further exploration include the intersection of the SOS and JP dikes on the south wall of the CX Pit (the largest block of decarbonatization on the Property); the Ogee pipe extension 457 to 549 m (1,500 to 1,800 ft) below the CX-C Pit, where historical hole HPC-070A intersected a 232 m (760 ft) interval of

low-to-moderate gold grades with high-grade mineralization near the proposed Ogee high-grade down-dip extension; the northern

continuation of the fault-propagated anticline in the western Mag Pit; the intersection of the Adam Peak Fault and the Mag Fault suite north of the Mag Pit; the CX-B Pit decarbonatization zone; and

decarbonatization zones on the Mag Pit west wall.

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Source: Fritz Geophysics, 2007

Figure 7-3: Pinson Local Gravity Interpretation

The existence of about 1,700 drillholes within the gravity survey area allowed a novel approach for interpreting the detailed gravity data

(Figure 7-3), correlating basement rock types defined by the basement gravity response with the general mapped geology. In 2008, Barrick interpreted the geophysical survey data at Pinson, combining the 2002 MT

survey, the 2006 gravity survey, and all available geological/geochemical information to define target areas requiring drillhole testing. Figure 7-4 through Figure 7-9

show MT resistivity depth inversions for each of the six survey lines, annotated with drilling, surface geological mapping, interpreted sectional geology, and structural targets.

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Source: i-80, 2026

Figure 7-4: MT Resistivity Depth Inversion for Line 6090

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Figure 7-5: MT Resistivity Depth Inversion for Line 12300

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Figure 7-6: MT Resistivity Depth Inversion for Line 13860

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Figure 7-7: MT Resistivity Depth Inversion for Line 15300

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Figure 7-8: MT Resistivity Depth Inversion for Line 17160

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Figure 7-9: MT Resistivity Depth Inversion for Line 19230

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A small exploration drifting program was conducted on the upper ‘B’ zone by

Cordex in the 1970s to conduct bulk testing (results unavailable). In May 2005, Small Mine Development (SMD) of Boise, Idaho, was contracted by Atna to drive exploration drifts, crosscuts, and drill stations to evaluate the Range Front and CX

resource areas, completing 606 m (1,988 ft) of adit, 115 m (378 ft) of decline, and six diamond drill stations; a mineability test on the newly defined Ogee Zone extracted approximately 363 t (400 st) of material and indicated the possibility of drift-and-fill as a potential mining method. During 2008, approximately 211 m (693 ft) of development drifting was completed with significant geological data recorded in the

RFZ. Throughout 2024 and 2025, approximately 173.7 m (570 ft) of development drifting was completed for the South Pacific exploration drill program, which began June 2025 and concluded December 13, 2025, targeting SPZ mineralization with a

total of 16,531 m (54,235.7 ft) drilled to pre-feasibility-study-quality resource and reserve declaration density. Salient results of the Ogee Zone channel sample assays are summarized in Table 7-1; assays

indicated no high-grade mineralization except where the main drift intersected the Ogee Zone on the 4770 elevation.

Table 7-1: Salient Results of the Ogee Zone Channel Sample Assays

Sample No.

From ft

(m)

To ft

(m)

Length ft

(m)

Gold Grade opt

(g/t)

North Rib

RFUG-055

76 (23.1)

81 (24.7)

5 (1.5)

0.144 (4.94)

RFUG-056

81 (24.7)

85 (25.9)

4 (1.2)

0.445 (15.26)

RFUG-059

85 (25.9)

88 (26.8)

3 (0.9)

0.274 (9.39)

RFUG-061

88 (26.8)

93 (28.3)

5 (1.5)

1.448 (49.65)

RFUG-063

93 (28.3)

97 (29.6)

4 (1.2)

0.176 (6.03)

RFUG-064

97 (29.6)

101 (30.8)

4 (1.2)

0.739 (25.34)

RFUG-067

101 (30.8)

110 (33.5)

9 (2.7)

0.996 (34.15)

Weighted Average

34 (10.4)

0.682 (23.38)

South Rib

RFUG-081

77 (23.5)

80 (24.4)

3 (0.9)

0.106 (3.63)

RFUG-082

80 (24.4)

83 (25.3)

3 (0.9)

0.065 (2.23)

RFUG-083

83 (25.3)

93 (28.3)

10 (3)

1.082 (37.10)

RFUG-084

93 (28.3)

96 (29.3)

3 (0.9)

0.894 (30.65)

RFUG-086

96 (29.3)

99 (30.2)

3 (0.9)

0.355 (12.17)

RFUG-087

99 (30.2)

107 (32.6)

8 (2.4)

0.028 (0.96)

RFUG-088

107 (32.6)

112 (34.1)

5 (1.5)

0.228 (7.82)

Weighted Average

(10.7)

7.2

Exploration Drilling

7.2.1

Drilling Type and Extent

Since 1970, a total of 2,083 drillholes totaling 291,312 m (955,747.9 ft) have been drilled within the Property area by previous operators

(Figure 7-10) (drillhole data from prior to 1970 is no longer available). PMC and its predecessors, Rayrock Mines and the Cordex Syndicate, account for most of these holes (1,434 holes totaling 168,991.8 m

(554,435 ft)); Homestake drilled 165 holes totaling 48,831.3 m (160,207.7 ft) and Barrick drilled 166 holes totaling 37,195 m (122,031 ft), both acting as operators for PMC; Atna, the last company to operate at the Granite Creek Mine prior to i-80, drilled 318 holes totaling 36,293.8 m (119,074.1 ft). Table 7-2 presents a summary of this historical drilling; each period is described in further detail below.

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Figure 7-10: Granite Creek Project Drill Plan by Operator

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Table 7-2: Summary of Historical Drilling (pre i-80) on the Granite Creek Property Since 1970

Company

Surface RC

Surface Core

UG RC

UG Core

Total Holes

Total Footage

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

PMC

1,426

546,313.0

8

8,122.0

1,434

554,435.0

PMC (Homestake)

136

108,335.0

29

51,872.7

165

160,207.7

PMC (Barrick)

39

35,645.0

67

65,700.1

4

930.0

56

19,756.0

166

122,031.1

Atna

29

18,672.0

65

52,847.6

176

32,068.0

48

15,486.5

318

119,074.1

Total

1,630

708,965.0

169

178,542.4

180

32,998.0

104

35,242.5

2,083

955,747.9

Many holes drilled by PMC during this period were development holes drilled in and adjacent to existing

pits — over 1,400 holes within the A, B, C, CX, Mag, CX West, Felix, and Blue Bell pit areas, nearly all conventional rotary or RC; the eight core holes drilled (2,475.6 m (8,122 ft)) tested stratigraphy, metallurgy, or deep mineralized

structures in the B, C, CX, and Mag Pit areas (Table 7-3).

Table 7-3: PMC Drilling 1970 to 1996

Company

Surface RC

Surface Core

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Total

Footage

(ft)

Total

Meters

(m)

PMC

1,426

546,313.0

166,516.2

8

8,122.0

2,475.6

1,434

554,435.0

168,991.8

Between 1997 and 2000, Homestake, as operator for PMC, drilled 165 holes, of which 136 (33,020.5 m

(108,335 ft)) were directed into the CX and RFF system (Table 7-4).

Table 7-4: PMC – Homestake Drilling 1997 to 2000

Company

Surface RC

Surface Core

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

PMC (Homestake)

136

108,335.0

33.020.5

29

51,872.7

15.810.8

165

160,207.7

48,831.3

Four exploration holes were drilled by Barrick, operator at the time for PMC, to test extensions of the

CX Fault Zone near its projected intersection with the Mag Pit fault system; the drilling did not identify significant mineralized zones, and no additional work was conducted by Barrick (Table 7-5).

Table 7-5: PMC – Barrick Drilling 2003

Company

Surface RC

Surface Core

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

PMC (Barrick)

3

3,340.00

1,018.0

1

3,003.30

915.4

4

6,343.30

1,933.4

Atna’s 2004 drilling followed up on mineralized zones previously identified by PMC and Homestake:

31 holes totaling 9,064.6 m (29,739.5 ft), comprising four RC holes and 27 core holes, with five objectives — improving grade / thickness definition where prior drilling was RC-only, infilling where

spacing exceeded 121.9 m (400 ft), expanding mineralized zones laterally and down-dip, obtaining rock quality data, and evaluating previously identified targets. Of the 31 holes, 13 (3,962.4 m (13,000 ft))

tested the CX Fault Zone and 18 (5,102.2 m (16,739.5 ft)) tested the RFF Zone (Table 7-6).

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Table 7-6: Atna Drilling 2004

Company

Surface RC

Surface Core

Total Holes

#

Holes

Footage

(ft)

Meterage

(m)

#

Holes

Footage

(ft)

Meterage

(m)

#

Holes

Total

Footage

(ft)

Total

Meterage

(m)

Atna

4

2,217.00

675.7

27

27,522.50

8,388.9

31

29,739.50

9,064.6

The 2005 to 2006 program aimed to define and delineate Measured and Indicated gold Mineral Resources in

the upper RFF Zone, testing the zone between 1,524 and 1,341 m (5,000 and 4,400 ft) amsl using both surface and underground drilling: 107 drillholes (16,818.9 m (55,180.1 ft)) in total, including 59 surface holes (12,098.6 m (39,693.6 ft), beginning

May 2005) and 48 underground holes (4,720.3 m (15,486.5 ft), beginning September 2005) in the Ogee, CX West, and Range Front targets (Table 7-7).

Table 7-7: Atna Drilling 2005 to 2006

Company

Surface RC

Surface Core

UG Core

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Total

Footage

(ft)

Meters

(m)

Atna

25

16,455.00

5,015.5

34

23,238.60

7,083.1

48

15,486.50

4,720.3

107

55,180.10

16,818.9

In August 2007, surface exploration and development drilling began using an Eklund RC drill rig and a

Major Drilling core rig, testing portions of the CX and RFF, the Ogee Zone, and the HPR104 area north of the Granite Creek Mine — 23 surface holes (5,765.7 m (18,916.2 ft)) were completed, with disappointing results (only thin, sub-economic zones of underground mining gold grades) (Table 7-8).

Table 7-8: PMC (Barrick) Drilling 2007

Company

Surface RC

Surface Core

Total Holes

#

Holes

Footage

(ft)

Meterage

(m)

#

Holes

Footage

(ft)

Meterage

(m)

Total

Holes

Total

Footage

Total

Meterage

(m)

PMC (Barrick)

7

4,935.00

1,504.2

16

13,981.20

4,261.5

23

18,916.20

5,765.7

Surface drilling began in January 2008 with three core drills and one RC drill testing areas north of

the CX West pit and the deep potential of the Getchell Fault system; underground exploration began in April 2008, rehabilitating existing workings and driving exploration headings into the Ogee and CX zones, followed by a second surface program in

August 2008 to twin RC holes suspected of downhole contamination. Total 2008 drilling comprised 29 surface RC holes (8,342.4 m (27,370 ft)), 50 surface core holes (14,848.5 m (48,715.6 ft)), four underground RC holes (283.5 m (930 ft)), and 56

underground core holes (6,021.6 m (19,756 ft)). Eight holes drilled north of the Pinson deposit to twin earlier drilling of the Range Front/Linehole Fault intersection (the HPR104 area) could not reproduce the earlier thick low-grade intercept (considered downhole contamination, and removed from the database), though follow-up core drilling intersected thin, higher-grade mineralization

structurally controlled by the Linehole Fault/Comus contact. Two deep drillholes (BPIN-010C to 867.3 m (2,845.5 ft) and BPIN-011A to 846.7 m (2,778 ft)) tested deep structural targets identified from the 2006 gravity survey; BPIN-010C returned

negative results, while BPIN-011A indicated a narrow, low-grade zone associated with decarbonatization and pyritized sediments (Table 7-9).

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Table 7-9: PMC (Barrick) 2008 Drilling

Company

Surface RC

Surface Core

UG RC

UG Core

Total

Holes

Total

Footage

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

#

Holes

Footage

(ft)

PMC (Barrick)

29

27,370.0

50

48,715.6

4

930.0

56

19,756.0

139

96,771.6

In 2012, Atna completed four PQ-size core holes (636 m (2,086.5

ft)) to acquire column leach metallurgical samples from the Mag Pit resource area, plus 56 underground exploration RC holes (2,284.5 m (7,495 ft)) in the Ogee Zone (Table 7-10).

Table 7-10: 2012 Atna Mag Pit Core Drilling

Company

Surface Core

UG RC

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Total Footage

(ft)

Meters

(m)

Atna

4

2,086.50

636.0

56

7,495.00

2,284.5

60

9,581.50

2,920.4

Between 2012 and 2015, Atna completed 120 underground RC holes totaling 7,489.9 m (24,573 ft), designed

to confirm continuity of mineralization and delineate stope configuration within the Ogee Zone for mining (Table 7-11).

Table 7-11: 2013 to 2015 Atna Underground Development RC Drilling

Company

UG RC

Total Holes

#

Holes

Footage

(ft)

Meters

(m)

#

Holes

Total Footage

(ft)

Meters

(m)

Atna

120

24,573.00

7,489.9

120

24,573.00

7,489.9

Figure 7-11 shows the drill plan of the Property in the area of

the 2021 Mineral Resource (red outline), coded by operator and time period, with section lines for the representative vertical sections presented in Section 7.2.3. All drilling described above predates

i-80’s involvement — i-80 had not conducted drilling on the Property at the time of the 2021 resource model.

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Figure 7-11: Plan View Section Lines of Granite Creek Mine Project

The Granite Creek land position expanded in 2021, with infill drilling targeting the underground Mineral Resources. Table 7-12 lists the number of holes drilled within the current property boundary by type and operator, including holes drilled from surface and underground and Figure 7-12

throughFigure 7-14 no discoveries have been made beyond the core land package, and holes outside the core land package (all drilled by PMC) serve primarily to augment geological knowledge rather than

contribute to the resource estimate.

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Table 7-12: Drillholes Within the Current Property

Boundary by Type and Operator

Company

Core Holes

(includes RC

pre-

collar with Core Tail)

Core

Footage

RC

Holes

RC

Footage

Total

Holes

Total

Footage

i-80 Gold

438

306,929

645

99,252

1,083

406,181

Atna

113

68,334

205

50,740

318

119,074

PMC (Barrick)

123

85,456

43

36,575

166

122,031

PMC (Homestake)

29

51,872.7

136

108,335

165

160,208

PMC

8

8,122

1,426

546,313

1,434

554,435

Totals

711

520,714

2,455

841,215

3,166

1,361,929

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Source: SRK, 2026

Figure 7-12: Drilling by Hole Type Completed by PMC

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Figure 7-13: Drilling by Type Completed by PMC with Barrick as Operator

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Figure 7-14: Drilling Completed by Atna

Drilling at Granite Creek is ongoing. The holes drilled by i-80 presented here were drilled from April

2021 through January 2026, with complete assay results by February 17, 2026 (the cutoff date for the current underground resource estimate). i-80 primarily uses core drilling for sample collection, with

most surface holes pre-collared using RC down to the water table and completed with HQ-size core, focused mainly on the Ogee and SPZ zones; underground holes were all

drilled as HQ-size core, focused on the Otto, Rangefront, Ogee, and SPZ zones. A Cubex RC rig is used by ore control geologists to assist short-term mining decisions. Figure

7-15 shows holes drilled by i-80.

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Figure 7-15: Drilling Completed by i-80

7.2.2

Drilling, Sampling, or Recovery Factors

Initial surface collar locations are based on drill plan targeting, marked in the field by a geologist using a handheld GPS device loaded with

coordinates from Leapfrog Geo drill plans; drill rigs are aligned to the planned azimuth and inclination using north-seeking gyroscopic rig alignment tools, re-confirmed by the geologist. A final collar

location survey is performed by a professional contractor or i-80 Gold surveyor (UTM NAD83 Zone 11N international feet). Downhole surveys are performed by International Directional Services (IDS) of Elko on 50-foot intervals using a north-seeking gyroscopic downhole survey tool. Reverse circulation (RC) samples are collected by the drill contractor on five-foot intervals through a rotating splitter (cyclone), with

contamination minimized by regular cyclone cleaning and sample IDs/footage checked against a written sample log. Core is recovered via wire line core tube, boxed in labeled, partitioned waxed cardboard boxes in sequential order, with wooden blocks

marking

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run footage and any drilling problems (caving, voids, core tube mismatches). Core recovery is measured as the ratio of recovered core length to drilled run length and has been excellent, at 99%

or greater, with any core loss (less than 1 m (2 ft)) limited to zones with stratigraphic voids. The authors believe the drilling procedures are adequate, and there are no drilling, sampling, or recovery factors that could materially impact the

accuracy and reliability of the results.

7.2.3

Drilling Results and Interpretation

Figure 7-16 through Figure 7-19 show representative vertical

sections through the four Open Pit areas (Mag, CX/C, A, and B), and Figure 7-20 shows a vertical section through the underground resource area, illustrating drilling results by operator and time period.

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Source: i-80, 2026

Note: Black line is a topographic surface. Not all items listed in the legend are on all sections

Figure 7-16: Vertical Section A-A’ of the Mag Pit

Area

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Note: Black line is a topographic surface. Not all items listed in the legend are on all sections.

Figure 7-17: Vertical Section B-B1 of the Pit CX and C Area

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Note: Black line is a topographic surface. Not all items listed in the legend are on all sections

Figure 7-18: Vertical Section C-C1 of the Pit A Area

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Source: i-80, 2026

Note: Black line is a topographic surface. Not all items listed in the legend are on all sections

Figure 7-19: Vertical Section D-D1 of the Pit B Area

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Note: Black line is a topographic surface. Not all items listed in the legend are on all sections

Figure 7-20: Vertical Section E-E1 of the Underground

Resource Area

Example sections showing drilling in the underground resource area are shown in Figure

7-21 through Figure 7-24. Holes drilled by i-80 are labeled with hole name and shown with thicker traces; faults and mineralized

envelopes modeled at 0.1 oz Au (3 g) cut-off grade are shown for reference.

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Figure 7-21: Plan View Showing Section Locations through the Underground Resource Area

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Figure 7-22: Section A-A’ Showing Drilling in the CX

Zone, 30.5 m (100 ft) thick, looking North

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Figure 7-23: Section B-B’ Showing Drilling in the

Otto and Ogee Zones, looking North

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Figure 7-24: Section C-C Showing Drilling in the SPZ, 130 m

(100 ft) thick, looking North

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7.3

Hydrogeology

The geotechnical database developed by i-80 Gold comprises approximately 6,096 m (20,000 ft) of

geotechnically logged core from 40 drillholes, mostly since 2024, intersecting all major geotechnical units relevant to proposed mine development. Hydrogeological data — water table measurements, pore pressure distribution, and groundwater

flow direction — are collected in conjunction with exploration and geotechnical investigations, supplemented once mining commenced by in-pit and underground operational data. Groundwater dewatering and

monitoring wells are the primary data-collection method, supplemented by pore pressure data from vibrating wire piezometers (VWPs) and hydrologic testing (injection/slug tests, air-lift tests, short- and

long-term pumping tests, and spinner logging), analyzed using industry-standard analytical methods. Analytical and numerical groundwater flow models have been developed using hydraulic parameters from testing, 3D geological modeling, and historic

site data. From approximately 1980 through 2008, 14 dewatering wells and 42 monitoring wells were completed; from 2022 through 2025, 23 VWPs were installed in the underground mining area, with additional dewatering wells completed or deepened in

2023, 2024, and 2025 to 2026. There are currently 4 active dewatering wells, 41 active monitoring wells, and 16 active VWPs across 6 locations (Figure 7-25), with pumping rates ranging from 100 to 750 gpm.

Monitoring wells are sampled routinely and analyzed for the State of Nevada Profile I suite at certified laboratories (currently Western

Environmental Testing Laboratory and Pace Analytical, Reno, NV), per permitting requirements; monitor wells and piezometer-equipped exploration drillholes are monitored for water levels and piezometric heads, and surface water is measured and

sampled routinely as required by permits. The Project area has been the subject of multiple hydrogeologic characterization studies by successive owners/operators (WMC 1998, 2000, 2002, and 2005; SWS 2014; Piteau Associates 2018), and more recently i-80 has contracted LRE Water (formerly HydroGeoLogica, now part of Spheros Environmental) for groundwater level and pore pressure monitoring, dewatering well planning and oversight, and groundwater flow modeling

(Figure 7-26).

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Figure 7-25: Well Locations

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Figure 7-26: Timeline for Hydrogeologic Characterization with Relationship to Mining

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The Granite Creek Mine lies in the Great Basin region of the Basin and Range Physiographic

Province, a closed drainage system; Granite Creek is an ephemeral stream sourced from seasonal snowmelt, diverted and routed around the CX Pit and typically infiltrating into permeable alluvial deposits within 762 to 1,500 m (2,500 ft) downgradient,

with mean annual flow estimated at 0.28 cfs. High spring flows have historically been routed to the floor of the mined-out A and B pits, infiltrating into the CX shear zone, which is hydrologically connected

to the CX Pit and underground development. Two groundwater systems are recognized: an alluvial system, moving south-southeast toward the Humboldt River and well understood from historical monitoring, with saturated alluvium concentrated east and

southeast of the Mag Pit; and a bedrock system, controlled predominantly by stratigraphy and geologic structure, with major faults acting as either groundwater flow pathways or gouge-filled barriers, favoring hydraulically isolated fault-bounded

bedrock blocks.

The Mag Pit, excavated mostly in Upper Comus calcareous mudstones, siltstones, and carbonaceous shales, was dewatered from

1987 through April 1998 (initially by sumps, then dewatering wells #12 and #13 from 1991), after which the pit lake was rapidly filled in 2000 and has since generally declined with alluvial groundwater levels, reaching an elevation of 1,411.3 m

(4,630.3 ft) amsl as of January 2026. The CX Pit and underground mine workings lie along east-northeast-trending, steeply southeast-dipping shear zones in the Lower Comus Formation; the CX block has been dewatered from 2007 to present via four

operating dewatering wells (APW-1, BPW-3, BPW-5, and GCW-6) at a combined average rate of

approximately 1,450 gpm, supplemented by approximately 1,350 gpm of underground sump pumping. Water from dewatering wells not used operationally is discharged to four permitted rapid infiltration basins (RIBs), with two additional cells constructed

in Q2 2026; wells with elevated arsenic (APW-1, BPW-5, GCW-6) require treatment prior to discharge at the Water Treatment Plant

(WTP), with a second WTP planned for Q3 2026 to increase capacity to 4,000 gpm.

Spheros Environmental (formerly LRE Water) constructed and

calibrated a numerical groundwater flow model to historic water level measurements and passive inflow rates, used to evaluate the effectiveness of current dewatering wells and estimate passive inflow into the underground workings (UGWs) for the

current Life of Mine plan through 2032 (Figure 7-27). The model simulates current dewatering wells plus two additional wells (GCW-14 and

GCW-15) coming online in Q3 2026 at a conservative 500 gpm each. Passive inflow is predicted at approximately 1,050 gpm in Q1 2026, rising to a maximum of 1,500 gpm in Q1 2029 before leveling as mining

progresses primarily within an already-established cone of depression. The model was also used to estimate groundwater elevation at the Otto, Ogee, and SPZ ore bodies, all of which are predicted to remain below their starting vertical centroid

elevations, with the SPZ and Ogee expected to see more drawdown than Otto due to greater nearby UGW development and proximity to dewatering wells (Figure 7-28). The model results are considered a reasonable

representation of predicted groundwater levels and passive inflow, given good calibration to local historical data; modeling work is ongoing and is being refined as additional data becomes available (Figure

7-29).

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Figure 7-27: Predicted Passive Inflow and Dewatering Well Pumping

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Figure 7-28: Predicted Groundwater Elevation in the Vicinity of Ore Bodies: Otto, Ogee, and South

Pacific

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Figure 7-29: Predictive and Passive Inflows from Scenarios One and Two

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7.4

Geotechnical Data, Testing and Analysis

The geotechnical database developed by i-80 Gold comprises approximately 6,096 m (20,000 ft) of

geotechnically logged core from 40 drillholes distributed throughout the Project area, mostly drilled since 2024 using HQ-diameter core, intersecting all major geotechnical units relevant to proposed mine

development (the Upper Comus, Lower Comus, and SPZ fault-hosted mineralization). Geotechnical Rock Mass Rating (RMR) logging was completed on the drill core, providing the basis for geotechnical domains, rock mass classifications, representative

design parameters, and ground support recommendations for underground mine development, discussed further in Mining Methods (Item 13). The dataset was supplemented by independent WSP photo-logging of three complete and four partial core holes as

part of the data verification program. Rock fabric characterization was completed using optical and acoustic televiewer surveys from 11 drillholes, used to define the orientation and spacing of bedding and major joint sets, develop stereographic

projections of the principal structural domains within the Upper and Lower Comus units, and support kinematic analyses of structurally controlled failure mechanisms and excavation orientation assessments. Dedicated core holes were also drilled along

the proposed Vent Raise 1 and Vent Raise 2 alignments to support geotechnical stability assessments for those excavations.

7.5

Property Plan View

A plan view of the Property showing drillhole locations, coded by operator and significant time period, is presented as Figure 7.2.1-2.

7.6

Exploration Target

Exploration targets consist of five zones at Granite Creek: the CX Fault depth/north extension, Rangefront and Adam Peak, the Ogee Zone, SPZ

infill and depth extension, and the SPZ/ Rangefront /Mag structural intersection.

Rangefront and Ogee are prioritized ahead of active and

planned underground development to reduce the risk of mining ahead of drill data, while the CX Fault program advances the known oxide-to-sulfide transition at depth. If

successful, the program has the potential to add additional Mineral Resources at Rangefront, Adam Peak, Ogee, and the SPZ / Rangefront / Mag intersection, and to upgrade a portion of the existing SPZ Inferred resource to the indicated category. In

addition, mineralization drilled at the CX fault target could result in this area being brought into the mine plan in the near future, allowing flexibility in the mine plan for more active faces and mined tonnage.

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8

Sample Preparation, Analysis and Security

8.1

Sample Preparation Methods and Quality Control Measures

The Granite Creek drillhole database comprises two distinct data sets that differ materially in operator, laboratory, analytical method, sample

security practice, and QA/QC framework: historical (pre-i-80) drilling conducted between 1970 and 2016 under successive operators (PMC,

PMC-Homestake, PMC-Barrick, Atna Resources, and Atna Underground), and i-80 Gold Corp drilling conducted from program inception

in April 2021 through Q1 2026. Sample preparation procedures applied prior to shipment for the historical era were not consistently recorded for the earliest programs (PMC 1970-1996); from 1997 onward, core was logged, and split samples (Jones split

or equivalent) were prepared for shipment to the contract laboratory of record. Under the i-80 Gold program, diamond core drilling predominates, with reverse-circulation drilling used only for pre-collar sections that are logged but not routinely production-sampled. Core is transferred from the core barrel to a wax-impregnated box at the drill rig, washed, and

delivered to i-80 Gold’s Lone Tree core shed for logging and sampling. Logging captures recovery, Rock Quality Designation (RQD), lithology, alteration, oxidation, structure, and sulfide mineralization.

Sample intervals are assigned by a project geologist subject to the rules that a sample must not cross a geological contact or an obvious alteration/oxidation boundary, and a sample length must not exceed 3 m (10 ft); average core sample length from

2023 through Q1 2026 has been approximately 1.3 to 1.5 m (4.3 to 4.8 ft). Core is cut lengthwise with a 5 hp diamond saw using a 14-inch diamond-impregnated blade; soft, clay-rich, or rubbly intervals are

split rather than sawn to preserve sample integrity. The half-core sample is bagged in a heavy-duty Protexo plastic bag with a stapled paper sample tag, tied closed, and held for collection; the remaining half-core is returned to the original core

box and stored under weather-resistant pallet coverings at the Lone Tree facility.

Sample security and chain of custody have been

maintained throughout the i-80 Gold program. Sample bags are collected from the Lone Tree core shed directly by laboratory personnel (ALS or MSA, depending on the program year) and transported by the

laboratory to its preparation facility. Each sample carries a unique numeric tag stapled to the bag, and bags are tied closed for transport. Each hole is tracked under a dedicated dispatch (prefix MSAGCS25_xxxx for surface and MSAGCU25_xxxx for

underground holes processed by MSA Labs) and a corresponding laboratory work-order number. No chain-of-custody breaches have been reported over the period. Drilling

activity predating 2005 predates formal documentation of these procedures; sample security practices for the historical era prior to that point are not separately documented in the source records available to the QP.

8.2

Sample Preparation, Assaying and Analytical Procedures

Historical (pre-i-80) sample preparation and analysis were

performed by a succession of operator-selected laboratories, summarized in Table 8-1: ALS Chemex (Reno, NV) for PMC-Homestake (1997-2000), crushing to -10 mesh with a 300 g Jones split, pulverizing to -150 mesh, and 30 g fire assay with AA finish (detection limit 5 ppb); American Assay Laboratories (AAL, Sparks, NV) for PMC-Barrick (2000-2008), one-assay-ton fire assay with AA finish and multi-element 69-element

aqua-regia ICP-AES analysis; Inspectorate American Laboratories (IAL, Reno, NV; ISO 9002 accredited) for Atna’s surface program (2004-2013), a two-stage crush and

300 g Jones split pulverized to greater than 90% passing -150 mesh with 30 g fire assay and AA finish; and an on-site mine laboratory adjacent to the administration

building for Atna’s underground program (2011 to 2016), following the same general jaw-crush/Jones-split/pulverize/30 g fire-assay-charge sequence. No relationship between these historical laboratories

and the registrant (i-80 Gold Corp, which did not yet hold the property) is applicable.

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Table 8-1: Historical (pre-i-80) Operators, Periods, and Analytical Laboratories at Granite Creek

Operator

Period

Laboratory

Primary Au method

PMC

1970-1996

Not recorded

AA on cyanide leach; FA-AA above 0.01 oz/t

PMC-Homestake

1997-2000

ALS Chemex (Reno)

30 g FA-AA

(Au-AA23); FA-Gravimetric overlimit

PMC-Barrick

2000-2008

American Assay Labs (Sparks)

1 AT FA-AA;

FA-Gravimetric above 5 ppm

Atna (Surface)

2004-2013

Inspectorate American Labs (Reno)

30 g FA-AA; gravimetric above 3 g/t

Atna Underground

2011-2016

On-site mine laboratory

30 g FA on -150 mesh pulp

All sample preparation and analytical work performed under i-80

Gold’s tenure has been completed by independent commercial laboratories holding internationally recognized accreditations (ISO/IEC 17025 for analytical testing, ISO 9001 for quality management) — neither laboratory is affiliated with or

related to the registrant. ALS (Sparks and Elko, Nevada) was the sole laboratory in 2023 and 2024 and remains in use in parallel with MSA Labs from Q1 2026 for confirmation check-assay work. ALS sample preparation follows protocol PREP-31 (PREP-31AY in 2026): air-drying, whole-sample crushing to better than 70% passing 2 mm, riffle/rotary splitting to a 250 g subsample, and pulverizing to better than

85% passing 75 microns. MSA Labs (355 W River Street, Elko, Nevada) became the principal laboratory in 2025, following codes CRU-CPA (crushing to 2 mm and splitting to 500 g),

SPL-410 (riffle splitting to a 250 g subsample), and PPU-510 (pulverizing to better than 85% passing 75 microns). Coarse reject and pulp are retained at the laboratory

for follow-up work.

The principal analytical change over the

i-80 Gold program has been the transition of the primary gold method from conventional fire assay to Chrysos Photon Assay beginning in 2025. Through 2023 and 2024, gold was determined by ALS method Au-AA23, a 30 g fire-assay fusion with atomic-absorption finish (working range 0.005 to 10 ppm Au), with samples above the upper limit re-assayed by gravimetric finish (Au-GRA21, working range 0.05 to 10,000 ppm); a 35-element aqua-regia/ICP-AES suite (ME-ICP41)

was also determined. Beginning in 2025, gold was determined by MSA method CPA-Au-1 (Chrysos Photon Assay, detection limit 0.015 ppm Au), a

non-destructive X-ray activation technique analyzing a substantially larger effective sample mass (typically 400 to 600 g) than conventional 30 g fire assay;

multi-element analysis under the MSA program uses a four-acid ICP-MS suite (IMS-230, 50-plus elements) with mercury by aqua-regia

ICP-MS (IMS-130-Hg). In Q1 2026, the tail of the 2025 underground program was submitted to ALS in parallel with the MSA Photon

Assay program (Au-AA23/Au-GRA21, with a 48-element four-acid ICP-MS suite, ME-MS61m); samples from the new iGU26 hole series are excluded from the PFS database cut-off and will be incorporated in future updates.

8.3

Quality Control Procedures/Quality Assurance

From 2005 to 2015, drilling programs under Atna Resources (surface, 2005-2013), Barrick (2005-2008), and the Atna underground operation

(2011-2016) incorporated systematic QA/QC sample insertion (certified reference materials [CRMs], blanks, and field and preparation duplicates) into the sample stream; no QA/QC data is available for work conducted prior to 2005. QA/QC performance

from these pre-i-80 programs was reviewed in the Granite Creek IA Technical Report (March 2025)

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by the previous QP (GRE), and SRK has independently regenerated the historical control charts from the legacy assay database to verify the previous QP’s conclusions (Figure 8.3-1 through Figure 8.3-3). A total of 1,081 CRMs were inserted between 2005 and 2015 (555 by Atna in 2005, 2006, 2013, and 2015; 526 by Barrick in 2007, 2008, and 2012),

drawn from thirty-seven Rocklabs (New Zealand) and CDN Resource Laboratories (Vancouver, BC) standards at insertion rates averaging 2.7% (peak 5.5%; generally 1-in-20 to

1-in-25 samples); the principal CRMs, certified gold values, and insertion counts are summarized in Table 8-2. A total of 1,249

blank samples were reviewed, of which 270 exceeded the five-times-detection-limit threshold (maximum 1.02 ppm Au, four notable outliers); the previous QP concluded blank contamination was not material to the resource estimate. A total of 287

duplicate samples were reviewed, with field-duplicate scatter returning an R-squared of 0.93 (increased scatter at higher grades attributed to inhomogeneous coarse-gold distribution); duplicate type is not

preserved in the legacy database, so historical pairs are combined under a conservative ±30% field-duplicate envelope (Figure 8-1 through Figure 8-3).

Table 8-2: Historical

(pre-i-80) CRMs at Granite Creek — Certified Au Values and Insertion Counts

CRM

Supplier

Certified Au

(ppm)

± SD

(ppm)

N

(Au inserts)

CDN-GS-7J

CDN Resource Labs

7.34

0.29

194

OxD57

Rocklabs

0.413

0.012

181

CDN-GS-P6E

CDN Resource Labs

0.572

0.031

163

CDN-GS-8C

CDN Resource Labs

8.59

0.52

160

SF23

Rocklabs

0.831

0.027

137

OxH52

Rocklabs

1.291

0.025

130

SG31

Rocklabs

0.996

0.028

116

OxI54

Rocklabs

1.868

0.066

113

CDN-GS-30C

CDN Resource Labs

32.14

0.89

112

OxG60

Rocklabs

1.025

0.028

106

CDN-GS-P1A

CDN Resource Labs

0.143

0.008

105

OxC58

Rocklabs

0.201

0.007

104

SJ32

Rocklabs

2.645

0.068

102

OxF53

Rocklabs

0.810

0.029

95

OxK48

Rocklabs

3.557

0.042

94

OxA59

Rocklabs

0.0817

0.0052

91

OxN49

Rocklabs

7.635

0.189

82

SI25

Rocklabs

1.801

0.044

68

SK33

Rocklabs

4.041

0.103

65

OxP50

Rocklabs

14.890

0.493

37

SN26

Rocklabs

8.543

0.175

22

Other (16 Rocklabs / CDN CRMs)

Rocklabs / CDN

0.024 -14.99 (range)

varies

< 20 each

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Source: SRK, 2026

Figure 8-1: Historical Rocklabs CRM Control Chart (OxD57, Representative)

Source: SRK, 2026

Figure 8-2: Historical Blank Control Chart, 2005 to 2018

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Source: SRK, 2026

Figure 8-3: Historical Combined Duplicate Scatter, 2005 to 2018

i-80 Gold Corp acquired the Granite Creek Property in 2021 and commenced drilling in April 2021. From

program inception through 2025 (the PFS database cut-off), approximately 1,065 drillholes have been sampled and analyzed under a formal QA/QC program designed and implemented by

i-80 Gold Corp. The program is a four-stream design comprising CRMs, coarse blanks, sample (field) duplicates, and preparation duplicates, supplemented by independent third-party check assays —

consistent with current industry best practice. From 2023 onward, the program has targeted an overall control-sample insertion rate of 20% of submitted samples (approximately greater than 5% each of CRMs, blanks, and field duplicates, and greater

than 2.5% preparation duplicates), exceeding the typical 1-in-20 (5%) per-stream insertion rate observed in recent industry

surveys of SK-1300 technical reports. The 2021 and 2022 programs, conducted under a transitional QA/QC framework reviewed by the previous QP, operated at lower aggregate insertion rates of approximately 13% to

21%.

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CRMs are inserted at approximately one per twenty samples, with CRM identity rotated to

maintain coverage across the expected grade range. Acceptance follows the Shewhart control chart convention: assays within ±2 standard deviations (2SD) of the certified inter-laboratory mean are ‘in control’; assays between 2SD

and 3SD are ‘warning’ values triggering batch review; and assays beyond 3SD are treated as batch failures requiring corrective action, using standard deviations published by the CRM supplier (Rocklabs, CDN Resource Laboratories, or

Oreas). The CRM catalog has evolved over the program (Table 8-3): the 2021 and 2022 programs used CDN Resource Laboratories Carlin-style standards (CDN-GS-7J, CDN-GS-8C, CDN-GS-30C, CDN-GS-P1A, CDN-GS-P6E); 2024 introduced Oreas Carlin-style standards (OREAS-277, OREAS-279, OREAS-282); and 2025 onward added OREAS-238B, OREAS-241, and OREAS-241B.

During QP review, a subset of records carrying a mismatched CRM label were identified as wrong-standard inserts or data-entry transcription errors (rather than genuine laboratory failures) and were excluded from the CRM performance assessment; the

issue has been raised with i-80 Gold for database hygiene remediation. SRK has regenerated CRM control charts for each of the principal standards used over the period (Figure

8-4 through Figure 8-9).

Table 8-3: i-80 Gold CRM Catalog Used at Granite Creek, 2021 to 2026 (year-to-date) — Certified

Au Values and Insertion Counts

CRM

Supplier

Certified Au (ppm)

± SD (ppm)

n (Au inserts)

Period used

CDN-GS-30C

CDN Resource Labs

32.14

0.445

500

2021-2025

CDN-GS-P6E

CDN Resource Labs

0.572

0.0155

402

2022-2024

CDN-GS-P2B

CDN Resource Labs

0.433

0.011

385

2022-2025

CDN-GS-7J

CDN Resource Labs

7.34

0.145

354

2022-2026

OREAS-277

Oreas

3.39

0.18

258

2024-2026

OREAS-282

Oreas

14.05

0.3855

248

2024-2026

OREAS-279

Oreas

6.55

0.327

196

2024-2026

OREAS-238B

Oreas

3.16

0.10

140

2025-2026

OREAS-241

Oreas

7.06

0.15

127

2025-2026

CDN-GS-P1A

CDN Resource Labs

0.143

0.012

84

2021-2022

CDN-GS-8C

CDN Resource Labs

8.62

0.72

83

2021-2022

OREAS-241B

Oreas

7.058

0.163

45

2025-2026

CDN-GS-20C

CDN Resource Labs

19.67

0.381

13

2025

OREAS-273

Oreas

10.76

0.50

9

2025-2026

OREAS-264

Oreas

0.307

0.0185

5

2025

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Source: SRK, 2026

Figure 8-4:

CDN-GS-7J Control Chart, i-80 Era

Source: SRK, 2026

Figure 8-5:

CDN-GS-30C Control Chart, i-80 Era

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Source: SRK, 2026

Figure 8-6:

CDN-GS-P6E Control Chart, i-80 Era

Source: SRK, 2026

Figure 8-7: OREAS-277 Control Chart, i-80 Era

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Source: SRK, 2026

Figure 8-8: OREAS-279 Control Chart, i-80 Era

Source: SRK, 2026

Figure 8-9: OREAS-282 Control Chart, i-80 Era

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Coarse blanks (Vigoro brand crushed marble chips) are inserted to monitor the full

sample-preparation circuit for contamination or carry-over, placed where possible directly behind the highest-grade samples in each batch. The acceptance threshold is set at five times the analytical lower limit of detection (LLD; 0.025 ppm Au for

the conventional fire-assay methods used through 2024, held at the same conservative level for Photon Assay despite its slightly higher 0.015 ppm Au detection limit); LLDs and thresholds by method are summarized in Table 8-4. The i-80 era blank control chart regenerated by SRK (Figure 8-10) shows 89.8% of blank inserts (2,112 of 2,353) below the

5×LLD design threshold; 241 exceed it. Applying a materiality floor at 10×LLD (0.05 ppm Au) reclassifies 74 of the 241 exceedances as within acceptable range, yielding a revised pass rate of 92.9% (2,186 of 2,353). Of the remaining 167

failures, 128 (76%) were assayed at the Lone Tree in-house laboratory, spread across 2022 through 2025 rather than confined to a single period — discussed further under Opinion on Adequacy as a

laboratory-performance matter.

Table 8-4: Lower Limit of Detection (LLD) and 5xLLD Blank

Acceptance Threshold by Gold Analytical Method

Method

Lab

LLD (ppm Au)

5×LLD threshold

(ppm)

Au-AA23

ALS

0.005

0.025

Au-GRA21

ALS

0.05

0.25 (held at 0.025 conservatively)

ME-ICP41

ALS

0.001

0.005

ME-MS61m

ALS

0.005

0.025

CPA-Au-1

MSA

0.015

0.075 (held at 0.025 conservatively)

IMS-230

MSA

(multi-element)

n/a

Source: SRK, 2026

Figure 8-10: i-80 Era Blank Control Chart, 2021 to Q1

2026

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Two duplicate streams are run: sample (field) duplicates, prepared at the Lone Tree core shed

by quarter-splitting selected core intervals (one quarter primary, one quarter duplicate — chosen over half-core duplication to preserve a portion of the original core for future reference), monitoring in situ sampling variance plus all

downstream preparation and assay variance; and preparation duplicates, prepared at the laboratory from the coarse-crushed reject of selected samples, monitoring laboratory preparation and analytical variance alone. The QP applied the duplicate-pair

regression framework of Thompson and Howarth (1978), with tolerance envelopes reflecting the additivity of in situ sampling, preparation, and analytical variance across stages: ±10% mean relative difference (MRD) for lab (pulp/analytical)

duplicates, ±20% MRD for preparation duplicates, and ±30% MRD for field (quarter-core) duplicates, reflecting the nuggety distribution of gold in Carlin-style mineralization (Table 8-5).

Preparation duplicates are additionally expected to return an R-squared of 0.99 or better across the analytical method’s mineralized working range. SRK has regenerated the duplicate scatter plots for the

i-80 era (Figure 8-11 through Figure 8-13).

Table 8-5: Duplicate Tolerance Envelopes Applied to the

i-80 Gold Drillhole Assay Database

Duplicate type

Variance sources

Envelope (MRD)

Basis

Lab (pulp)

Analytical only

±10%

QP-adopted basis: 9% RP observed for pulp >2 g/t Au

Preparation

Prep + analytical

±20%

QP judgment + typical CVAVE for Carlin-style prep

Field

(quarter-core)

In-situ + prep + analytical

±30%

QP judgment + typical CVAVE for nuggety Carlin-style gold

Historical (combined)

Type not preserved in

legacy DB

±30%

Conservative; assumes field-equivalent

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Source: SRK, 2026

Figure 8-11: Field (Quarter-core) Duplicate Scatter, i-80

Era

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Figure 8-12: Preparation (Coarse Reject) Duplicate Scatter,

i-80 Era

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Figure 8-13: Lab (Pulp / Analytical) Duplicate Scatter, i-80

Era

Annual QA/QC insertion counts for the 2021 through Q1 2026 i-80 Gold drilling programs are

summarized in Table 8-6. The 2021 and 2022 reviews are drawn from the previous QP’s analysis in the Granite Creek IA Technical Report (March 2025); the 2023 through Q1 2026 reviews are based on the eight

quarterly QA/QC reports issued by i-80 Gold Corp.

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Table 8-6: Annual QA/QC Insertion Summary, i-80 Gold Drillhole Sampling at Granite Creek, 2021 to 2026

Year

Au-method

CRM

Au-method

Blank

Field dup

Prep dup

Lab dup

Check assay

Total

2021

109

122

89

94

1,568

0

1,982

2022

825

844

688

356

683

0

3,396

2023

260

169

116

113

78

69

805

2024

332

237

152

95

129

54

999

2025

970

741

386

388

1,081

411

3,977

2026 (YTD)

353

240

99

139

330

9

1,170

Total

2,849

2,353

1,530

1,185

3,869

543

12,329

2021

Approximately 69 holes drilled (20 surface diamond core, 9 metallurgical, 40 underground diamond core). The previous QP reviewed 1,395

drillhole samples with control samples inserted at 3.9% CRMs, 3.7% blanks, 2.4% field duplicates, and 2.7% preparation duplicates using five Carlin-style CDN standards; performance was acceptable, with no systematic CRM bias, blanks predominantly

below the detection-limit threshold, and preparation-duplicate R-squared of 0.99.

2022

Approximately

154 holes drilled (30 surface diamond core, 56 underground diamond core, 62 underground production core, 5 underground Cubex, 1 water well). The previous QP reviewed a three-hole, 800-sample sub-set with control samples inserted at 5.9% CRMs, 6.4% blanks, 6.0% field duplicates, and 3.0% preparation duplicates; performance was again acceptable. Beginning with 2023, the QA/QC framework was formalized to

the current 20% target insertion rate.

2023

Approximately 278 holes drilled, including a twelve-hole deep SPZ infill program (~5,904 m / 19,369 ft). Control samples were inserted at the

20% program target using CDN-GS-30C, CDN-GS-7J, and CDN-GS-P6E; SRK’s regenerated control charts show no consistent CRM bias, and all duplicate variance fell within acceptance. 69 check-assay samples (8 holes) were submitted to American Assay

Laboratories (AAL).

2024

Approximately 185 holes drilled, continuing the deep SPZ infill program. The principal QA/QC change was migration from CDN to Oreas standards (OREAS-277/279/282). One CRM exceedance was investigated and resolved by re-assay; duplicate variance was within acceptance for most pairs. 54 check-assay samples (3 holes)

were submitted to AAL.

2025

Approximately 379 holes drilled, marking a substantial expansion with the introduction of underground production drilling. SRK’s

regenerated control charts show weighted-average observed-to-certified ratios of 84.6% (Q3) and 98.1% (Q4); OREAS-238B/241/241B remained within ±1.5% of certified

in both quarters, while OREAS-282 averaged 82.0% of certified in Q3 before recovering to 96.0% in Q4 — carried forward as a corrective-action item under Opinion on Adequacy, below. 93.7% of 741 blanks

returned values below the 5×LLD threshold. Q4 2025 included the program’s first material check-assay round: 209 samples (10.5% QC) submitted to a third-party laboratory, with 93.0% of primaries within ±15% of the check result.

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Q1 2026

Drilling continued as the tail of the 2025 SPZ infill campaign and the start of 2026 exploration drilling targeting the Ogee, Otto, Adam Peak,

and Range Front fault systems; the PFS database cut-off restricts the dataset to holes iGU25-37, iGU25-37A, and iGU25-38. Two laboratories were used in parallel (MSA for iGS25/iGU25-prefix holes; ALS for the new iGU26 prefix), with overall control-sample insertion of 19.5% (214 of 1,097 samples). ALS blank failures were

attributed to minor pulverizer carry-over, with a procedural correction implemented; four OREAS-282 failures on the ALS gravimetric method were due to undersized pulp aliquots, and CRM submitted mass has been

increased going forward. A cross-laboratory check assay (24 MSA Photon Assay pulps re-analyzed by ALS fire assay) returned R = 0.9972 with a small negative bias of

-3.48% (ALS lower than MSA).

i-80’s Check Sample

Report identifies check-assay activity in every year from 2023 onward — 69 samples (8 holes) in 2023, 54 (3 holes) in 2024, 411 (37 holes) in 2025, and 9 (2 holes) in Q1 2026, totaling 543 samples across 49 holes. Lab destination was

determined by cross-referencing each sample’s check-side assay-method code against the corporate QC database: the 2023-2024 population (123 samples) is American Assay Laboratories, and the 2025-Q1 2026

population (420 samples) is ALS — both independent third-party laboratories. Across the full 543-sample population, SRK’s check-assay precision analysis (Figure

8-14) found a mean bias of -6.5% and a median of -3.5% (check assay relative to primary), with 75.3% of pairs within ±10%.

Documented corrective actions include: re-assay of the bracketed batch following the Q3 2024 OREAS-279 failure (all bracketed primaries confirmed within tolerance on re-run); flagging of a Q1 2024 sample-duplicate failure for follow-up coarse-reject re-assay; instruction to submit twice the prior CRM

material mass for check-assay submissions following the Q4 2025 undersized-aliquot findings; and a procedural correction at the ALS pulverizer sequence following the Q1 2026 blank carry-over findings.

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Figure 8-14: Check Sample Report Primary vs. Check Assay Scatter, Au, 2023 to 2026

8.4

Opinion on Adequacy

Based on review of the eight quarterly QA/QC reports issued by i-80 Gold Corp over the period Q3 2023

to Q1 2026, the previous QP’s review of the 2021 and 2022 programs in the Granite Creek IA Technical Report (March 2025), the prior QPs’ reviews of the historical

(pre-i-80) data, and the Q1 2026 Photon Assay method-validation study, the QP concludes the following. Historical (pre-i-80) sample preparation, analytical, and QA/QC procedures from 2005 to 2016 are documented in prior technical reports and have been accepted as adequate for mineral resource estimation by all prior QPs,

including in the most recent (March 2025) IA Technical Report; no QA/QC data is available for the pre-2005 portion of the historical dataset, but the volume of subsequent drilling under the modern QA/QC

framework is sufficient to support the mineral resource estimate without exclusive reliance on the unverified historical data. i-80 Gold sample preparation was performed by independent ISO/IEC 17025-accredited

commercial laboratories (ALS Minerals and MSA Labs) using industry-standard preparation protocols appropriate for Carlin-style gold mineralization, consistently delivering the target product (greater than 85% passing 75 microns) with high

preparation precision (preparation-duplicate R-squared of 0.99 across the program; 0.9997 in Q1 2026).

Analytical methods used by i-80 Gold over the period are appropriate for the deposit style and grade

distribution. The 2021 to 2024 fire-assay approach (Au-AA23 with Au-GRA21 gravimetric overlimit) is the industry-standard method for sediment-hosted gold deposits. The

transition to Chrysos Photon Assay from 2025 onward has been independently validated against both conventional Fire Assay and Screen Fire Assay (R-squared of 0.994 or greater; mean bias +3% to +6%,

mechanistically explained by larger effective sample mass), and is fit-for-purpose as the primary analytical method for resource

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estimation at Granite Creek. The i-80 Gold QA/QC program has consistently achieved or exceeded the design control-sample insertion rate of 20% across the

2023 to Q1 2026 period; the 2021 and 2022 programs operated at lower insertion rates within a transitional framework reviewed and accepted by the previous QP. CRM, blank, and duplicate performance has been within industry-accepted acceptance limits

in every quarter, with all identified failures appropriately investigated and corrected, or appropriately accepted; the Q3 2025 OREAS-282 low bias and the concentration of blank exceedances in Lone Tree in-house-assayed batches are the two items the QP has carried forward as ongoing laboratory-performance matters for continued monitoring. Sample security and chain of custody for the

i-80 Gold program are appropriate: samples were collected at i-80 Gold Corp’s secure Lone Tree core shed and transported in sealed sample bags directly by

laboratory personnel to the preparation facility, with no chain-of-custody breaches reported over the period.

In the QP’s opinion, the sample preparation, analytical, and security procedures used to generate the combined Granite Creek drillhole

assay database — historical (1970 to 2016) and i-80 Gold (2021 to Q1 2026) — are adequate to support mineral-resource estimation, and mineral-reserve and economic studies on the property. SRK

recommends that every QC sample record (CRM, blank, field duplicate, preparation duplicate, pulp duplicate, and check assay) be flagged in acQuire with its specific QC type and identity at the time of data entry or LIMS import, rather than relying

on generic, ambiguous, or system-generated placeholder identifiers; where a standard or blank record cannot be immediately matched to a certified value or known material at the time of entry, it should be flagged explicitly as ‘unresolved,

pending confirmation’ rather than left unflagged or carried under a temporary label. SRK further recommends that i-80 and its laboratories capture and retain

per-sample analysis order (or equivalent sample-sequence data) alongside QC results, so that future blank exceedances can be evaluated by a carry-over-percentage method rather than the coarser absolute-value

materiality floor applied in this report.

8.5

Non-Conventional Industry Practice

In Q1 2026, i-80 Gold Corp commissioned a three-way analytical

comparison of MSA Photon Assay (CPA-Au-1), conventional Fire Assay (AAL method IO-FAAu30), and Screen Fire Assay (AAL method G-FASFAu11) to assess whether Photon Assay — a non-destructive, higher-sample-mass technique that is comparatively new relative to conventional fire assay —

produces results statistically comparable to the established fire-assay methods, and to characterize any systematic bias. Photon Assay consistently returned slightly higher gold values than both fire-assay variants (a small positive bias of

approximately +0.55 ppm Au vs. Screen Fire Assay and +0.69 ppm Au vs. conventional Fire Assay), an effect that is expected and widely documented in the literature, arising from the non-destructive nature of

Photon Assay and its substantially larger effective sample mass (typically 400 to 600 g, versus a 30 g fire-assay charge), which better represents coarse-gold and nugget-effect populations. The coefficient of determination exceeded 0.994 against

both fire-assay reference methods across the analyzed grade range, and the Bland-Altman limits of agreement are narrow (approximately ±1 to 2 ppm Au) with no grade-dependent behavior. In the QP’s opinion, this validation supports the

reliability of Photon Assay as the primary gold analytical method for the Granite Creek resource estimation, with Screen Fire Assay recommended as a supplementary check method on samples flagged for coarse-gold or nugget-effect risk.

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9

Data Verification

Data supporting this technical report summary has been verified through two successive programs: a comprehensive verification program completed

by GRE in 2021 in support of the Granite Creek IA Technical Report (March 2025), and a further, independent verification program completed by SRK in 2026 in support of this Pre-Feasibility Study, comprising a site inspection, review of i-80’s Standard Operating Procedures, statistical review of the drilling database, and a systematic comparison of the assay database against original laboratory certificates. Both programs are described in

Section 9.1.

9.1

Data Verification Procedures

GRE completed a comprehensive, SK-1300-compliant data verification program to support the previously

disclosed IA of the Project (Initial Assessment SK-1300 Technical Report, Granite Creek Mine Project, Humboldt County, Nevada, USA, issued March 26, 2025). The GRE QPs carried out an on-site inspection on April 20, 2021, including review of key geological formations, structural controls, and mineralization, confirming that field observations were consistent with existing geological mapping

(Osgood, 2016); core and RC sample storage, handling, and condition were also inspected and found appropriate for that level of study. Approximately 752 RC and core sample intervals from four drillholes were visually checked against drill logs and

database records, confirming that logged lithologies and sample descriptions accurately matched the physical samples. To verify assay reliability, seven core/chip check samples and two surface rock chip samples were collected and submitted to Hazen

Research Inc. for fire assay; the check assays showed excellent correlation with the original results (R-squared approximately 0.99), with a t-Test indicating no

statistically significant bias between the original and check datasets. GRE’s Geology QP manually audited roughly 10% of original assay certificates from the 2021 to 2022 drilling campaigns against the digital database and identified no

material errors, and recommended that i-80 implement routine automated database audits after significant programs or database updates. Discipline-specific QPs concluded the data were adequate for IA purposes:

the Geology QP considered lithology, mineralization, and assay data reasonably accurate and suitable for mineral resource estimation; the Metallurgy QP found the metallurgical test work representative, professionally executed, and appropriate for

recovery assumptions; the Mine Planning QP verified key operating and cost assumptions against comparable operations and industry benchmarks; and the Environmental QP confirmed current compliance with Nevada permits and inferred that environmental

data met state QA/QC requirements for that study level. Overall, GRE’s QPs reported no limitations or failures in data verification and considered the datasets used in the Granite Creek IA reliable and appropriate for IA-level evaluation.

During February 2026, SRK conducted a site visit and tour of the Granite Creek

operation, including underground mine headings and the Lone Tree facility (core processing and sample storage). Mr. Burkett reviewed and discussed with site personnel the site geology, structure, mineralization, drilling, core logging and

sampling, density measurements, site security, core and sample storage, and data management procedures; the visit was hosted by a combination of i-80 management and site-based staff, with unrestricted access

to all aspects of the Project and all questions satisfactorily addressed by the Company. WSP coordinated and reviewed the 3D structural interpretation and geologic model constructed by i-80, and separately

undertook an exercise to calibrate the ventilation model to data collected during the site visit; the data used in the ventilation evaluation is considered adequate for the purpose used in this technical report.

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The QP completed a phased approach to verification of the data provided by i-80, comprising a site inspection, review of i-80 Standard Operating Procedures (SOPs), statistical review of the databases, review of mineralized drill core intervals, an

underground mine tour focused on active headings in mineralized material, and verification of assay certificates against original laboratory-issued documents; the QP also independently reviewed i-80’s

internal quality control program for analytical data and drilling survey confidence. It is the QP’s opinion that the current procedures and protocols in place for data collection and validation are acceptable for use in mineral resource

estimation. WSP separately spent two days on site comparing core shed observations of all major lithologic units with conditions underground, the geotechnical logging database, and the geologic model; in-place

Cemented Rock-fill (CRF) backfill was observed underground, backfill QA/QC procedures were reviewed, and the CRF backfill QA/QC database was reviewed for completeness and consistency of backfill strength verification over time. The data used in the

geotechnical evaluation is considered adequate for the purpose used in this technical report.

As part of the data verification program,

the QP conducted an underground mine tour of the Granite Creek operation focused on direct observation of mineralization in active working headings, to visually confirm the geological controls on mineralization, the geometry of the mineralized

structures, and the consistency of the lithological and structural relationships captured in the i-80 drillhole database and 3D geological model. The QP visited multiple working headings in mineralized

material and directly observed the hanging wall and footwall contacts of the mineralized structures, the character and continuity of vein and breccia textures, the distribution of sulfide and oxide mineralogy, and the structural attitudes of the

controlling faults (Figure 9-1 and Figure 9-2). These observations were compared against the i-80 structural and lithological

model and used to independently validate the geologic model and the mineralization domaining strategy applied during resource estimation; the character of the mineralization observed in situ is consistent with the logged descriptions, domain

boundaries, and assay distributions stored in the database for the corresponding drillhole intervals.

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Figure 9-1: Underground Working Heading at Granite Creek Showing Mineralized Structure with Sample

Outline and Drill Pattern Marks Observed During the QP Site Visit

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Figure 9-2: Underground Working Heading at Granite Creek Showing Exposed Mineralized Face Observed

During the QP Site Visit

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Following the underground tour, the QP spent time at the

on-site core logging facility reviewing retained drill core from recent infill and definition drilling, selecting core boxes spanning a range of grade intervals and mineralization styles represented in the

resource model. For each interval reviewed, the QP placed printed assay sheets for the corresponding drillhole alongside the core box and cross-checked the visual character of the mineralization against the Au and Ag grades recorded in the database,

interval by interval (Figure 9-3 and Figure 9-4). The visual character of the mineralized intervals (degree of structural damage) was consistent with the recorded Au and

Ag grades — elevated-grade intervals corresponded to visibly mineralized and damaged core, and low-grade intervals corresponded to weakly mineralized or barren core with little structural deformation;

lithology was not part of this review. No discrepancies were identified between the observed mineralization and the corresponding database Au and Ag grades. Together, the mine tour and the core review provided the QP with an independent, field-based

check on the integrity of the mineralization-domain and assay data used in mineral resource estimation.

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Figure 9-3: Drill Core from Hole iGS22-07 (1,499 to 1,482 m

(581 ft)) with Printed Assay Sheet Used by the QP to Cross-Check Mineralization Against Database-Recorded Au and Ag Grades

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Figure 9-4: Drill Core from Hole iGS23-05 (1,498 to 1,491 m

(610 ft)) with Printed Assay Sheet Used by QP to Cross-Check Mineralization Against Database Au and Ag Grades

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The QP conducted a systematic comparison of the Au and Ag values stored in the i-80 acQuire drilling database against the original analytical values reported in laboratory certificates issued by ALS Minerals, MSALABS, and American Assay Laboratories (AAL), to confirm that the database

accurately reflects the laboratory-reported results with no transcription or data-management errors. The comparison was conducted programmatically using CSV-format laboratory certificate exports: 138

certificate files were processed, encompassing 14,127 sample rows across the surface and underground drilling programs. Database records were matched to certificate entries by sample ID — of the certificate samples, 10,139 were present in the

assay database and 3,988 were lab-inserted QC materials (blanks, CRMs, laboratory duplicates) not retained in the project assay table. A 5% discrepancy threshold was applied to flag samples with differing Au

or Ag values.

The initial comparison flagged 13 samples exceeding the 5% threshold, all involving the Au value; these were individually

reviewed, and the laboratory was contacted for clarification. MSALABS provided updated certificate documentation for the affected Photon Assay jobs, consisting of dedicated Au-only Photon Assay re-run reports issued separately from the primary multi-element certificates for samples flagged as Heterogeneous Block (‘HB’) during the initial Photon Assay screen. Once incorporated, the 13 flagged

samples resolved as follows: six samples (MSALABS job ELK2510227) were Photon Assay re-run reconciliations — the database had correctly stored the Au value from the authoritative Au-only re-run certificate, and the initial comparison had referenced only the multi-element certificate without accounting for the sibling

re-run file, so no database correction was required; the remaining seven samples (across MSALABS lab jobs ELK2510221, ELK2510270, and ELK2610263) were database discrepancies — drillhole sample rows

existing in the database with no assay value populated, despite valid laboratory certificates (six at or below the 0.015 ppm Au detection limit and one at 0.018 ppm Au), representing database ingest gaps rather than transcription errors, now

documented and provided to i-80 for re-ingest; given the very low grades and limited count, the impact on the mineral resource estimate is considered negligible.

Following resolution of all 13 samples, the comparison was re-run against the updated database and revised certificates, confirming a 99.90% match (7,181 of 7,188 compared samples) between the database Au

values and the original laboratory-reported values; the seven residual samples (non-mineralized) were not received prior to the PFS database cut-off date (February 2,

2026).

The QP performed statistical analyses on the drilling database to check for potential erroneous data in the collar, downhole

survey, geology/lithology, specific gravity, and assay tables, including descriptive statistics, charting, and review of potential outlier and erroneous data, to identify errors common among drilling databases (zero values, treatment of

below-detection-limit values, negative or non-numeric values, extreme outliers, and interpretation of the distribution of mineralization across the Property). As part of the updated Mineral Resource reporting,

SRK completed a detailed review of the database provided; through discussion with the i-80 geological team, several issues were noted and addressed prior to modeling. Seven holes

(GCPR25-050, GCPR25-309, GCPR24-108, GCPRPR25-054,

GCPR25-045, GCPR25-349, GCPR25-330) showed potential downhole contamination with high-grade smearing down the hole; in all but GCPR25-045, selected intervals were reset to trace to avoid local overstatement of gold. Two holes (GCPU22-34 and GCPU22-55) contained low-grade assays within areas of expected grade but had no collar records in hard copy, and were potentially misplaced in the model — these were removed from the estimation dataset. A further hole with a low-grade interval (iGU22-44) was reviewed but kept in place, as records confirmed its location.

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i-80 has a well-documented series of SOPs covering

core handling, geological and geotechnical logging, sample collection, chain of custody, QA/QC submission protocols, sample preparation and analysis guidelines, and sample retention policy, applicable to all

i-80 employees and contractors at any i-80 site including Granite Creek; SOPs are updated on an as-required basis. Data

verification occurs prior to incorporation of data into final databases. i-80’s SOPs are listed in Table 9-1, with the

sub-sections below summarizing the SOPs reviewed by the QP. The QP reviewed the SOPs and found them to meet SME standards; the SOPs are specifically designed to comply with U.S. SEC S-K 1300 disclosure requirements for exploration projects. During the site visit, core was not being processed, so the QP did not observe site personnel acting in accordance with the SOPs.

Table 9-1: i-80 Standard Operating Procedure List

SOP Protocol

Version

SOP Updated Date

Core Logging Procedure

Rev. 3

October 2024

Exploration Drill Sampling and Chain of Custody

v1.2

December 2021

Sampling and Assay Quality Control Guidelines

v1.2

December 2021

Sample Analysis Guidelines

v1.2

March 2022

Sample Retention Guidelines

v1.3

April 2022

Core drilling is carried out by a reputable contractor using a diamond bit and core barrel to obtain

runs of up to 3 m (10 ft) (Figure 9-5); the standard diameter for exploration drilling at Granite Creek is HQ (21⁄2-inch

diameter core). Core boxes are labeled with the hole ID and footage interval and are collected from the drill site daily by i-80 geologists or geotechnicians and transported to the logging facility (Figure 9-6). Sample intervals are selected by the project geologist based on geological interpretation, lithology, alteration, and structural contacts; sample lengths are nominally 1 to 2 m (5 ft) in ore-zone intervals and up to 3 m (10 ft) in waste. Regular samples consist of one half of the core cut lengthwise with a diamond saw, with the remaining half retained in labeled boxes for the life of the project.

Sample identifications follow a 13-character alphanumeric format (the Granite Creek underground program prefix is GCEXU, e.g., GCEXU25002980); scannable sample tags are affixed to sample bags, with the sample

ID also transcribed by permanent marker as a backup, and each bag sealed with a tie. All samples remain under the control of i-80 management from collection through delivery to the laboratory, stored while

awaiting shipment in a secure facility not accessible to unauthorized personnel (Figure 9-7), and transported by i-80 staff or a laboratory-approved carrier.

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Source: SRK, 2025

Figure 9-5: Diamond Core Saw Station at the Lone Tree Laboratory Facility

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Source: SRK, 2025

Figure 9-6: Palleted and Tarped Coarse Reject and Core Storage Yard at the Lone Tree Facility

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Figure 9-7: Pulp Sample Storage at the Lone Tree Facility

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Prior to logging, core is washed and scrubbed to remove drill fluids and mud, so geological

textures can be clearly observed, with core boxes oriented in ascending order and box labels verified for accuracy before logging commences. All geological and geotechnical data are recorded directly in acQuire database software via remote desktop

connection to the i-80 server; drillhole metadata (hole ID, start/end dates, actual depth, drilling company, hole type) are entered upon hole completion using the project Drillhole Tracker spreadsheet as the

source document. Rock Quality Designation (RQD) is measured for each 2 m (5 ft) interval by summing the lengths of competent core pieces above a minimum threshold (NQ: 0.34 ft; HQ: 0.40 ft; PQ: 0.55 ft), recording total recovery and RQD footage;

geological logging captures lithology, alteration, oxidation, sulfide mineralogy, veining, and structures at a level of detail appropriate for resource modeling, with non-overlapping, consecutive logging

intervals. Bulk density (specific gravity) is measured using either the wax immersion method or the vacuum sealer method (Figure 9-8), with one density sample selected per stratigraphic unit and increased

sampling density in the ore zone and altered intervals; samples are dried for a minimum of two hours before measurement, and density data are recorded directly in acQuire. Core is photographed in labeled boxes after logging and before cutting, taken

wet to improve color and texture contrast, and named by hole ID and footage interval for upload to the i-80 SharePoint.

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Source: SRK, 2025

Figure 9-8: Specific Gravity (Water Immersion) Measurement Station at the Lone Tree Laboratory

Facility

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Sample preparation is completed at ALS located in Sparks, Nevada; the prepared pulp is a

minimum of 30 g, ground to 85% passing 200 mesh, with prepared pulps returned from the commercial laboratory after analysis and stored in a dry, secure location. The primary analytical program for all samples includes: gold fire assay (AuFA) with

atomic absorption (AA) finish for all samples (samples returning greater than 10 ppm Au at commercial laboratories are re-analyzed with a gravimetric finish); gold cyanide (AuCN) assay with AA finish for all

underground exploration and production samples where AuFA exceeds 2 ppm, and for open-pit samples where AuFA exceeds 0.2 ppm; carbon and sulfur analyses by LECO method (total carbon, carbonate, sulfate sulfur,

sulfide sulfur, and total sulfur) on selected exploration and production samples; and multi-element ICP analyses on exploration samples where AuFA exceeds 2 ppm, on all surface exploration samples regardless of gold grade, and on every fourth

underground production drillhole, covering a 40-element suite (Ag, Al, As, Ba, Bi, Ca, Cd, Ce, Co, Cr, Cu — total and cyanide-soluble, Fe, Hg, K, Li, Mg, Mn, Mo, Na, Nb, Nd, Ni, P, Pb, Rb, Sb, Sc, Se,

Sn, Sr, Te, Th, Tl, U, V, W, Zn, and Zr).

i-80’s QA/QC program is designed to meet the

minimum standards of the SME and to comply with U.S. SEC S-K 1300 disclosure requirements, documented in the i-80 Sampling and Assay Quality Control Guidelines (v1.2,

December 2021). The overall QA/QC insertion rate is a minimum of 17.5% of total submitted samples, comprising: CRMs — commercial standards supplied by CDN Resource Laboratories (or equivalent), inserted at a minimum rate of 5% of submitted

samples with at least two standards of different grades per submission, required to fall within ±3 standard deviations of the certified expected value (failure triggers re-assay of all samples between

adjacent passing standards); blanks — coarse marble-chip blank material inserted at a minimum rate of 5%, acceptable when less than five times the laboratory’s gold detection limit, preferentially placed after high-grade standards to

detect contamination (failure triggers re-assay of all samples between adjacent passing blanks); and duplicates — field duplicates (second half of split core) at 5% of submitted samples targeting

competent rock intervals (±20% relative-difference acceptance, pending accumulation of project-specific variance data), preparation duplicates (coarse reject split at the laboratory) at 2.5% (±15% relative difference), and analytical

duplicates (pulp re-split at the laboratory) assessed as required (±10% relative difference). External check analyses are conducted by submitting 5% to 10% of pulps to an independent secondary

laboratory using the same digestion and analytical finish as the primary laboratory, acceptable within ±10% accuracy and ±10% precision; serious discrepancies (greater than 15% for fire assay, greater than 20% for cyanide methods)

trigger submission to a third ISO-certified laboratory. All QA/QC data are reviewed on receipt of each assay certificate, and no results are released until QA/QC controls are validated; results are compiled

and reported quarterly to the i-80 Senior Resource Geologist, including performance statistics for standards, blanks, and duplicates, Thompson-Howarth precision plots, and documentation of any failures,

corrective actions, and re-assays.

i-80’s Sample

Retention Guidelines (v1.3, April 2022) govern the storage and disposal of all sample materials from the Granite Creek Project: one half of the split diamond drill core is retained in labeled boxes for the life of the project at the Lone Tree core

shed; coarse reject material generated during laboratory sample preparation is retained for a minimum of three years, stored in dry conditions (palleted and tarped, or in labeled barrels) to allow retrieval by batch; sample pulps are stored for the

life of the project at the Lone Tree Laboratory or commercial laboratory facility; and disposal of coarse reject, core, or sample pulps requires written approval from the VP Geology and must follow i-80

environmental disposal protocols.

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9.2

WSP Verification

WSP spent two days on site comparing core shed observations of all major lithologic units with conditions underground, the geotechnical logging

database, and the geologic model. Selected core intervals were independently reviewed against available core photographs and geotechnical logging records as part of the data verification process. Observations of

in-placed cemented rockfill (CRF) backfill were made underground, backfill QA/QC procedures were reviewed, and the CRF backfill QA/QC database was reviewed for completeness and consistency of backfill strength

verification over time. Data used in the geotechnical evaluation is considered adequate for the purpose used in this technical report.

WSP

spent one day on site and completed a ventilation assessment including inspection of primary, booster, and auxiliary ventilation systems, as well as a review of active and planned mine workings, haulage routes, escapeways, and proposed ventilation

raise locations. Temperature and humidity data were collected for Ventsim™ model calibration. The data used in the ventilation evaluation is considered adequate for the purpose used in this

technical report.

9.3

Limitations

The following items are identified as known limitations on data verification and are accounted for in mineral resource classification by the

QP: the QP did not directly supervise or oversee data acquisition of any drilling, logging, or analytical data used in the determination of mineral resources — the QP reviewed summary data, technical reporting, and supporting documentation,

and performed independent verification, with all descriptions of procedures and methods provided by i-80; and the QP has not conducted a visit or inspection of the analytical laboratories providing the

baseline analytical data supporting resources, though the laboratories used are considered reputable and independent facilities suitable for the analyses performed.

9.4

Opinion on Data Adequacy

It is the QP’s opinion that the data provided by i-80 for the Granite Creek Project are

reasonable, reliable, and of acceptable quality and quantity to support the mineral resource estimate and the technical disclosures presented in this Technical Report Summary. The drilling, sampling, sample preparation, analytical, data management,

and QA/QC procedures applied at Granite Creek are consistent with SME Best Practice Guidelines and satisfy the disclosure requirements of U.S. SEC S-K 1300. This opinion is based on the data verification

activities completed by the QP, summarized as follows: the assay database vs. analytical certificate comparison confirmed a 99.90% match between the i-80 database Au values and the original laboratory

certificates across the 7,188 samples available for direct comparison (the seven residual non-mineralized samples were not received prior to the PFS database cut-off

date); the underground mine tour, through direct observation of mineralization in active working headings, confirmed the geometry, hanging wall and footwall contacts, and structural controls represented in the

i-80 geological model and mineralization domaining; the core inspection, through cross-check of database-recorded Au and Ag grades against retained drill core, confirmed that the database accurately represents

the in situ mineralization distribution and grade; and the SOP and QA/QC program review found i-80’s Standard Operating Procedures, QA/QC insertion rates, acceptance criteria, and sample retention

practices fit for purpose and consistent with industry best practice. This opinion is subject to the limitations on data verification identified in Section 9.2, which are accounted for in the QP’s mineral resource classification.

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10

Mineral Processing and Metallurgical Testing

Metallurgical test work supporting the Granite Creek Underground Project comprises five test programs completed between 2005 and 2026: an

initial autoclave pre-treatment program by Dawson Metallurgical Laboratories (2005-2006); a comprehensive sample characterization, comminution, and pressure oxidation test program by FLSmidth (2022-2023); a

pilot-scale continuous pressure oxidation program by FLSmidth (2025-2026) on a blended feed representing the early years of commercial operation across three i-80 Gold properties; a pressure oxidation and

cyanidation program by SGS focused on the South Pacific Zone (2025-2026); and ongoing bench top autoclave testing performed by the third-party toll milling facility that currently processes Granite Creek Underground production. The nature and extent

of this testing (Section 10.1), the representativeness of the samples tested (Section 10.2), the laboratories involved (Section 10.3), the resulting gold recovery estimates (Section 10.4), and the adequacy of the supporting data (Section 10.5) are

discussed below. Note that all metallurgical test work described in this section pertains to Granite Creek Underground. Testing on Granite Creek open pit samples is historical. No current metallurgical testing has been completed on Granite Creek

Open Pit samples. The current Granite Creek Open Pit work is acceptable for the purposes of resource estimation.

10.1

Metallurgical Testing Programs

A list of the metallurgical test work programs conducted on Granite Creek Underground samples is shown in Table

10-1: an initial autoclave pre-treatment program completed by Dawson Metallurgical Laboratories in 2005 and 2006; a comprehensive sample characterization, comminution,

pressure oxidation, and downstream processing program completed by FLSmidth in 2022 and 2023; a continuous pilot-scale pressure oxidation program completed by FLSmidth in 2025 and 2026 on a blended feed representing the early years of commercial

operation; a pressure oxidation and cyanidation program completed by SGS in 2025 and 2026 focused on the South Pacific Zone (SPZ); and ongoing bench top autoclave testing performed by the third-party toll milling facility currently processing

Granite Creek Underground production. The nature and extent of the testing and analytical procedures used in each program are described below.

Table 10-1: Metallurgical Test Work Programs

Date

Laboratory

Accreditation

Description

2005 / 2006

Dawson Metallurgical Laboratories

No information available. Independent of i80 Gold.

Bench Top Autoclave and CIL

2022

FLSmidth Minerals Testing and Research Center

No accreditations provided on their website.

Independent of i80 Gold.

Bench Top Autoclave and CIL

Bench Top Roaster

and CIL

Continuous Autoclave and CIL

Cyanide Destruction

Solid-Liquid Separation

2025

FLSmidth Minerals Testing and Research Center

No accreditations provided on their website.

Independent of i80 Gold.

Continuous Autoclave and CIL (blend)

Cyanide

Destruction (blend)

Solid-Liquid Separation (blend)

2025

McClelland Laboratories Inc.

No accreditations provided on their website.

Independent of i80 Gold.

Sample preparation and particle Size Analysis

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2025

SGS Lakefield Research Ltd.

Conforms to the requirements of the ISO/IEC 17025 standard for specific registered tests.

Independent of i-80 Gold.

Bench Top Autoclave and CIL

2023 and Ongoing

3rd Party Toll Milling Facility Laboratory

No accreditations provided on their website.

Independent of i80 Gold.

Bench Top Autoclave and CIL

Dawson Metallurgical Program (2005-2006)

Dawson Metallurgical Laboratories completed autoclave metallurgical test work programs on samples from the Granite Creek underground workings

(Ogee samples) on behalf of Atna (Dawson, 2005; Dawson, 2006a; Dawson, 2006b), reported in an initial 2005 program (Ogee Right Rib/Left Rib and RF Zone samples RF_Met-1,

RF_Met-2, and RF_Met-4) followed by a 2006 program (CX Zone samples APCX-204, APCX-211, APCX-219, and APCX-226; the undefined sample AMW-002; and re-tests of the 2005 program’s

Met1 and Met2 samples). The objective of both programs was to determine whether the underground samples, identified as refractory, could be treated using autoclave pre-treatment, as Atna was considering

contracting with a third party for autoclave treatment and downstream processing of the underground material. The scope of the test work included head assays (gold, sulfur speciation, and carbon speciation), baseline cyanide leach shake-out tests on ground feed samples, and pressure oxidation test work: samples were ground to a P80 of 75 µm or 45 µm, acidulated with sulfuric acid to a pH of 1.8 to 2.0 for one hour to digest

carbonate minerals ahead of autoclave treatment (standard methodology for whole-material autoclave treatment in Nevada), then processed in an autoclave at 225°C (437°F) for a one-hour residence time,

35% solids w/w pulp density, and an oxygen overpressure of 3,176 kPa (460 psi). Lime was added to the autoclave residue to raise the pH to 10.0-10.5 prior to a cyanide leach test simulating CIL (carbon-in-leach) processing to determine gold recovery.

FLSmidth Metallurgical Program (2022-2023)

In early 2022, FLSmidth (FLS) was contracted to undertake a series of tests on Granite Creek underground samples, completed in early 2023. The

program’s objectives were sample characterization, mineralogical studies, comminution testing, acid/alkaline batch pressure oxidation (POX) followed by batch cyanidation and preg-robbing tests, a continuous POX test followed by batch

neutralization, cyanidation, cyanide detoxification, and solids-liquids separation testing, and a batch of POX-CIL tests on three composite blends.

Comminution testing (Bond ball mill work index, BWi) was conducted on selected samples, limited to Bond ball mill work index testing since no

comminution circuit design is required for this study — production will initially be toll milled at a Nevada Gold Mines process facility, and eventually processed through the refurbished Lone Tree facility. Tests were run using a 106 µm

closing screen size, targeting a grind size k80 of 75 µm. The overall average work index was 16.4 (imperial, classified as hard), with

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a 75th-percentile (normal design) value of 19.0 (very hard). By zone, the OG Zone averaged 18.9 (very hard, 75th percentile 20.3, very hard) with the oxide samples showing similar hardness to the

sulfide samples; the Otto Zone averaged 16.2 (hard, 75th percentile 17.4, hard); the Adams Peak Zone averaged 12.8 (medium, 75th percentile 13.2, medium); and the Deep Range Front and Range Front Zone samples were classified as medium hardness while

the South Pacific Zone (SPZ) sample was classified as very hard. Mineralogy (XRD) and swelling clay analysis showed gangue mineralogy of quartz, k-feldspar, muscovite, clays (kaolinite and swelling clay), and

calcite, with pyrite and marcasite present in all samples as the primary sulfides; two samples (APL and APLG) showed higher percentages of swelling clay that may require attention during POX treatment — addressed in later test conditions

through trona addition (see below).

Analytical direct cyanide leach shake tests were conducted on pulverized samples from each of the 17

selected samples in centrifuge tubes on a shaker for 60 minutes, with pH adjusted to 10.5 using lime slurry, followed by centrifuging and gold analysis by AA on the pregnant solution, to provide a baseline recovery. Preg-robbing leach tests were

performed similarly, on samples spiked with a stock solution containing a known amount of gold, to measure a preg-rob index (PRI) comparing the amount of gold adsorbed onto the solids to the leached gold.

All metallurgical samples (with the exception of OGOX) were subjected to a series of Batch Autoclave (BTAC) tests under six different

conditions (A through F), summarized in Table 10-2. Test conditions B, C, D, and F replicate Lone Tree autoclave operating conditions. Conditions A and E used acidic

pre-acidulation (one hour, 98% concentrated sulfuric acid) at two temperatures — Condition A added acid in a stoichiometric ratio for carbonate destruction, and Condition E targeted a CO3/S2- weight ratio of 1 — while Conditions C and D tested trona addition (10 kg/t and 5 kg/t respectively) to counteract swelling-clay effects under alkaline conditions. Continuous POX testing was

subsequently used to confirm the batch test results (see below). Carbon-in-Leach bottle roll tests were performed on each sample and on the discharge of each BTAC test,

at CIL conditions of 21°C, 35% solids, 20 g/L carbon, pH 10.5-11, initial cyanide addition of 2.5 g/t NaCN, and a 24-hour residence time.

Table 10-2: Underground Samples Batch Pressure Oxidation Conditions from FLS Program

POX Condition

A

B

C

D

E

F

Acidulation

Yes

No

No

No

Yes

No

POX

Acid

Alkaline

Alkaline

Alkaline

Acid

Alkaline

Trona Dosage (kg/t)

None

None

10

5

None

None

Temperature (oC)

225

199

199

199

199

199

O2 Overpressure (kPag)

689

689

689

689

689

689

Target Gauge Pressure (kPag)

3137

2103

2103

2103

2103

2103

Pulp Density (% solids)

30

30

30

30

30

30

Particle Size, k80 (µm)

75

75

75

75

75

53

Retention Time (min)

60

45

45

45

45

45

Follow-up tests were conducted on three sample blends (Table 10-3) to optimize alkaline POX conditions against benchtop roasting, evaluating longer retention time under alkaline (Lone Tree) conditions and confirming the effect of trona addition, using the five test conditions

summarized in Table 10-4. Benchtop roasting (BTR) was performed in a Carbolite HTR rotary reactor tube furnace: a dry ground sample was weighed into a tared borosilicate glass reactor and subjected to a two-stage roast with oxygen (99.9% purity) gas applied across the roaster bed, at a first-stage temperature of 986°C for 30 minutes, a second-stage temperature of 1,058°C for 15 minutes, a 5°C/minute

ramp rate, an oxygen rate of 2 cc/minute, and a sample feed weight of 500 to 550 g.

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Table 10-3: Granite Creek Underground

Metallurgical Sample Blends for Additional Testing

Sample

Composition

Blend 1

65% OTHG, 34% APHG

Blend 2

50% OTLG, 34% OUT, 16% APL

Blend 3

60% OUT, 27% APHG, 13%

OTLG

Table 10-4: Granite Creek Underground Metallurgical Testing Program

Follow-Up BTAC Test Conditions

Test Condition

I

II

III

IV

V

Acidulation

No

No

No

No

No

POX Condition

Alkaline

Alkaline

Alkaline

Alkaline

Alkaline

Trona Dosage (kg/t)

None

None

20

10

None

Temperature (°C)

199

199

199

199

199

O2 Overpressure (kPag)

689

689

689

689

689

Pulp Density (% solids)

30

30

30

30

30

Particle Size, k80 (µm)

75

75

75

75

75

Retention Time (min)

60

45

45

45

75

A continuous POX test was completed using composite sample OAPC across three conditions (Table 10-5), selected from the batch BTAC results as the conditions producing the highest sulfide oxidation and gold recovery for that sample: Run 1-Condition B (alkaline, no

reagents, six 45-minute turnover intervals), Run 2-Condition E (partially acidulated, six 45-minute turnover intervals), and Run 3-Condition A (fully acidulated, six 60-minute turnover intervals at 225°C/437°F). The three runs operated back-to-back for approximately 16 hours, excluding autoclave heating and cooling time; profile samples were collected once steady state was achieved and again after three volume changeovers, and 16 L of POX

discharge were collected and weighed at each changeover for downstream testing. The continuous autoclave feed rate was approximately 13.31 kg/hour of solids for Runs 1 and 2 (45-minute residence time) and

approximately 9.98 kg/hour for Run 3 (60-minute residence time). Figure 10-1 shows the resulting sulfide oxidation profile by autoclave compartment, from autoclave feed

to autoclave discharge: oxidation began quickly in compartments 1 to 2 and reached near-full oxidation by compartment 3 for Runs 2 and 3, while Run 1 (alkaline) did not reach the same level of oxidation as the other two runs.

Table 10-5: Granite Creek Underground Metallurgical Testing Program Continuous POX Run Test

Conditions

Test Condition

B (Run 1)

E (Run 2)

A (Run 3)

POX Condition

Alkaline

Acid

Acid

Acidulation

No

Partial

Complete

Trona Dosage (kg/t)

0

0

0

Temperature (oC)

199

199

225

O2 Overpressure (kPag)

689

689

689

Total Pressure (kPag)

2096

2096

3130

Pulp Density (% solids)

30

30

30

Particle Size, k80 (µm)

75

75

75

Retention Time (min)

45

45

60

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Source: TR Raponi Consulting Ltd. (2023)

Figure 10-1: Granite Creek POX Pilot Plant Sulfide Oxidation Profile

Cyanide destruction tests were performed using the SO2/air process, originally

developed and patented by Inco Ltd. (now Vale), in which SO2 plus air oxidizes cyanide into cyanate, catalyzed by the addition of copper ions; typical retention times to achieve less than 5

mg/L weak-acid-dissociable cyanide (CNWAD) are 1 to 2 hours at pH 7.5 to 9.5, and the process is capable of achieving discharge concentrations below 1 mg/L CNWAD (though it is not suited to directly reducing total cyanide (CNT). SO2 is typically supplied as sodium metabisulphite (SMBS) solution

dissolved on site, or as elemental sulfur combusted to generate SO2 on site for larger users. A total of 32 L of transitional continuous POX discharge material, collected from the beginning of

Run 2, was used to conduct a bulk 24-hour Carbon-in-Leach test at standard CIL conditions; the resulting slurry then underwent

cyanide detox testing with continuous SMBS and copper sulfate addition, at the two sets of conditions summarized in Table 10-6. The objective in this program was to achieve a discharge concentration of <50

mg/L CNWAD.

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Table 10-6: Underground Blend Samples Cyanide

Detox Conditions from FLS Program

Test Number

1

2

Feed Slurry

Bulk CIL

Oxidizing Agent

SMBS

Feed CNWAD (mg/L)

177

182

SO2:CNWAD (g/g)

8

5

Retention Time (h)

1

1

Copper (mg/L)

25

25

pH range

8.0

-

9.0

8.0

-

9.0

Number of Turnovers

6

6

Total Continuous Time (h)

4.5

4.5

Temp °C

25

25

POX discharge slurry samples from Runs 1, 2, and 3, plus the product slurry from Detox Test 1 (the POX Run 2

sample), were retained for solids-liquids separation testing to determine optimum thickener sizing and operating parameters, evaluating pre-leach thickener duty for the POX discharge samples and tailings

thickener duty for the cyanide detox sample. Flocculant screening on the Run 1 sample tested five flocculants (905VHM, 910VHM, 913VHM, 923VHM, and 934VHM); 913VHM (an anionic polyacrylamide flocculant of medium molecular weight and low charge

density, substitutable with any comparable product) was found to provide the best overflow clarity and settling velocities and was selected for subsequent testing on all samples. FLSmidth also conducted pressure filtration tests using a bench-scale

unit capable of simulating FLSmidth’s recessed-chamber and membrane-squeeze-chamber configurations across a range of feed solids concentrations, pressure profiles, and cake thicknesses; the Detox

Tailings sample was tested at a feed of 48% solids by weight to simulate feeding from a thickener underflow.

FLSmidth Metallurgical

Program (2025-2026)

In 2025, a program was undertaken at the pilot-scale autoclave facility at the FLSmidth Minerals Testing and

Research Center in Salt Lake City, Utah, to simulate continuous pressure oxidation conditions similar to those expected in the commercial plant, using a blended feed made up of ore from Granite Creek, Ruby Hill, and Cove — the three i-80 operations expected to supply feed in the early years of commercial operation. The block flow diagram of the pilot-scale flow sheet is shown in Figure 10-2. The program

comprised feed preparation (crushing and grinding), blending, acidulation (where specified), continuous pressure oxidation, neutralization, CIL, and cyanide detoxification; various products of the program were also used for bulk CIL, cyanide detox,

solids-liquid separation, and slurry rheology testing.

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Source: i-80 Gold

Figure 10-2: Continuous Pressure Oxidation Program Flow Sheet

For the Granite Creek portion of the composite, rejected sample material from active underground mining was used; for Ruby Hill, available core

from the 426 Zone (planned for the early years of mining) was used; and for Cove, the UHPRODCOMP sample, composited to represent the early years of mining, was used. Blend proportions (Table 10-7) were

determined from the expected feed tonnages of the early years of plant operation from the three sources: 26% Granite Creek, 58% Ruby Hill, and 16% Cove.

Table 10-7: Proportions of Continuous Pressure Oxidation Feed from

i-80 Sources

Project

Proportion of Blend (%)

Granite Creek

26

Ruby Hill

58

Cove

16

The three source materials were prepared individually before blending. Samples were stage-crushed to -10 mesh, then ground in 12 kg batches at 60% solids with water using a rod mill, targeting a P80 of 75 microns; grind studies determined the correct grind time for each sample via dip sampling, wet screening

(coarse +20 micron and fine -20 micron fractions), drying at 50°C (122°F), and Ro-Tap sizing of the coarse fraction. Leftover material from the continuous test

(constituents and blend) was analyzed by McClelland Labs of Reno, Nevada for particle size (Table 10-8), and McClelland subsequently performed bench-scale autoclave tests on each of the components.

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Table 10-8: Particle Size Analysis of Continuous

Pressure Oxidation Blend and Components

Sieve Size (µm)

Cumulative Weight Passing (%)

Granite Creek

Ruby Hill

Cove

Blend

150

98.9

96.1

98.4

96.2

104

91.3

90.7

92.6

90.6

74

76.9

82.0

79.1

77.5

53

67.4

74.9

66.6

69.8

43

61.7

70.0

60.3

63.9

38

58.9

67.6

57.6

62.5

P80 (µm)

79.3

67.3

75.3

78.7

Three conditions were tested over three continuous runs, selected to match the expected temperature and

pressure conditions of the Lone Tree autoclave, in either alkaline (Run 1) or acidic (Runs 2 and 3) mode, with 45-minute (Runs 1 and 2) or 60-minute (Run 3) retention

time (Table 10-9). Samples for Runs 2 and 3 were acidulated by slowly adding the stoichiometric quantity of sulfuric acid required for full carbonate destruction over 2 to 3 hours to prevent bubbling and

frothing; the acidified slurry was noticeably more viscous than the alkaline Run 1 feed. Testing used a titanium continuous autoclave with a dynamic operating volume of 27 L (including gas hold-up), a flash

discharge system, and six compartments (the first two only partly divided and treated as a single compartment, with cooling water addition available to manage temperature there); a photograph of the autoclave is shown in Figure 10-3. Set points for pre-heat and shell temperature were determined from prior projects of similar sulfide grade and target temperature. At the end of each run, a set of seven

profile samples was taken (POX feed; compartment samples 1-2, 3, 4, 5, and 6; and POX discharge), and a total of 50 L of continuous autoclave discharge from the end of Run 2 was collected as a bulk sample.

Table 10-9: Continuous Pressure Oxidation Test Conditions

Test Condition

Run 1

Run 2

Run 3

Acidulation

No

Yes

Yes

Temperature (°C)

199

199

199

O2 Overpressure (kPag)

689

689

689

Total Pressure (kPag)

2,103

2,103

2,103

% Solids by Weight

30

30

3,0

Retention Time (min)

45

45

60

Total Run Time (h)

3.5

4.0

4.5

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Source: FLSmidth (2026)

Figure 10-3: Continuous Pressure Oxidation Autoclave at FLSmidth

The slurry temperature profile of the autoclave compartments over the course of testing is shown in Figure

10-4; the target temperature of 199°C (390°F) was reasonably met, with some notable excursions in Run 2. The overall autoclave pressure, shown inFigure 10-5,

showed no significant variation from the target of 2,103 kPag (305 psig) throughout testing. Vent oxygen purity, targeted at 90%, is shown in Figure 10-6; purity dropped below target at times, likely due to

carbonate decomposition during the alkaline phase and at the start of acidic pressure oxidation as acidulated feed mixed with the autoclave’s alkaline contents, then increased as the acidic runs proceeded. The higher viscosity of the

acidulated slurry caused several upset events where feed pumping was halted due to obstruction; water was added during the acidulated runs to improve pumpability, with pumping rate slowed to compensate and maintain the required retention time,

reflected in the calculated feed flow rate shown in Figure 10-7. The progression of sulfide oxidation over autoclave residence time, from the profile samples, is shown in Figure

10-8.

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Source: FLSmidth (2026)

Figure 10-4: Continuous Pressure Oxidation Slurry Temperature Profile

Source: FLSmidth (2026)

Figure 10-5: Continuous Pressure Oxidation Vessel Pressure Profile

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Source: FLSmidth (2026)

Figure 10-6: Continuous Pressure Oxidation Vent Oxygen Purity

Source: FLSmidth (2026)

Figure 10-7: Continuous Pressure Oxidation Calculated Feed Flow Rate

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Source: FLSmidth (2026)

Figure 10-8: Continuous Pressure Oxidation Sulfide Oxidation Profile

Following pressure oxidation, samples were neutralized in a hot water bath at 60°C (140°F) with overhead mixing and air sparging, with

lime added over a 3-hour period to target a pH of 10.5 (the bulk sample collected during pressure oxidation was not neutralized). Neutralized samples were then subjected to bench-scale CIL leach tests using

activated carbon, similar to the bench-scale program; the bulk sample was leached in a bulk CIL reactor for 24 hours using the same conditions but with pH adjustment during leaching rather than bulk neutralization, and was subsequently used for

cyanide detox and solids-liquids separation testing. Each of Runs 1 through 3 was tested in duplicate CIL runs. Benchtop autoclave tests on the same blended feed used in the continuous test were CIL-leached

using the same conditions as the prior bench top program for comparison.

The bulk sample, after 24 hours of CIL leaching, was screened to

remove carbon and dosed with sodium metabisulfite (SMBS) as the SO2 source and copper sulfate as catalyst, in a series of two tests at the conditions shown in Table 10-10. In both tests, CNWAD concentrations

were reduced to below 5 mg/L after one hour of retention time, as shown in Figure 10-9.

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Table 10-10: Continuous Pressure Oxidation

Testing, Cyanide Detoxification Conditions

Test 1

Test 2

Feed CNWAD (mg/L)

110

106

SO2/CNWAD (g/g)

11

6

Retention Time (h)

1

1

Copper Concentration (mg/L)

43

31

pH Range

8-10

8-10

Number of Turnovers

4

4

Total Continuous Time (h)

4

4

Temperature (°C)

25

25

Lime Consumption (kg/t)

1.87

1.87

SO2 Consumption (kg/t)

2.77

1.40

CuSO4 Consumption (kg/t)

0.10

0.07

Lime:SO2 (w/w)

0.67

1.34

Source: FLSmidth (2026)

Figure 10-9: CNWAD Concentration in Cyanide

Detoxification

Subsamples of material from the continuous pressure oxidation program were used for thickening, rheology, and

filtration test work at FLSmidth. Malvern laser diffraction particle size analyses of the autoclave feed, autoclave discharge, and detox tailings samples are shown in Table 10-11 (the AC Feed sample,

representative of autoclave feed during the continuous test, was analyzed after acidulation). Flux testing determined the optimum feedwell suspended solids concentration ranges for flocculation shown in Table

10-12 — quite low ranges, indicating significant internal dilution will be required in thickener feedwells for optimum settling. Continuous fill tests measuring predicted thickener underflow densities

are shown in Table 10-13: acceptable underflow densities were achieved for the AC Discharge and Detox Tails samples at bed retention times of up to two hours (typical for an operating thickener), though it is

unclear why the AC Feed sample produced a comparatively low

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underflow density; rise rates were appropriate for a high-rate thickener while providing acceptable overflow clarity. Dynamic rheology testing, performed at expected underflow densities using a cup-and-bob apparatus on underflow from the continuous fill tests, is shown in Table 10-14; underflow yield stresses were found to be

acceptable for centrifugal pumping, and dynamic viscosity analysis showed reduced viscosity with increasing shear rate. Pressure filtration testing on the Detox Tails sample, using recessed chambers with a bench-scale filtration unit across a range

of feed solids concentrations, pressure profiles, and cake thicknesses, is shown in Table 10-15; the final cake moisture was high relative to the specification needed for filtered tailings stacking, and

filtration rates of 63 to 91 kg/m2/h were roughly an order of magnitude lower than typical tailings filtration, attributed to the fine feed particle size distribution and the oxidized nature of the solids — further thickening and filtration

test work on the products of pressure oxidation, CIL, and cyanide detoxification will be required to reach targeted moisture and filtration rates.

Table 10-11: Malvern Laser Diffraction Particle Size Analysis of Continuous Pressure Oxidation

Samples

AC Feed

AC Discharge

Detox Tails

D90 (µm)

75

76

50

D80 (µm)

~49

43

~23

D50 (µm)

13

12

6

D20 (µm)

~2.3

2.7

~1.7

D10 (µm)

1.2

1.4

1.0

Table 10-12: Required Solids Feed Concentration Ranges

Sample

Solids

Feed

Concentration Range for

Highest Flux Rate (%)

AC Feed

1 – 2

AC Discharge

5 – 7

Detox Tails

3 – 5

Table 10-13: Continuous Fill Tests on Continuous Pressure Oxidation

Samples

AC Feed

AC Discharge

Detox Tails

Feed Suspended Solids Concentration (%)

1.5

6

5

Flocculant Dose Range Tested (g/t)

50 - 100

60 - 80

40 -65

Unit Area Range Tested (m2/tpd)

0.13 - 0.25

0.04 -0.10

0.042 - 0.10

Internal Rise Rate Range Tested (m/h)

12.0 - 23.3

10.6 - 18.7

9.1 - 21.7

Est. Bed Suspended Solids – 0.5 hr (wt%)

15.3

35.9

36.5

Est. Bed Suspended Solids – 1 hr (wt%)

18.7

41.0

41.5

Est. Bed Suspended Solids – 2 hr (wt%)

23.2

44.8

45.8

Est. Bed Suspended Solids – 3 hr (wt%)

26.5

47.1

47.0

Est. Bed Suspended Solids – 6 hr (wt%)

28.3

48.2

47.5

Recommended Floc Dose Range (g/t)

70 - 80

70 - 80

55 - 65

Design Unit Area (m2/tpd)

0.25

0.07

0.07

Design Overflow Clarity (mg/L)

55 - 65

75 - 90

< 50

Table 10-14: Slurry Rheology on Continuous Pressure Oxidation

Samples

AC Feed

AC Discharge

Detox Tails

Underflow Density (wt%)

19

40

41

Yield Stress (Pa)

1.65

40.5

18.7

Bingham Plastic Viscosity (cP)

14.4

20.6

28.7

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Table 10-15: Pressure Filtration on Cyanide

Detoxification Samples

Detox Tails

Chamber Type

Recessed

Filter Media

POPR 966

Feed Suspended Solids (wt%)

41

Feed Pressure (kPa)

1,000 & 1,500

Fill Time (min)

5.25 – 9.0

Cake Thickness (mm)

32 & 50

Air Blow Pressure (kPa)

310 – 700

Air Blow Time (min)

10

Filtrate Quality (ppm)

5,243 – 7,219

Ultimate Cake Moisture (wt%)

28.7 – 29.9

Dry Cake Density (kg/m3)

1,202 – 1,236

Final Filtration Rate (kg/m2/h)

63 – 91

SGS Metallurgical Program (2025-2026)

In 2025, a metallurgical test program was conducted at SGS in Lakefield, Ontario, Canada, focused on the South Pacific Zone (SPZ), for which

there was not a significant amount of prior test work; the program focused on pressure oxidation followed by gold recovery by CIL and was completed in 2026.

Bond ball mill grindability tests were performed on samples SPZN, SPZCOM1, SPZCOM2, and SPZCOM3 at a 105 µm grind size (Table 10-16).

Table 10-16: Bond Ball Mill Grindability Test

Results

Sample ID

Grind Size

(µm)

F80

(mm)

P80

(mm)

Work Index

(kWh/t)

Imperial

Work Index

(kWh/t)

Metric

SPZN

105

2,485

85

16.7

18.4

SPZCOM1

105

2,446

80

14.4

15.9

SPZCOM2

105

2,108

73

13.8

15.3

SPZCOM3

105

2,618

76

18.0

19.8

Pressure oxidation test work was undertaken on the eight SPZ samples to evaluate the effectiveness of

oxidative treatment in overcoming their refractory nature, at a grind target P80 of 75 µm, pre-acidulation residence time of 6 hours (where required), pulp density of 35% by weight, temperature of

199°C (390°F), pressure of 2,048 kPag, and residence time of 45, 60, or 75 minutes. Four tests per sample were run: one with no pre-acidulation and three with

pre-acidulation at varying residence time, using 2 L titanium autoclave vessels; acid addition (where required) was applied over a six-hour pre-acidulation period with hourly pH and ORP readings. Initial tests based pre-acidulation on the carbonate assay of each feed (targeting a 1:1

carbonate-to-sulfide ratio), but early results (POX tests 2R, 6, 10, 14, 22, and 26) showed solids clumping, unreacted solids, and poor sulfide oxidation; pre-acidulation was subsequently modified to target a pH of 2, which increased sulfide and arsenopyrite oxidation and gold recovery from CIL. Final POX pulp pH across the modified tests averaged 1.24, with an

average ORP of 581 mV (excluding POX 7, whose ORP of 21 mV likely reflected a malfunctioning probe).

Standard CIL bottle roll tests were

conducted on each of the eight samples, without POX pre-treatment, to provide a baseline recovery, at pH 10.5, NaCN concentration of 1.0 g/L, a 6-hour pre-aeration duration, and a 24-hour CIL residence time. POX residues were subsequently forwarded for CIL testing using the same conditions as the baseline CIL tests.

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Third-Party Toll Milling Bench Top Autoclave Tests

As part of the toll milling process, bench top autoclave tests are conducted at the on-site laboratory

of the third-party toll milling facility that currently processes Granite Creek Underground production, on a composite of each monthly delivery lot, to assign a gold recovery value to that lot’s production; the result is used, together with

the results from other feed components delivered that month, to reconcile the plant’s actual gold recovery to each individual component. The bench top tests are conducted at conditions analogous to the plant: a temperature of 225°C

(437°F), a pressure of 3,172 kPag (460 psig), and a residence time of 60 minutes.

10.2

Sample Representativeness

The degree to which the metallurgical test samples are representative of the various types and styles of mineralization at Granite Creek

Underground, and of the deposit as a whole, is discussed below by test program and supported by the spatial distribution and grade/composition variability of the samples tested.

FLSmidth 2022 Program Samples

The FLSmidth 2022 program tested 17 samples selected to span the primary mineralized zones of the Granite Creek Underground deposit — OG

Zone (upper and lower sulfide, oxide, and high-/low-grade variability samples), Otto Zone (upper and lower sulfide and high-/low-grade variability samples), Adams Peak

Zone (upper, lower, and high-/low-grade variability samples), the Otto/Adams Peak composite (OAPC), and variability samples from the Deep Range Front, Range Front, and South Pacific Zones — together with

dedicated comminution samples for each zone. Head assays for the sample set are shown in Table 10-17, ranging from 4.56 g/t (Adams Peak Zone lower) to 39.43 g/t Au (OG Zone comminution sample), with sulfide

sulfur content ranging up to 5.84% (Adams Peak Zone upper), reflecting the range of mineralization styles and grades present across the deposit.

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Table 10-17: Underground Samples Head Assays from

FLS Program

Sample

Number

Description

ID

Au g/t

Au oz/

st

Ag g/t

Ag oz/

st

S=%

Corg%

1

OG Zone Upper (Sulfide)

OGU

14.54

0.424

4.63

0.135

1.96

0.44

2

OG Zone Lower (Sulfide)

OGL

28.46

0.830

9.39

0.274

2.98

0.12

OG Zone Oxide

OGOX

12.48

0.364

11.55

0.337

0.43

0.34

3

OG Zone High Grade (Sulfide) Variability Sample

OGHG

37.71

1.100

7.41

0.216

3.89

0.10

4

OG Zone Low Grade (Sulfide) Variability Sample

OGLG

8.43

0.246

22.29

0.650

2.73

0.36

OG Zone Comminution Sample No. 1

OGCOM1

12.75

0.372

153.26

4.470

2.33

0.33

OG Zone Comminution Sample No. 2

OGCOM2

39.43

1.150

9.77

0.285

0.04

0.03

OG Zone Comminution Sample No. 3

OGCOM3

10.39

0.303

7.10

0.207

0.15

0.02

OG Zone Comminution Sample No. 4

OGCOM4

9.94

0.290

4.22

0.123

0.00

0.19

OG Zone Comminution Sample No. 5

OGCOM5

26.23

0.765

3.70

0.108

0.00

0.03

OG Zone Comminution Sample No. 6

OGCOM6

12.89

0.376

1.92

0.056

1.41

1.12

5

Otto Zone Upper (Sulfide)

OTU

16.70

0.487

4.01

0.117

1.14

0.18

6

Otto Zone Lower (Sulfide)

OTL

14.98

0.437

10.94

0.319

1.67

0.38

7

Otto Zone High Grade (Sulfide) Variability Sample

OTHG

20.40

0.595

12.89

0.376

0.79

0.19

8

Otto Zone Low Grade (Sulfide) Variability Sample

OTLG

6.24

0.182

1.10

0.032

1.48

0.22

Otto Zone Comminution Sample No. 1

OTCOM1

8.40

0.245

0.00

0.000

3.01

0.11

Otto Zone Comminution Sample No. 2

OTCOM2

13.92

0.406

0.99

0.029

1.81

0.47

Otto Zone Comminution Sample No. 3

OTCOM3

9.50

0.277

1.06

0.031

2.84

0.27

9

Adams Peak Zone Upper

APU

7.65

0.223

0.93

0.027

5.84

0.33

10

Adams Peak Zone Lower

APL

4.56

0.133

4.90

0.143

4.29

0.25

11

Adams Peak Zone High Grade Variability Sample

APHG

20.43

0.596

0.99

0.029

2.86

0.20

12

Adams Peak Zone Low Grade Variability Sample

APLG

6.45

0.188

8.23

0.240

3.20

0.13

Adams Peak Zone Comminution Sample No. 1

APCOM1

11.04

0.322

3.12

0.091

3.12

0.37

Adams Peak Zone Comminution Sample No. 2

APCOM2

11.76

0.343

15.05

0.439

3.85

0.17

13

Otto/Adams Peak Zone Composite

OAPC

11.35

0.331

3.67

0.107

2.25

0.24

14

Deep Range Front Zone Variability Sample

DRFV

6.31

0.184

2.78

0.081

1.73

0.26

15

Range Front Zone Variability Sample

RFV

7.30

0.213

2.85

0.083

2.33

0.17

16

South Pacific Zone Variability Sample

SPZV

20.16

0.588

1.23

0.036

2.79

0.75

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FLSmidth 2025 Program Feed Blend

The FLSmidth 2025 continuous pressure oxidation program used a blended feed representing the early years of commercial operation across three i-80 properties — 26% Granite Creek, 58% Ruby Hill, and 16% Cove (Table 10-18) — rather than Granite Creek material alone; the assayed head composition of each

source and of the blend is shown in Table 10-18. The blend’s composition (11.8 g/t Au, 18.2% carbonate, and 1.94% sulfide sulfur, assayed) indicates a high level of carbonate and sulfide sulfur relative

to the design parameters of the Lone Tree pressure oxidation plant. Because the blend includes material from the Adam’s Peak Zone, which is not part of the Granite Creek Underground mine plan, recovery data from the continuous tests on this

blend were not used directly in Granite Creek Underground recovery modeling (see Section 10.4); the blend nonetheless demonstrates the pilot-scale continuous process response of Granite Creek material processed together with other i-80 feed sources under commercial-scale conditions.

Table

10-18: Composition of Continuous Pressure Oxidation Feed Blend

Project

Au

(g/t)

Au

(oz/st)

Corg

(%)

Ctotal

(%)

CO3

(%)

Stotal

(%)

Ssulfide

(%)

Granite Creek

9.5

0.277

0.39

3.2

13.8

1.8

1.52

Ruby Hill

6.8

0.198

0.11

3.3

16.1

2.7

2.43

Cove

15.1

0.440

0.11

6.7

32.9

0.9

0.86

Blend (assayed)

11.8

0.344

0.16

3.8

18.2

2.0

1.94

SGS 2025 Program Samples

The SGS 2025 program tested eight half-HQ-core samples selected

to represent the South Pacific Zone (SPZ) — south, middle, and north variability samples, high- and low-grade variability samples, and three comminution samples (Table

10-19). Gold grades of the delivered samples ranged from 5.34 to 21.8 g/t, with composite samples displaying a consistent gold grade around 15 g/t; gold cyanide analysis showed three samples (SPZM, SPZN, and

SPZCOM2) were highly refractory (AuCN less than 10% of total gold) while three others (SPZS, SPZHG, and SPZCOM3) displayed partial free gold or weak refractoriness. Sulfide sulfur levels varied from 1.3% to 3.4%, mercury from 42 to 256 g/t (likely

cinnabar), arsenic from 0.19% to 1.91%, total carbon from 0.47% to 4.7%, and carbonate from 0.49% to 20.2% — head assays for the sample set are shown in Figure 10-10.

Table 10-19: South Pacific Test Program Sample List

Sample ID

Sample Description

SPZS

South Pacific Zone South Variability Sample

SPZM

South Pacific Zone Middle Variability Sample

SPZN

South Pacific Zone North Variability Sample

SPZHG

South Pacific Zone High Grade Variability Sample

SPZLG

South Pacific Zone Low Grade Variability Sample

SPZCOM1

Comminution Sample No. 1

SPZCOM2

Comminution Sample No. 2

SPZCOM3

Comminution Sample No. 3

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Source: SGS

Figure 10-10: SGS Metallurgical Program Head Assays

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Mineralogical Representativeness (SGS Samples)

Bulk modal mineralogy conducted on four SGS samples (SPZS, SPZN, SPZHG, and SPZCOM2), shown in Figure

10-11, found the samples primarily composed of quartz (41% to 68%), calcite (4.6% to 32.7%), sericite/muscovite (5.8% to 11%), and clays (9.3% to 10.2%), with SPZCOM2 reporting the highest sulfide

concentration at 6.1%. The degree of particle liberation of pyrite-marcasite, arsenopyrite, silicates, and carbonates by mass is shown in Figure 10-12 through Figure

10-15: SPZHG showed the highest liberation of pyrite and arsenopyrite, while SPZCOM2 showed poor sulfide liberation and highly locked arsenopyrite; all samples showed poor liberation of arsenopyrite, with no

free arsenopyrite present, suggesting the refractory gold in the SPZ is likely associated with arsenic.

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Source: SGS, 2025

Figure 10-11: Bulk Modal (QEMSCAN) Results

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Source: SGS (2026)

Figure 10-12: Pyrite/Marcasite Liberation by Mass

Source: SGS (2026)

Figure 10-13: Arsenopyrite Liberation by Mass

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Source: SGS (2026)

Figure 10-14: Silicate Liberation by Mass

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Source: SGS (2026)

Figure 10-15: Carbonate Liberation by Mass

Location of Metallurgical Samples

Figure 10-16 and Figure 10-17 show the locations of the test

samples for the recent FLSmidth and SGS test programs within the mining shapes of the Granite Creek deposits, in section and plan view respectively. From a spatial perspective, selection of samples for the Granite Creek metallurgical test programs

appears to represent the extents of the deposit.

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Source: i-80 Gold

Figure 10-16: Locations of Metallurgical Samples (Section View)

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Source: i-80 Gold

Figure 10-17: Locations of Metallurgical Samples (Plan View)

Variability and Representativeness of Metallurgical Samples

Figure 10-18 through Figure 10-21 are histograms illustrating

the variability of key attributes across the samples of the Granite Creek Underground test programs — gold grade, organic carbon content, carbonate content, and sulfide sulfur content.

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Source: i-80 Gold

Figure 10-18: Distribution of Gold Grade in Metallurgical Samples

Source: i-80 Gold

Figure 10-19: Distribution of Organic Carbon in Metallurgical Samples

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Source: i-80 Gold

Figure 10-20: Distribution of Carbonate in Metallurgical Samples

Source: i-80 Gold

Figure 10-21: Distribution of Sulfide Sulfur in Metallurgical Samples

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Table 10-20 shows the projected life-of-mine range of gold grade, organic carbon, carbonate, and sulfide sulfur content for Granite Creek Underground autoclave feed, and Table

10-21 compares the average and range of these feed characteristics for the FLS 2023 and SGS 2025 test programs against that projected feed. Based on this comparison, the grades of the samples tested — in

terms of gold, organic carbon, carbonate, and sulfide sulfur — are considered representative of the feed expected during mining.

Table 10-20: Projected Feed from Granite Creek Underground

Gold Grade

(g/t)

Gold Grade

(oz/st)

Organic

Carbon

(%)

Carbonate

(%)

Sulfide

Sulfur

(%)

Life of Mine

7.85

0.229

0.30

11.3

1.24

Annual Maximum

9.77

0.285

0.38

12.7

1.71

Annual Minimum

7.65

0.223

0.20

9.4

0.78

Table 10-21: Summary of Feed Values for Granite Creek Test

Programs

Gold Grade

(g/t)

Gold Grade

(oz/st)

Organic Carbon

(%)

Carbonate

(%)

Sulfide Sulfur

(%)

FLS 2023 Program

Average

14.6

0.426

0.28

8.8

2.2

Maximum

39.5

1.152

1.12

21.4

5.8

Minimum

4.5

0.131

0.02

0.01

0.0

SGS 2025 Program

Average

12.9

0.376

0.70

8.3

2.1

Maximum

21.8

0.636

1.30

20.2

2.7

Minimum

5.3

0.155

0.33

0.5

1.1

10.3

Laboratories

Metallurgical testing on Granite Creek Underground samples has been performed at five independent, commercial or contract laboratories and

facilities, none of which is affiliated with i-80 Gold Corp. (the registrant): Dawson Metallurgical Laboratories (Salt Lake City, Utah), which completed the 2005-2006 autoclave

pre-treatment program on behalf of Atna; FLSmidth, whose FLSmidth Minerals Testing and Research Center pilot-scale autoclave facility in Salt Lake City, Utah performed both the 2022-2023 bench-scale/continuous

test program and the 2025-2026 continuous pilot-plant program; McClelland Laboratories, Inc. (Reno, Nevada), which performed particle size analysis and bench-scale autoclave testing on the components and blend of the 2025 FLSmidth continuous

program; SGS, whose Lakefield, Ontario, Canada facility performed the 2025-2026 South Pacific Zone pressure oxidation and cyanidation test program; and the on-site laboratory of the third-party toll milling

facility that currently processes Granite Creek Underground production, which performs ongoing bench top autoclave testing on monthly production lots as part of the toll milling reconciliation process. Each of these laboratories is an established,

independent commercial metallurgical testing facility with no ownership, financial, or other relationship to i-80 Gold Corp. The source technical documentation reviewed by the QP does not state the specific

accreditation or certification status (e.g., ISO 17025) held by each laboratory for the metallurgical testing performed; accordingly, no representation of formal accreditation is made here beyond the independent, commercial nature of each facility.

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10.4

Relevant Results

The results of the metallurgical test programs, and the basis for recovery estimates and assumptions used in this technical report summary, are

presented below by test program, followed by the QP’s synthesis of expected gold recovery by mining zone and process route.

Dawson Program Results

The cyanide leach gold recoveries from the Dawson baseline tests and autoclave pre-treatment tests are

shown in Table 10-22. The refractory material responded well to autoclave pre-treatment: for samples with both a baseline cyanide leach (no pre-treatment) and an autoclave-residue cyanide leach, average recovery increased from 52% (baseline) to 92% (post-autoclave). There is a wide range of baseline gold recoveries — from 11% to 86% — among

samples with measured sulfide sulfur; an attempted correlation between sulfide sulfur content and baseline gold recovery had a very low coefficient of determination and is not presented. There was an inverse relationship between sulfide sulfur

content and cyanide solubility (no autoclave treatment), shown in Figure 10-22, with a reasonable correlation between solubility and sulfide grade; prior test work by McClelland Labs had found that sulfide

sulfur content did not correlate well with cyanide solubility, and other factors (such as the presence of organic carbon, and gold not directly associated with pyrite) also appear to influence cyanide solubility.

Table 10-22: Autoclave Pre-Treatment Tests from Dawson Test

Work Program

Sample

Year of

Test

Work

Program

Grind

P80

(µm)

Gold

Head

Assay

(g/t)

Gold

Head

Assay

(oz/st)

Total

Sulfide

Sulfur

(%)

Total

Carbon

(CO2)

Head

Assay

(%)

Cyanide Leach Gold

Recovery (%)

Baseline

Tests

Tests on

Autoclave

Residue

Ogee Samples

Ogee (Right Rib + Left Rib)

2005

75

13.71

0.400

0.0

0.82

86

93

RF Zone Samples

RF_Met-1 (33941)

2005

75

8.23

0.240

1.21

2.27

52

93

RF_Met-2 (33942)

2005

75

14.74

0.430

2.61

1.76

61

95

RF_Met-4 (34259)

2005

75

14.74

0.430

2.32

2.43

11

89

CX Zone Samples

APCX-204

2006

75

9.26

0.270

0.00

5.34

94

N/A

APCX-211

2006

75

11.31

0.330

0.00

4.29

85

N/A

APCX-219

2006

75

11.31

0.330

0.84

0.77

60

91

APCX-226

2006

45

19.20

0.560

1.53

2.70

42

94

Undefined Samples

AMW-002

2006

75

11.31

0.330

0.03

0.35

77

N/A

MET 1

2005

75

17.49

0.510

1.21

2.27

N/

A

93

MET 2

2005

75

10.97

0.320

2.61

1.76

N/

A

95

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Source: i-80 Gold

Figure 10-22: Gold Cyanide Solubility and Sulfide Influence – Ogee Samples

FLSmidth 2022 Batch Autoclave Results

Averaged baseline and batch pressure oxidation CIL results by zone are shown in Table 10-23 Baseline

CIL tests confirmed the refractory nature of the samples, with an average recovery of 31.3% Au across all zones. Acidic BTAC conditions produced the highest gold recoveries: Condition A produced the highest average sulfide oxidation (95.6%) and

average recovery (88.6% Au), ranging from 70.2% (OGLG, 99.5% sulfide oxidation) to 95.7% (SPZV, 99.6% sulfide oxidation); Condition E produced the next-highest average sulfide oxidation (71.8%) and average recovery (79.0% Au), ranging from 42.2%

(APHG, 31.8% sulfide oxidation) to 94.5% (OGL, 61.7% sulfide oxidation). Alkaline conditions produced lower sulfide oxidation and recovery: Condition F (finer grind) performed best among the alkaline conditions, averaging 60.3% sulfide oxidation and

72.6% recovery; Condition B averaged 59.8% sulfide oxidation and 69.2% recovery; and the trona-addition conditions (C and D) produced the lowest results — Condition C averaged 44.1% sulfide oxidation and 61.5% recovery, and Condition D

averaged 46.7% sulfide oxidation and 64.6% recovery. By zone: Adam’s Peak and Deep Range Front were the most refractory based on baseline CIL recoveries, while the SPZ variability sample (SPZV) responded well to all BTAC conditions, averaging

90.5% recovery across sulfide oxidations ranging from 43.0% (Condition C, 79.0% recovery) to 99.5% (Condition A, 95.7% recovery). OG Zone samples showed a positive trend between gold head grade and recovery and had the highest recoveries of the

primary zones regardless of POX condition, with the highest average recoveries under Condition F followed by Condition A. Otto Zone and Adam’s Peak Zone samples showed no clear grade-recovery relationship; Otto Zone recoveries were highest

under acidic conditions (E followed by A), and Adam’s Peak Zone recoveries were highest under Condition A, with the other three conditions notably lower. The OAPC composite responded best to acidic Conditions A and E and

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poorly to alkaline conditions. RFV and DRFV samples responded well only to Condition A (95.0% and 89.1% recovery respectively), with recoveries below 60% under all other conditions. Overall,

there is a positive trend between sulfide sulfur oxidation and gold recovery, shown in Figure 10-23

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Table 10-23: Underground Samples Baseline and

Batch Pressure Oxidation CIL Results from FLS Program

Parameter

OG Zone

Otto Zone

Adam’s Peak

Otto /Adam’s Peak

Deep Range Front

Range Front

South Pacific

N/A

Baseline Recovery (% Au)

41.50

46.75

9.25

24.00

10.00

34.00

43.00

A

S Oxidation (%)

86.26

99.49

98.87

99.33

95.87

96.62

99.48

Recovery (% Au)

84.12

88.77

90.65

83.18

89.05

94.96

95.65

NaCN Consumption (kg/t)

0.66

0.75

0.66

0.50

0.61

0.31

0.80

B

S Oxidation (%)

71.11

56.10

57.71

44.00

32.95

42.06

98.68

Recovery (% Au)

85.17

67.12

61.33

56.88

42.79

59.32

94.19

NaCN Consumption (kg/t)

1.41

1.07

0.73

1.81

1.62

1.31

1.82

C

S Oxidation (%)

66.56

38.66

32.35

38.22

35.26

38.63

43.01

Recovery (% Au)

75.75

66.71

45.43

52.17

43.33

57.64

79.09

NaCN Consumption (kg/t)

1.11

1.07

1.60

1.79

1.71

1.91

1.17

D

S Oxidation (%)

60.60

37.91

42.81

35.56

31.21

34.76

80.47

Recovery (% Au)

80.40

66.93

52.71

50.99

40.10

57.79

84.98

NaCN Consumption (kg/t)

0.75

1.11

1.65

0.92

1.65

1.90

1.65

E

S Oxidation (%)

65.20

92.98

67.32

68.67

36.42

42.92

98.84

Recovery (% Au)

86.59

89.31

70.02

81.70

41.47

63.04

93.95

NaCN Consumption (kg/t)

1.31

0.99

1.42

0.38

0.47

1.80

1.85

F

S Oxidation (%)

67.68

60.17

58.13

44.89

32.95

45.06

98.80

Recovery (% Au)

86.40

72.76

65.74

61.13

45.18

60.99

94.83

NaCN Consumption (kg/t)

0.76

1.33

0.71

1.07

1.51

1.39

2.08

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Source: TR Raponi Consulting Ltd. (2023)

Figure 10-23: CIL Gold Recovery as a Function of Sulfide Sulfur Oxidation – Underground

Samples

FLSmidth 2022 Follow-Up BTAC and Roasting Results

Actual and predicted sulfide oxidation results from the follow-up BTAC and roasting tests are compared

in Table 10-24, and gold recovery results in Table 10-25. Overall, actual sulfide oxidation exceeded predicted values, and gold recoveries slightly exceeded predicted

values; extended retention times did not significantly improve recovery, and trona addition again did not prove beneficial. Benchtop roasting (BTR) provided superior sulfide oxidation (average 97.0%) and gold recovery (average 83.0%) compared to all

alkaline BTAC blend results.

Table 10-24: Granite Creek Underground Metallurgical Testing

Program Follow-Up BTAC Sulfide Oxidation Results Compared to Predicted Results

Test Condition

Actual Results

Predicted Results

Blend 1

Blend 2

Blend 3

Average

Blend 1

Blend 2

Blend 3

Average

I

48

57

49

51

—

—

—

—

II

43

52

42

46

43

60

50

51

III

39

49

39

42

39

26

33

33

IV

37

35

41

38

36

33

37

35

V

55

52

45

51

—

—

—

—

BTR

99.0

93.9

98.2

97.0

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Table 10-25: Granite Creek Underground

Metallurgical Testing Program Follow-Up BTAC Gold Recovery Results Compared to Predicted Results

Test Condition

Actual Results

Predicted Results

Blend 1

Blend 2

Blend 3

Average

Blend 1

Blend 2

Blend 3

Average

I

46

75

71

64

—

—

—

—

II

65

74

70

70

63

69

66

66

III

67

73

68

69

63

64

64

64

IV

52

73

69

65

62

68

65

65

V

69

77

66

71

—

—

—

—

BTR

84.0

85.0

80.0

83.0

FLSmidth 2022 Continuous POX Results

Continuous POX discharge leached in a hot stir tank showed no significant difference in gold recovery compared to POX discharge leached in a

bottle roll at ambient temperature; the primary effect of the higher temperature was increased lime and cyanide consumption. Table 10-26 compares results across the three continuous conditions: Conditions A

and E (acidic) achieved much higher gold recovery and sulfide oxidation than Condition B (alkaline),but also consumed significantly more lime and recovered much less silver. The acid POX conditions achieved higher sulfide oxidation and therefore

higher gold recoveries but required significantly more lime for neutralization; the resulting increased lime consumption likely caused much of the silver to be locked within jarosite, which did not appear to form under alkaline POX conditions

— a deleterious effect on silver recovery specific to the acidic POX route.

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Table 10-26: Underground Samples (OAPC) Continuous

Autoclave Tests from FLS Program

Parameter

Baseline Results

Run 1 (B) – Alkaline

Run 2 (E) – Partial Acidulation

Run 3 (A) – Full Acidulation

Bottle Roll Conditions

Hot Stir

Bottle Roll

BTAC

Hot Stir

Bottle Roll

BTAC

Hot Stir

Bottle Roll

BTAC

Sulfide Oxidation (%)

49

44

97

69

96

99

CIL Recovery (% Au)

24

63

63

57

89

91

81.7

91

91

83

CIL Recovery (% Ag)

16

50

42

—

3

3

—

2

1

—

Lime Consumption (kg/t)

1.3

5.23

2.99

—

43.3

39.19

—

55.88

52.75

—

Cyanide Consumption (kg/t)

2.45

2.95

1.79

1.81

2.56

1.83

0.38

3.3

1.64

0.50

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FLSmidth 2022 Cyanide Detoxification Results

Reagent consumption in the two cyanide detox tests is shown in Table 10-27 (lime was used during bulk

CIL testing to prepare the feed for both detox tests). Final cyanide (CNWAD) results from both tests, sampled from the feed every two hours and from reactors 1, 2, and the discharge every hour,

are shown in Table 10-28; the additional SMBS added in Test 1 resulted in significantly lower CNWAD compared to Test 2.

Table 10-27: Underground Cyanide Detox Reagent Consumption from FLS Program

Reagent

Detox 1

Detox 2

Lime Consumption (kg/t)

1.44

1.44

SO2 Consumption (kg/t)

3.45

3.42

Lime:SO2 (wt/wt)

0.42

0.60

CuSO4 Maintained (mg/L)

29.6

32.1

CuSO4 Consumption (kg/t)

0.07

0.09

Table 10-28: Underground Cyanide Detox WAD from FLS Program

Time (h)

Detox 1

Detox 2

CNWAD (mg/L)

1

22.5

51.1

2

25.9

42.8

3

23.2

46.1

4

33.2

50.5

FLSmidth 2025 Continuous POX Results

Sulfide oxidation and carbonate decomposition results from the 2025 continuous pressure oxidation test are shown in Table 10-29. Run 2 (partial acidulation) showed lower-than-expected sulfide oxidation in the middle compartments, though results approached expectations by discharge; because acidic Run 2 immediately followed alkaline Run

1, higher residual carbonate levels in the autoclave may have created a lower-oxygen environment due to CO2 evolution, resulting in lower-than-expected sulfide oxidation for Run 2. As a point

of comparison, bench top autoclave tests on the same feed blend at the same conditions are shown in Table 10-30. The amount of lime required to raise the pH of the autoclave discharge slurry to its target

cyanide-leaching pH is shown in Figure 10-24. Gold recovery results from the continuous POX CIL tests (each run tested in duplicate) are shown in Table 10-31, and gold

recovery results from the corresponding bench top autoclave tests, CIL-leached under the same conditions, are shown in Table 10-32. The reason for the lower recoveries

achieved by the continuous pressure oxidation tests compared to the bench top tests, despite similar levels of sulfide sulfur oxidation, is not fully understood; as noted in Section 10.2, this continuous test blend includes Adam’s Peak

Zone material not part of the Granite Creek Underground mine plan, and its recovery results were accordingly not used directly in the zone-by-zone recovery modeling

presented later in this section.

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Table 10-29: Pressure Oxidation Continuous Test,

Sulfide Sulfur Oxidation and Carbonate Decomposition

Run

Sample

Sulfide Sulfur

Assayed

(%)

Carbonate

Assayed

(%)

Sulfide Oxidation,

Relative to Head

(%)

Carbonate

Decomposition,

Relative to Head

(%)

1

Feed

1.94

18.2

C1-2

1.49

16.9

24.4

8.55

C3

1.34

16.7

32.0

9.36

C4

1.32

16.5

33.0

10.5

C5

1.20

16.3

29.1

11.5

C6

1.10

16.0

44.2

13.4

Discharge

1.08

16.4

45.2

11.3

2

Feed

1.94

0.30

98.4

C1-2

1.23

0.08

37.6

99.6

C3

0.74

0.13

62.5

99.3

C4

0.47

0.15

76.2

99.2

C5

0.58

0.16

70.6

99.2

C6

0.52

0.15

73.6

99.2

Discharge

0.25

0.12

87.3

99.4

3

Feed

1.94

0.25

98.6

C1-2

0.51

0.10

74.1

99.5

C3

0.29

0.05

85.3

99.7

C4

0.20

0.05

89.9

99.7

C5

0.12

0.05

93.9

99.7

C6

0.20

0.05

89.9

99.7

Discharge

0.19

0.05

90.4

99.7

2

Bulk

0.41

0.16

78.9

99.1

Table 10-30: Bench Top Autoclave Tests on Continuous Pressure

Oxidation Feed Blend

Sample

Sulfide Oxidation

45-min Baseline

POX

(%)

45-min Acidic

POX

(%)

60-min Acidic

POX

(%)

75-min Acidic

POX

(%)

CONT

48

82

89

80

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Source: FLSmidth (2026)

Figure 10-24: Neutralization Lime Usage in Continuous Pressure Oxidation Discharge

Table 10-31: Continuous Pressure Oxidation, CIL Au Recovery

Test

Au

Recovery

(%)

NaCN

Consumption

(kg/t)

NaCN

Consumption

(lb/st)

Lime

Consumption

(kg/t)

Lime

Consumption

(lb/st)

Run 1a

77

1.38

3.35

2.40

5.83

Run 1b

77

1.31

3.18

2.49

6.05

Run 2a

80

1.82

4.42

42.86

104.16

Run 2b

76

2.02

4.91

34.96

84.96

Run 3a

88

1.62

3.94

49.89

121.24

Run 3b

88

1.49

3.62

56.10

136.33

Bulk CIL

68

0.84

2.04

66.54

161.70

Table 10-32: Bench Top Autoclave, CIL Au Recovery

Sample

45-min

Baseline POX

(%)

45-min

Acidic POX

(%)

60-min

Acidic POX

(%)

75-min

Acidic POX

(%)

CONT

77

94

96

96

SGS Program Results

Results from the SGS pressure oxidation tests on the eight SPZ samples are shown in Table 10-33. The

initial tests, with insufficient acid addition, achieved low sulfide oxidation of 15% to 30%, corresponding to low free acidity, minimal iron in solution, and high residual sulfide and sulfur content; once sufficient acid was added during the pH-2 pre-acidulation, good to excellent sulfide oxidation efficiencies were achieved, confirming that free acid availability is the leading control on POX

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performance for all SPZ samples tested. Results from the baseline CIL tests (without POX pre-treatment) are shown in Table

10-34; these tests averaged 26.9% gold recovery and ranged from 3.6% to 56.0%, confirming the refractory nature of the SPZ material. Results from CIL testing of the POX residues, using the same conditions as

the baseline tests, are shown in Table 10-35: the extent of sulfide oxidation was the controlling factor on gold recovery, with recovery consistently above 90% where POX oxidation was good to excellent, and

lower recoveries always associated with poor oxidation. This dependence of gold recovery on sulfide oxidation is illustrated in Figure 10-25.

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Table 10-33: POX Results

Test ID

Sample ID

Acid

Addition

(kg/t)

Acid Addition (lb/st)

POX Res.

Time

(min)

POX Disch.

Free Acid

(g/

L H2SO4)

POX

Residue S=

(%)

POX

Mass Loss

(%)

POX S=

Oxidation

(%)

POX 1

SPZS

0

0.00

45

0.0

0.82

4.8

28.3

POX 2R

SPZS

356

865.14

45

9.41

0.20

-7.3

80.3

POX 2RR

SPZS

281

682.88

45

12.6

0.27

-2.7

74.6

POX 3

SPZS

290

704.75

60

8.92

0.19

-7.6

81.2

POX 4

SPZS

325

789.81

75

11.3

0.08

-11.4

91.8

POX 5

SPZM

0

0.00

45

0.0

1.45

4.0

23.1

POX 6

SPZM

80

194.41

45

0.0

1.50

-2.2

15.3

POX 6R

SPZM

121

294.05

45

10.3

0.47

-2.1

73.5

POX 7

SPZM

215

522.49

60

20.0

0.16

-6.1

90.6

POX 8

SPZM

148

359.67

75

39.3

0.08

-8.0

95.2

POX 9

SPZN

0

0.00

45

0.0

1.76

2.3

23.9

POX 10

SPZN

177

430.14

45

5.0

0.84

-6.2

60.5

POX 10R

SPZN

217

527.35

45

13.5

1.19

-3.9

45.3

POX 11R

SPZN

298

724.19

60

14.1

0.34

-45.8

78.1

POX 12

SPZN

227

551.65

75

18.5

< 0.05

-8.1

97.6

POX 13

SPZHG

0

0.00

45

0.0

1.57

-0.1

25.1

POX 14

SPZHG

13

31.59

45

0.0

1.52

-0.4

27.3

POX 14R

SPZHG

55

133.66

45

16.0

0.06

2.9

97.2

POX 15

SPZHG

60

145.81

60

16.0

0.51

-16.4

71.7

POX 16

SPZHG

84

204.13

75

23.7

< 0.05

3.9

97.7

POX 17

SPZLG

0

0.00

45

0.0

1.92

-1.7

27.6

POX 18

SPZLG

107

260.03

45

20.4

0.07

1.9

97.5

POX 19

SPZLG

111

269.75

60

23.4

< 0.05

2.8

98.2

POX 20

SPZLG

123

298.91

75

21.7

< 0.05

-18.2

97.8

POX 21

SPZCOM1

0

0.00

45

0.0

1.69

2.7

24.9

POX 22

SPZCOM1

56

136.09

45

0.0

1.84

0.2

16.2

POX 22R

SPZCOM1

96

233.30

45

15.8

0.27

-0.5

87.6

POX 23

SPZCOM1

182

442.29

60

19.5

0.06

11.4

97.6

POX 23R

SPZCOM1

120

291.62

60

10.2

0.79

-24.4

55.1

POX 24

SPZCOM1

116

281.90

75

16.2

0.30

7.1

87.3

POX 25

SPZCOM2

0

0.00

45

0.0

2.24

0.9

15.3

POX 26

SPZCOM2

137

332.93

45

0.0

2.30

4.3

16.0

POX 26R

SPZCOM2

187

454.44

45

23.2

0.05

-2.7

98.0

POX 27

SPZCOM2

145

352.38

60

10.3

0.91

-2.6

64.4

POX 28

SPZCOM2

184

447.15

75

8.25

< 0.05

1.5

98.1

POX 29

SPZCOM3

0

0.00

45

25.1

0.35

2.8

84.4

POX 30

SPZCOM3

12

29.16

45

33.1

< 0.05

10.4

97.9

POX 31

SPZCOM3

0

0.00

60

26.3

< 0.05

4.3

97.8

POX 32

SPZCOM3

0

0.00

75

23.3

< 0.05

1.7

97.7

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Table 10-34: Baseline CIL Results

Test

ID

Sample

ID

Calc

Head

Au (g/t)

Calc Head

Au

(oz/st)

Assayed

Residue

Au

(g/t)

Assayed

Residue Au

(oz/st)

Au

Extraction

(%)

NaCN

Consumption

(kg/t)

NaCN

Consumption

(lb/st)

CaO

Consumption

(kg/t)

CaO Consumption

(lb/st)

CIL 1

SPZS

14.42

0.421

6.35

0.185

56.0

1.15

2.79

1.35

3.28

CIL 2

SPZM

7.84

0.229

7.56

0.221

3.6

0.70

1.70

0.95

2.31

CIL 3

SPZN

12.81

0.374

12.15

0.354

5.2

0.67

1.63

0.99

2.41

CIL 4

SPZHG

29.60

0.863

14.15

0.413

52.2

1.31

3.18

1.96

4.76

CIL 5

SPZLG

5.79

0.169

5.2

0.152

10.2

0.96

2.33

0.97

2.36

CIL 6

SPZCOM1

16.70

0.487

13.15

0.384

21.2

1.69

4.11

2.79

6.78

CIL 7

SPZCOM2

16.08

0.469

14.15

0.413

12.0

1.02

2.48

1.19

2.89

CIL 8

SPZCOM3

16.00

0.467

7.28

0.212

54.5

3.09

7.51

9.07

22.04

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Table 10-35: POX Residue CIL Results

Test ID

Sample ID

POX

Test ID

Calc Head Au

(g/t)

Calc

Head Au (oz/st)

Assayed Residue Au

(g/t)

Assayed

Residue Au (oz/st)

Au

Extraction

(%)

NaCN

Consumption

(kg/t)

NaCN

Consumption (lb/st)

CaO

Consumption

(kg/t)

CaO

Consumption (lb/st)

CIL 9

SPZS

1

10.9

0.318

2.44

0.071

77.6

0.92

2.24

0.97

2.36

CIL 10

SPZS

2R

9.53

0.278

0.68

0.020

92.9

1.09

2.65

2.16

5.25

CIL 12

SPZS

2RR

10.2

0.298

1.06

0.031

89.6

0.88

2.14

2.15

5.22

CIL 36

SPZS

3

9.75

0.284

0.64

0.019

93.5

0.75

1.82

3.43

8.34

CIL 37

SPZS

4

9.48

0.277

0.50

0.015

94.8

0.64

1.56

2.33

5.66

CIL 5

SPZM

5

8.19

0.239

4.07

0.119

50.3

1.13

2.75

0.89

2.16

CIL 6

SPZM

6

8.02

0.234

3.90

0.114

51.4

1.24

3.01

0.81

1.97

CIL 15

SPZM

6R

7.54

0.220

1.52

0.044

79.9

1.34

3.26

4.00

9.72

CIL 38

SPZM

7

11.7

0.341

1.45

0.042

87.6

1.65

4.01

6.98

16.96

CIL 39

SPZM

8

7.27

0.212

0.44

0.013

94.0

0.78

1.90

4.61

11.20

CIL 17

SPZN

9

12.3

0.359

5.59

0.163

54.4

1.15

2.79

0.99

2.41

CIL 18

SPZN

10

11.6

0.338

1.66

0.048

85.7

1.45

3.52

1.87

4.54

CIL 26

SPZN

10R

12.0

0.350

1.94

0.057

83.8

1.58

3.84

4.76

11.57

CIL 45

SPZN

11R

11.6

0.338

0.88

0.026

92.4

1.12

2.72

5.49

13.34

CIL 40

SPZN

12

11.4

0.333

2.41

0.070

78.9

0.76

1.85

3.99

9.70

CIL 13

SPZHG

13

23.6

0.688

6.85

0.200

70.9

1.16

2.82

0.91

2.21

CIL 14

SPZHG

14

22.8

0.665

6.85

0.200

69.9

1.13

2.75

0.93

2.26

CIL 16

SPZHG

14R

22.3

0.650

1.51

0.044

93.2

0.68

1.65

2.56

6.22

CIL 41

SPZHG

15

13.1

0.382

0.95

0.028

92.7

1.26

3.06

4.74

11.52

CIL 22

SPZHG

16

18.4

0.537

1.11

0.032

94.0

0.91

2.21

3.31

8.04

CIL 23

SPZLG

17

5.61

0.164

2.22

0.065

60.5

1.03

2.50

0.63

1.53

CIL 25

SPZLG

18

5.42

0.158

0.50

0.015

90.9

0.91

2.21

3.53

8.58

CIL 19

SPZLG

19

5.25

0.153

1.53

0.045

71.0

1.05

2.55

2.58

6.27

CIL 20

SPZLG

20

5.16

0.151

0.28

0.008

94.6

1.05

2.55

3.12

7.58

CIL 29

SPZCOM1

21

15.8

0.461

7.95

0.232

49.7

1.57

3.82

1.67

4.06

CIL 30

SPZCOM1

22

15.5

0.452

8.39

0.245

45.7

0.95

2.31

1.13

2.75

CIL 43

SPZCOM1

22R

15.3

0.446

2.29

0.067

85.1

0.81

1.97

4.29

10.43

CIL 46

SPZCOM1

23R

15.2

0.443

4.36

0.127

71.3

0.84

2.04

5.23

12.71

CIL 24

SPZCOM1

24

16.7

0.487

1.39

0.041

91.7

1.38

3.35

5.27

12.81

CIL 33

SPZCOM2

25

14.8

0.432

6.75

0.197

54.3

1.16

2.82

0.85

2.07

CIL 34

SPZCOM2

26

14.3

0.417

7.98

0.233

44.3

1.02

2.48

0.73

1.77

CIL 44

SPZCOM2

26R

14.4

0.420

3.01

0.088

79.2

0.61

1.48

3.63

8.82

CIL 21

SPZCOM2

27

14.4

0.420

1.19

0.035

91.7

1.39

3.38

3.31

8.04

CIL 28

SPZCOM2

28

16.2

0.473

0.44

0.013

97.3

1.08

2.62

4.48

10.89

CIL 29

SPZCOM3

29

16.8

0.490

0.89

0.026

94.7

1.14

2.77

5.50

13.37

CIL 35

SPZCOM3

30

15.8

0.461

0.49

0.014

96.9

1.01

2.45

3.41

8.29

CIL 31

SPZCOM3

31

16.2

0.473

0.62

0.018

96.2

1.13

2.75

3.64

8.85

CIL 32

SPZCOM3

32

15.1

0.440

0.69

0.020

95.5

1.13

2.75

4.13

10.04

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Source: i-80 Gold

Figure 10-25: South Pacific Gold Recovery as a Function of Sulfide Oxidation (SGS 2026)

Third-Party Toll Milling Results

Gold recovery results from the monthly toll milling bench top autoclave tests, from October 2023 through December 2025, are shown in Table 10-36; recovery averaged 91.1%, ranging from 88.9% to 95.4%.

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Table 10-36: Gold Recovery Results from Toll

Milling BTAC Tests

Lot

Month

Calculated

Head

Grade Au

(g/t)

Calculated Head

Grade Au (oz/st)

Residue

Grade Au

(g/t)

Residue

Grade

Au (oz/

st)

Gold

Recovery

(%)

Oct-2023 (#1)

10.02

0.292

0.89

0.026

90.5

Oct-2023 (#2)

10.13

0.295

1.07

0.031

88.9

Mar-2025

6.78

0.198

0.59

0.017

90.1

Apr-2025

7.10

0.207

0.62

0.018

90.5

May-2025

5.83

0.170

0.57

0.017

89.1

Jun-2025

6.25

0.182

0.65

0.019

89.6

Jul-2025

9.60

0.280

0.86

0.025

90.2

Aug-2025

8.06

0.235

0.48

0.014

93.6

Sep-2025

7.86

0.229

0.34

0.010

95.4

Oct-2025

8.07

0.235

0.55

0.016

93.3

Nov-2025

6.47

0.189

0.50

0.015

91.4

Dec-2025

7.64

0.223

0.65

0.019

90.5

Gold Recovery — Basis for Recovery Estimates

Expected gold recoveries for the Granite Creek Underground Project were developed from the BTAC test data described above, by mining zone and

material type.

Oxide-type material will not require pressure oxidation to achieve acceptable gold recovery and will be processed directly

by CIL. Gold recovery in CIL bottle roll tests on Granite Creek Underground samples, as a function of cyanide-soluble gold percent and of sulfide sulfur content, is shown in Figure 10-26 and Figure 10-27, respectively; at the lower range of sulfide sulfur content, CIL recovery decreases approximately linearly with increasing sulfide sulfur, and this portion of the data was used to model oxide CIL recovery

(only samples below 1% sulfide sulfur were considered for the sulfide sulfur-based function, since samples above 1% sulfide sulfur would not be processed by CIL alone without pressure oxidation). Combining the two regressions as an average yields an

expression for gold recovery by CIL without pressure oxidation as a function of cyanide-soluble gold, fire-assayed gold, and sulfide sulfur content, capped at a maximum recovery of 87.4% (the maximum gold recovery achieved in any of the tests).

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Source: i-80 Gold

Figure 10-26: CIL Recovery vs Cyanide Soluble Gold Percent

Source: i-80 Gold

Figure 10-27: CIL Recovery vs Feed Sulfide Sulfur Content

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Gold recovery by pressure oxidation and CIL at the Lone Tree facility is estimated by mining

zone, based on bench top autoclave tests conducted at FLSmidth and SGS at the expected operating conditions of the Lone Tree facility.

OG

Zone gold recovery is plotted as a function of gold head grade and organic carbon content in Figure 10-28 and Figure 10-29, respectively, for tests at the expected Lone

Tree operating conditions; combining these two regressions as an average yields an expression for gold recovery by CIL after pressure oxidation as a function of gold head grade and organic carbon content.

Source: i-80 Gold

Figure 10-28: OG Zone Gold Recovery vs Gold Head Grade

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Source: i-80 Gold

Figure 10-29: OG Zone Gold Recovery vs Organic Carbon Content

Gold recovery of Otto Zone material was consistent across the four samples tested at expected Lone Tree operating conditions, shown in Table 10-37; the average of these values, 89.3%, was selected as the expected recovery by pressure oxidation-CIL at Lone Tree for the Otto Zone.

Table 10-37: BTAC Gold Recovery Results on Otto Zone Samples

Program

Sample ID

Au Recovery (%)

FLS-2023

OTU

90.0

FLS-2023

OTL

90.3

FLS-2023

OTHG

87.0

FLS-2023

OTLG

90.0

Recovery for the SPZ was estimated using data from tests conducted at expected Lone Tree conditions from the

2025 SGS program. Figure 10-30 through Figure 10-32 show sulfide oxidation versus carbonate content, mass loss in pressure oxidation versus carbonate content, and CIL

gold recovery as a function of CIL feed sulfide sulfur content (after POX), respectively. Samples with higher carbonate content tended to be less oxidized after acid pressure oxidation at Lone Tree conditions, and higher carbonate content resulted

in a mass gain as gypsum formed with sulfuric acid addition; the resulting sulfide sulfur grade of the POX residue controlled downstream CIL gold recovery, with higher residual sulfide values corresponding to lower gold recovery. Combining the

resulting regression lines with head sulfide sulfur content yields an expression for SPZ gold recovery as a function of carbonate and sulfide sulfur content in the feed, capped at 98.0% (the highest recovery achieved in any of the testing at

expected Lone Tree conditions).

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Source: i-80 Gold

Figure 10-30: Sulfide Oxidation vs Carbonate Content,

45-Minute Retention Time, Acidic Conditions

Source: i-80 Gold

Figure 10-31: Mass Loss versus Carbonate Content, 45-Minute

Retention Time, Acidic Conditions

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Source: i-80 Gold

Figure 10-32: CIL Gold Recovery vs CIL Feed Sulfide Sulfur, All Conditions

The toll milling facility performs bench top autoclave testing on each monthly lot of sulfide material shipped to it, as described in

Section 10.1; the average recovery of these tests, 91.1%, has been relatively consistent within a tight range since inception of toll milling operations, and 91.1% is accordingly the expected recovery for material processed by pressure oxidation-CIL at the toll milling facility.

Several processing factors and deleterious elements were

identified across the test programs that affect, or could affect, gold and silver recovery: all FLSmidth 2022 samples demonstrated some degree of preg-robbing (average preg-robbing index 17.9%, ranging from 4.4% to 54.1%), consistent with historical

preg-robbing observed in Mag Pit samples, though CIL improved recoveries over baseline cyanide shake tests by only an average of 4.1%, indicating CIL counteracts preg-robbing only to a small extent; two FLSmidth 2022 samples (APL and APLG) showed

elevated swelling clay content that required attention (trona addition) during POX treatment; acidic POX conditions, while producing higher sulfide oxidation and gold recovery than alkaline conditions, consume substantially more lime and appear to

lock silver within jarosite, which does not form under alkaline conditions, resulting in markedly lower silver recovery under acidic POX; and SGS mineralogical work on the South Pacific Zone found no free arsenopyrite liberation in any sample,

suggesting the zone’s refractory gold is likely associated with arsenic. These factors are accounted for in the zone-specific recovery models described above.

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10.5

Adequacy of Data and Non-Conventional Industry Practice

The metallurgical test programs described in Sections 10.1 through 10.4 — spanning autoclave pre-treatment testing (Dawson, 2005-2006), comprehensive characterization and pressure oxidation testing (FLSmidth, 2022-2023), pilot-scale continuous pressure oxidation testing on a representative blended feed

(FLSmidth, 2025-2026), pressure oxidation and cyanidation testing focused on the South Pacific Zone (SGS, 2025-2026), and ongoing production-scale bench top autoclave testing at the third-party toll milling facility — collectively provide, in

the opinion of the QP, an adequate basis for the gold recovery estimates used in this technical report summary. The test samples span the primary mineralized zones of the Granite Creek Underground deposit (OG, Otto, Adams Peak, Deep Range Front,

Range Front, and South Pacific), and their grade and compositional characteristics (gold, organic carbon, carbonate, and sulfide sulfur) are consistent with the range expected in the projected life-of-mine feed (Section 10.2). Recovery estimates for each zone are based on regression relationships developed from batch and, where available, continuous test data at the expected operating conditions of

the Lone Tree pressure oxidation facility, informed by ongoing production-scale reconciliation data from the toll milling facility currently processing Granite Creek Underground material — providing an actual operating cross-check on the

bench-scale recovery estimates. The pressure oxidation (POX) and carbon-in-leach (CIL) processes tested for the refractory sulfide gold mineralization at Granite Creek

Underground, and the associated comminution, cyanide destruction, and solids-liquids separation test methods, are conventional, industry-standard metallurgical testing procedures for refractory Carlin-style gold ore; the analytical procedures used

across all five test programs are consistent with conventional industry practice, and no non-conventional testing procedures were identified in the source technical documentation reviewed by the QP.

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11

Mineral Resource Estimate

The Mineral Resource for the Granite Creek Underground Project has been developed as two separate estimates, reflecting the deposit’s two

distinct extraction methods. The open pit Mineral Resource was estimated by GRE and remains unchanged from the previously disclosed estimate in terms of underlying drilling, sampling, and technical information; GRE updated only its assessment of

reasonable prospects for economic extraction (RPEE) to reflect the current resource gold price. The underground Mineral Resource was estimated by SRK Consulting using the exploration database supporting the underground operation completed by i-80 Gold Corp. The effective date of both the open pit and underground Mineral Resource estimates is March 31, 2026.

11.1

Open Pit

The Mineral Resource Statement presented herein is updated from the 2025 Technical Report, “Initial Assessment SK-1300 Technical Report, Granite Creek Mine Project, Humbolt County, Nevada, USA,” with an effective date of December 31, 2024, and an issue date of March 26, 2025, with the following new data:

•

Seventeen new drillholes with 2,195 new gold assays have been added to the database.

•

Major faults used for structural trends of gold grade continuity have been updated with input from i-80 Geology team.

•

Economically feasible gold price based on a 36-month trailing average has

been calculated for defining potentially minable resources.

The Mineral Resources were estimated in conformity with the

Society for Mining, Metallurgy and Exploration (SME) best practice guidelines and are reported in accordance with the S-K 1300 Requirements. There is also no certainty that the Inferred Mineral Resources will

be converted to the Measured or Indicated categories through further drilling or into Mineral Reserves, once economic considerations are applied. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no

certainty that all or any part of the mineral resource will be converted into Mineral Reserves. The open pit project presently has no Mineral Reserves. Whittle Pit optimization was applied to the open pit MRE to assess the reasonable prospects for

economic extraction (RPEE) for the resource.

The open pit MRE for the Granite Creek Mine Project was completed by GRE an independent QP as

defined in SK-1300. The effective date of the open pit resource statement is March 31, 2026. In the opinion of GRE, the MRE reported here is a reasonable representation of the Mineral Resources found in

the open pit portion of the Granite Creek Mine Project at the current level of sampling.

11.1.1

Drillhole Database

GRE performed a data validation of the drillhole database prepared by i-80 for the Granite Creek

deposit and determined it to be of suitable accuracy to perform an MRE for the open pit portion of the property. More details regarding the validation of the drillhole database can be found in Section 12. The drill hole data for the Granite

Creek Mine Project was delivered as separate .csv files that contained exploration and production collar locations, drillhole survey orientations, sample intervals with gold assays in ppm, geologic intervals with lithology, alteration type, and

alteration strength. The collar locations are projected in a local grid system, with planar and elevation units in feet. All downhole intervals are captured in feet.

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The complete data set contained assays, collar, and survey data for a total of 2,855

exploration holes (surface, underground, and trench samples) and 695 production holes (surface and underground). Drilling is a mix of RC drilling, diamond drilling, and RC pre-collar with diamond drilling to

final depth. The exploration assay file contains 212,839 gold assays. The production data assay file contains 1,477 gold assays. The drillhole collar locations are shown in Figure 11-1.

Source: GRE, 2021, reviewed and confirmed 2026

Note: This figure is intended to show the relative distribution of surface drillhole collars on topography around the areas of interest. This

figure does not show all collars that have been drilled.

Figure 11-1: Drillhole Used Plan View

on Topography

A number of negative, missing, and blank assay values exist in the drillhole data files provided to GRE by i-80. Missing intervals and values were assumed to be non-mineralized and therefore assigned a value of half of the most common detection limit used to assay the samples.

Negative assay values were replaced according to Table 11-1.

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Table 11-1: Negative Values in Drillhole Database

Non-Positive

Value

Interpretation

Count In Data Set

Action Taken

-0.005

below detection limit of 0.005 opt

6417

replaced with 0.0857 ppm

-0.003

below detection limit of 0.003 ppm

2723

replaced with 0.0015 ppm

-0.9943

below detection limit of 0.029 opt

429

replaced with 0.0857 ppm

-5557

Sample Not Received

3

Omit

-5556

Sample Not Received

80

Omit

-0.0343

Half the detection limit of 0.002opt

926

replaced with 0.0343 ppm

-0.1714

below detection limit for 0.005 opt

52

replaced with 0.0857 ppm

-3394.2842

conversion of -99 opt to ppm

13

replaced with 0.0857 ppm

Source: GRE, 2026

11.1.2

Topography

Topography was provided by i-80 as .dxf files with triangulated surfaces. The files included both as

built surfaces showing dimensions of previously mined pits at their maximum depths, and present topography which includes backfill, pits, dumps, and surrounding topography. The current topographic data was loaded into Leapfrog Geo and used to

constrain the block model. The topographic data provided by i-80 was not rectangular, which is required within Leapfrog to generate models; therefore, GRE extrapolated topographic data around the edges to form

a rectangular surface (Figure 11-2). The extrapolated area, however, is not part of the resource estimate.

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Figure 11-2: Current Topography Used for Open Pit Resource Estimation

11.1.3

Geologic Model

The geologic model used to complete the MRE was developed by Dr. Samari of GRE using grouped majority composites for lithology based on

data provided to GRE as part of the drillhole database. Material below the current topography and above the as-built surface was classified as backfill and assigned an Au ppm grade of zero. Figure 11-3 illustrates the geologic model used in the resource estimation. The model was validated for geologic accuracy and found to be suitable for the purpose of mineral resource estimation by GRE as defined in SK-1300.

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Figure 11-3: Open Pit Geologic Model, Oblique View

11.1.4

Estimation Domains

Estimation zones were created by constraining the assays that surround the existing open pit areas (Figure

11-4). The underground zone to the North of Zone 3 / CX Pit was not considered since this area was estimated in the underground resource section of this technical report. Vertical extents of the estimation

zones range from 1,676 m (5,500 ft) amsl to 762 m (2,500 ft) amsl.

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Figure 11-4: Open Pit Estimation Zones

Table 11-2 summarizes the estimation zone numbers along with the corresponding pits.

Table 11-2: Open Pit Estimation Zone and Pit Name

Estimation Zone

Pit Name

Zone 1

A Pit

Zone 2

Mag Pit

Zone 3

CX Pit

Zone 4

B Pit

Source: GRE, 2021

Numeric indicator models using gold were constructed to better define the high-grade mineralized domains contained within the generalized

estimation domains to fit around the existing pits. The parameters used to define the indicator models are shown in Table 11-3.

Table 11-3: Open Pit Numeric Indicator Model Parameters

Estimation Zone

Indicator

Model

Cut-off (Au

ppm)

ISO Value

Search

Distance

(ft)

Search Distance

(m)

Dynamic Anisotropy

Zone 1

1.0

0.4

200

61.0

CX Fault

Zone 2

1.0

0.3

250

76.2

Mag Fault

Zone 3 CX

0.1

0.4

180

54.9

CX Fault

Zone 3 SOS Dike

0.1

0.4

100

30.5

SOS Dike

Zone 3 SOS XSECTION

0.1

0.3

150

45.7

SOS X Section Fault

Zone 4

1.0

0.3

200

61.0

NA*

*Because Zone 4 did not use dynamic anisotropy, a global trend set to the following parameters was used:

dip 90, dip Azimuth 100, pitch 75, and ellipse ratios max. 200, int. 200, min. 100.

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Source: GRE, 2021

Initial search ellipse orientations were set from examining the spatial orientation of the composites greater than 1 g/t (0.029 oz/st) (Figure 11-5). After initial construction of the high-grade solids using indicator models, it was noted that some of the high-grade numeric models had voids or otherwise poor geometry. When examining the fault structures,

it was noted that the high-grade domain corresponded well with the location and orientation of several fault structures. Dr. Samari of GRE then attempted to add a structural trend using dynamic anisotropy to the numeric model by constructing a

structural trend from the fault meshes. This improved the continuity of the indicator models and helped eliminate voids that were previously present in the indicator models. ISO factor values, which are defined as a specified numeric threshold value

used for indicator Radial Basis Function (RBF) interpolants to generate a volume, for interpolants were based on visually examining the mineralized body and using an iterative process to select a value that produced a reasonable geologic shape that

respected the informing data. If small islands of volume were created off the major trend, they were clipped out.

Source: GRE, 2021, reviewed and confirmed 2026

Note: The opaque red solid represents the high-grade domain. The transparent red solid is the low-grade

domain. Composite grades are shown at a cut-off of 1 ppm.

Figure 11-5: Example of Numeric Indicator High-Grade Trend Analysis, Mag Pit

One resource estimation zone,

Zone 3, was broken into several sub-domains after failing to produce a reasonable high grade numeric indicator model using a single global trend within the Zone 3 domain. It was noted that during the initial

attempt, branching solids formed displaying three distinct trends. Upon further investigation it was found that these trends corresponded to fault structures that crosscut the Zone 3 domain. The identified structural trends are the CX fault, the SOS

Dike, and the SOS X-Section.

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To construct estimation sub-domains, GRE offset the

fault meshes both forwards and backwards to a thickness that contained most of the high-grade intercept (Figure 11-6). Separate numeric estimators were constructed within these domains, and high-grade and low-grade zones were defined within the zones that contained the mineralized trends (Figure 11-7). To later avoid estimation boundary issues during resource estimation, the

volumes on either side of the mineralized domains were separated into sub domains. These include the HW 1, HW 2, HW 3, and FW zones.

Source: GRE, 2021, reviewed and confirmed 2026

Figure 11-6: Open Pit Zone 3 Sub-Domains

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Note: The high-grade zone is shown in green, the low-grade domain is shown in purple.

Figure 11-7: High-Grade and Low-Grade Open Pit Domains in the CX Fault

Domain Analysis

To

check the validity of the high grade and low-grade estimation domains box and whisker plots were constructed, as shown in Figure 11-8. Generally, a good correlation was observed between the high -grade and

low-grade solids and the distribution of the grades contained within them.

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Figure 11-8: Box and Whisker Plot of Open Pit Estimation Domains

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Source: GRE, 2026

Figure 11-9: Histograms of the High-Grade and Low-Grade Distributions in Zones 1 and 2

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Source: GRE, 2026

Figure 11-10: Histograms of the High-Grade and Low-Grade Distributions in Zones 3 and 4

11.1.5

Compositing

Sample data was composited to intervals of equal length as a bias reduction measure to ensure that the samples used in statistical analysis and

estimations were equally weighted. To accomplish compositing, Ms. Lane of GRE first examined the sample length interval histogram to determine the most common assay length (Figure 11-11).

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Source: GRE, 2026

Figure 11-11: Open Pit Sample Interval Length Statistics of Au ppm Assays

Once it was determined that 2 m (5 ft) was the primary assay length, Ms. Lane of GRE evaluated various compositing lengths using 2 m (5

ft) intervals to avoid splitting assays. It was decided that compositing on a 6 m (20 ft) interval represented a significant decrease in the variance of the data while not adversely decreasing the mean of the data set, as shown in Table 11-4. Therefore, Ms. Lane of GRE selected a 26 m (0 ft ) composite interval.

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Table 11-4: Open Pit Compositing Interval Statistics

Statistic

AU_ppm_Assays

Composite Interval

5 ft

10 ft

15 ft

20 ft

25 ft

Count

221,062

264,712

133,121

88,217

66,103

52,752

Length

1,323,435.3

1,323,396.9

1,323,367.4

1,322,037.1

1,321,859.4

1,321,327.9

Mean

0.362

0.362

0.361

0.358

0.357

0.354

SD

2.58

2.49

2.31

2.08

1.95

1.87

CV

7.1

6.9

6.4

5.8

5.5

5.3

Variance

6.6

6.2

5.4

4.3

3.8

3.5

Minimum

0

0

0

0

0

0

Q1

0.0170

0.0171

0.0171

0.0171

0.0171

0.0171

Q2

0.0170

0.0171

0.0274

0.0343

0.0343

0.0343

Q3

0.0860

0.0857

0.0857

0.0857

0.0857

0.0857

Maximum

290.06

290.06

290.06

119.89

94.33

85.08

Source: GRE, 2026

A box plot comparison of the 6 m (20 ft) composited and the uncomposited assays is shown in Figure

11-12. This comparison shows that compositing, while not changing the mean or quartiles, does drastically reduce the maximum value of grades. Table 11-5 shows the open

pit compositing comparison at 6 m (20 ft) intervals.

Source: GRE, 2026

Figure 11-12: Open Pit Compositing Comparison, 6 m (20 ft) Intervals

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Table 11-5: Open Pit Compositing Comparison, 6 m (20 ft) Intervals

Statistic

Composited

Uncomposited

Count

67,143

224,077

Length

1,342,141.4

1,343,991.7

Mean

0.36

0.36

SD

1.95

2.58

CV

5.45

7.11

Variance

3.81

6.65

Minimum

0

0

Q1

0.017

0.017

Q2

0.034

0.017

Q3

0.086

0.086

Maximum

94.329

290.057

Source: GRE, 2026

11.1.6

Evaluation of Outliers

Cumulative probability plots for gold were completed for the composites within each estimation domain (Figure

11-13). A break in the population was identified and marked with the clipping line. Based on this analysis, Ms. Lane of GRE applied a maximum allowable value for the gold grade within each separate

domain, as shown in Table 11-6.

Source: GRE, 2026

Figure 11-13: Example of Open Pit Cumulative Log Probability Plot, Zone 1 HG

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Table 11-6: Open Pit Upper Clipping (Capping) Au ppm Values by Domain

Zone

Sub-Domain

Clipping Value

Zone 1

HG

35

LG

NA

Zone 2

HG

10

LG

5

Zone 3

CX HG

20

CX LG

10

SOS DIKE HG

NA

SOS DIKE LG

NA

SOS XSECTION HG

NA

SOS XSECTION LG

NA

HW 1

7

HW 2

3

HW 3

7

FW

12

Zone 4

HG

NA

LG

3

Source: GRE, 2026

11.1.7

Density

Density was assigned to each domain in the block model based on a combination of rock type and grade, as shown in Table 11-7. The bulk densities are the same as those used in the 2020 Getchell Project Technical Report (AMC, 2020) and were originally supplied by OMC. The results for each domain fit well with GRE’s experience

with similar rock types.

Table 11-7: Open Pit Domain Density Summary

Unit

Au>=0.008 opt

(tonne/m3)

Au<0.008 opt

(tonne/m3)

Backfill

1.85

1.85

Alluvium

1.85

1.85

Granodiorite

2.7

2.7

Upper Comus

2.5

2.7

Lower Comus

2.51

2.64

Preble

2.42

2.6

Source: GRE, 2026

11.1.8

Variography

After iterative analysis, a good fit for the gold grade variography was found using pairwise relative variograms. The pairwise relative

variogram helps to smooth the variogram by scaling g(h) using the square of the mean of each sample pair of the data from calculating g(h). This helps make the

interpretation of the variogram model easier, and all variances calculated this way are relative to the mean of the sample pairs within the distribution.

Variogram analysis was completed on the samples within each of the high-grade and low-grade estimation domains to establish the direction of

maximum continuity between sample pairs. The range for each variogram was found using a global variogram. The nugget was calculated by examining the downhole variograms and determining where the short-range trend crossed the y-axis. Variograms were

oriented along the strike and dip of the visually observed high-grade trend of the composites, with the major axis oriented along the direction of maximum continuity (Table 11-8).

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Table 11-8: Open Pit Variogram Parameters

Zone

Sub-Domain

Dip

Dip Azimuth

Pitch

Major Axis

(ft)

Major Axis

(m)

Semi-Major

Axis (ft)

Semi-Major

Axis (m)

Minor Axis

(ft)

Minor Axis

(m)

Zone 1

Overall

38

134

75

200

61.0

200

61.0

100

30.5

HG

38

134

160

160

48.8

160

48.8

75

22.9

LG

38

134

75

175

53.3

175

53.3

60

18.3

Zone 2

Overall

50

70

105

250

76.2

250

76.2

100

30.5

HG

50

70

75

180

54.9

125

38.1

75

22.9

LG

50

70

105

300

91.4

200

61.0

200

61.0

Zone 3

Overall*

56

135

75

160

48.8

160

48.8

125

38.1

CX

56

135

75

180

54.9

160

48.8

70

21.3

CX HG

56

135

75

80

24.4

80

24.4

50

15.2

CX LG

56

135

75

50

15.2

50

15.2

50

15.2

SOS Dike

65

170

45

100

30.5

80

24.4

50

15.2

SOS DIKE HG

65

170

45

100

30.5

80

24.4

75

22.9

SOS DIKE LG

65

170

45

100

30.5

50

15.2

25

7.6

SOS XSECTION

80

170

80

150

45.7

125

38.1

50

15.2

SOS XSECTION HG

80

170

80

80

24.4

80

24.4

80

24.4

SOS XSECTION LG

80

170

105

150

45.7

125

38.1

50

15.2

HW 1

56

135

150

160

48.8

100

30.5

125

38.1

HW 2

56

135

100

160

48.8

160

48.8

125

38.1

HW 3

56

135

75

160

48.8

160

48.8

125

38.1

FW*

56

135

70

140

42.7

125

38.1

125

38.1

Zone 4

Overall

90

100

75

200

61.0

200

61.0

100

30.5

HG

90

100

90

60

18.3

60

18.3

25

7.6

LG

90

100

75

300

91.4

225

68.6

75

22.9

Source: GRE, 2026

11.1.9

Block Model Parameters

A 3D block model was developed to represent the open pit deposit using a block size of 25 ft x 25 ft x 20 ft. The block model dimensions and

model limits are shown in Table 11-9. The coordinate system used for the 3D modeling was based on the local grid system using imperial units of feet. The block model is un-rotated and contains no sub blocking.

Table 11-9: Open Pit Block Model Parameters

Parameter

Value

Base point

6000,7000,5800 (X,Y,Z)

Parent block size

25x25x20 (X,Y,Z)

Azimuth Rotation

0

Boundary size

9000,7000,3000 (X,Y,Z)

Size in Blocks

360x280x150 (X,Y,Z)

Sub-blocking

None

Source: GRE, 2026

11.1.10

Estimation Domains

The estimation domains used to constrain the MRE resulted from the numeric indicator models developed as part of the geologic model as

discussed in Section 14.2.4.

Figure 11-14 shows an overview of the estimation domains that

were used to constrain the mineral resource estimation for the open pit MRE.

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Figure 11-14: Open Pit Numeric Indicator Models

11.1.11

Estimation Parameters

Estimation within mineralized domain boundaries was performed using an inverse distance squared method with a minimum of 4 samples, a maximum

of 20 samples, and a drillhole limit of 2. Declustering objects were applied to all high-grade estimation domains. Dynamic anisotropy was applied where it was applicable based on faults that structurally control mineralization. The exception to this

was Zone 4, which has no apparent structural control that has yet been mapped. In this case, search ellipse orientation was determined from examining the spatial orientation of the composites greater than 1 g/t. (0.029 oz/st)

Search distances of the domained estimators were based on the variography for each sub-domain, as discussed in Section 14.1.7. Search

distances for domains that showed poor variography were replaced by the overall sub-domain or overall domain search distances. This method was used for the Zone 4 HG and LG estimations and the Zone 3 sub-domains. All estimations used hard boundaries

except for the boundaries between Zone 1 HG and Zone 3 HG and between Zone 1 LG and Zone 4 LG. Soft 40-foot boundaries were set up with filters between these boundaries since they are immediately adjacent to each other and could potentially have

continuity in grade estimation across these boundaries. The inverse distance estimation parameters for each domain are given in Table 11-10.

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Table 11-10: Open Pit ID2 Estimation Parameters

Zone

Sub-Domain

Dynamic

Anisotropy

Trend

Major

(ft)

Major

(m)

Semi-Major

(ft)

Semi-Major

(m)

Minor

(ft)

Minor

(m)

Zone 1

HG

Yes

CX Fault

160

48.8

160

48.8

75

22.9

LG

Yes

CX Fault

175

53.3

175

53.3

60

18.3

Zone 2

HG

Yes

Mag Fault

180

54.9

125

38.1

75

22.9

LG

Yes

Mag Fault

300

91.4

200

61.0

200

61.0

Zone 3

CX HG

Yes

CX Fault

180

54.9

160

48.8

70

21.3

CX LG

Yes

CX Fault

180

54.9

160

48.8

70

21.3

SOS DIKE HG

Yes

SOS Dike

100

30.5

80

24.4

50

15.2

SOS DIKE LG

Yes

SOS Dike

100

30.5

80

24.4

50

15.2

SOS XSECTION HG

Yes

SOS Xsection

150

45.7

125

38.1

50

15.2

SOS XSECTION LG

Yes

SOS Xsection

150

45.7

125

38.1

50

15.2

HW 1

Yes

CX Fault

160

48.8

100

30.5

125

38.1

HW 2

Yes

CX Fault

160

48.8

160

48.8

125

38.1

HW 3

Yes

CX Fault

160

48.8

160

48.8

125

38.1

FW

Yes

CX Fault

160

48.8

160

48.8

125

38.1

Zone 4

HG

No

90,100,90

200

61.0

200

61.0

100

30.5

LG

No

90,100,75

200

61.0

200

61.0

100

30.5

Source: GRE, 2026

After each domained estimator was constructed, a combined estimator was used to assign a hierarchical value to each domained estimation to

produce a single gold grade value. The combined estimator hierarchy is shown in Table 11-11.

Table 11-11: Open Pit Combined Estimator Hierarchy

Priority

Domained Estimation

Priority

Domained Estimation

1

Zone 1 HG

11

Zone 3 HW 3

2

Zone 1 LG

12

Zone 3 HW 1

3

Zone 2 HG

13

Zone 3 HW2

4

Zone 2 LG

14

Zone 3 FW

5

Zone 3 CX HG

15

Zone 4 HG

6

Zone 3 CX LG

16

Zone 4 LG

7

SOS Dike HG

17

Zone 1

8

SOS Dike LG

18

Zone 2

9

SOS Xsection HG

19

Zone 3

10

SOS Xsection LG

20

Zone 4

Source: GRE, 2026

The overall domain grade estimations were assigned to priority 17 to 20 so that they would fill in areas of the numeric estimator domains that

lacked the required number of samples to be able to estimate grade due to the small volume being estimated that excluded drillholes.

11.1.12

Geometallurgical Modeling

Section 13.5 goes into detail of how the gold recovery model was estimated and implemented in the block model.

Cyanide solubility was compared to all available interval information from the drilling data: gold assay, alteration, lithology, depth, etc.

From this available data, a principal component analysis, regression tree, and multivariate adaptive regression spline analysis were performed. A multivariate adaptive regression spline analysis model was created to predict cyanide solubility in

different zones using the available drilling data. Heap Leach (HLCL) recovery was determined by plotting the cyanide solubility with the column recovery. The carbon in leach (CIL) recovery equation was determined by plotting the trend with the

calculated head grade and CIL recovery.

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The input fields required in the recovery equations were added to the block model. Then, the

recovery equations were applied to the block model for HLCH and CIL recoveries. These recoveries, along with the Whittle inputs from Table 14 5, were used to determine which of the two processes would be applied to each block.

11.1.13

Block Model Validation

Validation of the estimated block grades for the open pit MRE were completed for each of the estimation domains. The resource block model

estimate was validated by:

•

Completing a series of visual inspections by comparisons of gold assay and composite grades to estimated block

values across the deposit in both horizontal and vertical sections.

•

Statistical comparison of parameters such as means, quantiles, and variance between 6 m (20 ft) composites,

Nearest Neighbor (NN), Inverse Distance squared (ID2), and Ordinary Kriged (OK) estimators to ensure that the grade estimations are representative of the composites they are based on.

•

Comparing average composite sample values with average estimated block grades along east, north, and elevation

orientations using swath grade trend plots.

Visual Inspection

The block model was examined in plan and section views to compare to drillhole locations and grades. Plan views and section views for each of

the estimation areas are shown in Figure 11-15 through Figure 11-22. Comparison of the block model grade from the assays did not reveal any major discrepancies. In the

section views, blocks above current topography have been removed.

Source: GRE, 2026

Figure 11-15: Open Pit Zone 1 Visual Comparison Composite to Block Model Grade, Plan View

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Figure 11-16: Open Pit Zone 2 Visual Comparison Composite to Block Model Grade, Plan View

Source: GRE, 2026

Figure 11-17: Open Pit Zone 3 Visual Comparison Composite to Block Model Grade Plan View

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Figure 11-18: Open Pit Zone 4 Visual Comparison Composite to Block Model Grade Plan View

Source: GRE, 2026

Figure 11-19: Open Pit Zone 1 Section Composites and Block Model Cross Sections

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Source: GRE, 2026

Figure 11-20: Open Pit Zone 2 Section Composites and Block Model Cross Section

Source: GRE, 2026

Figure 11-21: Open Pit Zone 3 Section Composites and Block Model Cross Section

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Figure 11-22: Open Pit Zone 4 Section Composites and Block Model Cross Section

Statistical Comparison

To ensure

that the grade estimations are representative of the composites they are based on and validate the resource estimation results, the block model grade estimation statistics were analyzed. Ms. Lane of GRE compared the means, quantiles, and

variance between 20-foot composites, Nearest Neighbor (NN), Inverse Distance squared (ID2), and Ordinary Kriged (OK) estimators, as shown in Table 11-12. Blocks are confined to the 3000 US$/tr oz Whittle pit

at a cutoff grade of zero.

Table 11-12: Open Pit Comparison of Composite Values to Grade Estimation Methods

Parameter

Composites

Parameter

NN

ID2

OK

Count

14,380

Block Count

304,317

256,088

256,088

Mean

0.49

Mean

0.28

0.33

0.34

SD

1.86

SD

1.08

0.80

0.78

CV

3.81

CV

3.80

2.41

2.29

Variance

3.47

Variance

1.17

0.65

0.61

Minimum

0

Minimum

0

0.0

-0.5

Q1

0.02

Q1

0.02

0.03

0.03

Q2

0.04

Q2

0.04

0.08

0.08

Q3

0.18

Q3

0.09

0.25

0.26

Maximum

78.71

Maximum

35.00

32.19

27.27

Source: GRE, 2026

As expected, the NN estimator generates more blocks due to the lack of restrictions of having to use multiple samples and multiple drillholes

to estimate a block grade. Both ID2 and OK produced similar quantiles, means, variance, etc. The one marked difference seen between ID2 and OK is that is possible to have a negative value in Kriging as seen in the minimum value of the OK estimator

(Figure 11-23).

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Source: GRE, 2026

Figure 11-23: Cumulative Frequency of Composite and Estimated Block Grades

Swath Plots

Swath

plots of the various estimation methods (NN, ID2, and OK) were used to compare the results from each estimation method to the informing composite values and examine which method smoothed the estimated grades. As an example, swatch plots are provided

below for the Zone 1 high-grade and low -grade domains (Figure 11-24 though Figure 11-26). The swatch plots show a general trend that the ID2 estimator smoothed out

drastic swings in grade while not over smoothing local variability.

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Source: GRE, 2026

Figure 11-24: X Axis, Zone 1 High-Grade Domain

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Source: GRE, 2026

Figure 11-25: Open Pit Swath Plot, Y Axis, Zone 1 High-Grade Domain X axis, Zone 1 High-Grade Domain

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Source: GRE, 2026

Figure 11-26: Open Pit Swath Plot Z Axis, Zone 1 High-Grade Domain

11.1.14

Mineral Resource Classification

Block model quantities and grade estimates for the Granite Creek open pit deposit were classified according to the Society for Mining,

Metallurgy and Exploration (SME) guidelines.

Mineral Resources were classified as Measured, Indicated, or Inferred.

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical

characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit. Geological evidence is derived from detailed and

reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation (SME, 2017). Due to the density of drill hole spacing around the historical pit and underground

excavations, GRE found it reasonable to classify portions of the resource as Measured.

Mineral resource classification involved a two-step

process using minimum distances and minimum numbers of samples to define resource classification initially before applying numeric indicator model to define a more continuous and reasonable resource classification. The criteria used in the first

step of the resource classification are listed in Table 11-13. Parameters used for the numeric indicator models for the second step of the resource classifications are listed in Table 11-14.

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Table 11-13: Open Pit Mineral Resource Classification Parameters

Resource Class

Minimum Distance (ft)

Minimum Distance (m)

Minimum Number of Samples

Measured

50

15.2

7

Indicated

100

30.5

5

Inferred

150

45.7

NA

Source: GRE, 2026

Table 11-14: Open Pit Parameters for Resource Class Numeric Indicator Model

Resource Class

Shape

Interpolant Distance (ft)

Interpolant Distance (m)

ISO

Measured

Isotropic

250

76.2

0.4

Indicated

Isotropic

250

76.2

0.4

Inferred

NA

NA

NA

Source: GRE, 2026

Because the classification was performed across all resource estimation domains, an isotropic search was used along with an interpolant

distance of 250, which was based on the average continuity of grade seen in the deposit. No numeric indicator model was constructed for the Inferred resource class, rather it was defined as any block with a calculated gold grade that did not fall

within the Measured or indicated numeric indicator domains that had a calculated gold grade. A plan view of the estimated resource classes is shown in Figure 11-27.

Source: GRE, 2026

Figure 11-27: Open Pit Constrained Resource Class All Areas Plan View

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11.1.15

Open Pit Mineral Resource Statement

Society for Mining, Metallurgy and Exploration (2017) defines a mineral resource as: “a concentration or occurrence of solid

material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. The location, quantity, grade or quality, continuity, and other geological

characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.” The Mineral Resources may be impacted by further infill and exploration drilling that may result in

increases or decreases in future resource evaluations. The Mineral Resources may also be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, socio-economic, and other factors. Mineral Resources are not

Mineral Reserves and do not have demonstrated economic viability. Mineral Reserves can only be estimated based on the results of an economic evaluation as part of a Preliminary Feasibility Study or Feasibility Study. As a result, no Mineral Reserves

have been estimated as part of this study. There is no certainty that all or any part of the Mineral Resources will be converted into a mineral reserve.

The requirement, “reasonable prospects for economic extraction,” generally implies that the quantity and grade estimates meet

certain economic thresholds and that the Mineral Resources are reported at a cut-off grade considering appropriate extraction scenarios and processing recoveries. To meet this requirement, Ms. Lane of GRE considered that major portions of the

Granite Creek deposit are amenable for open pit extraction.

To determine the quantities of material offering “reasonable prospects

for economic extraction” by an open pit, Ms. Lane of GRE constructed open pit scenarios developed from the resource block model estimate using Whittle’s Lerchs-Grossman miner “Pit Optimizer” software. Reasonable mining

assumptions were applied to evaluate the portions of the block model (Measured, Indicated, and Inferred blocks) that could be “reasonably expected” to be mined from an open pit. The optimization parameters presented in Table 11-15 were selected based on experience and benchmarking against similar projects. The results are used as a guide to assist in the preparation of a mineral resource statement and to select an appropriate resource

reporting cut-off grade. Ms. Lane of GRE considers that the blocks located within the resulting conceptual pit envelope show “reasonable prospects for economic extraction” and can be reported as a mineral resource.

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Table 11-15: Granite Creek Resource Parameters for Open Pit Optimization

Parameter

Items

Unit

Value

Costs

Mining Cost (waste/mineralized material)

US$/tonne mined

2.46

Heap Leach*

US$/tonne mineralized material treated

9.04

Carbon in Leach**

US$/tonne mineralized material treated

17.22

Recovery

Heap Leach (HLCH) Recovery with CN

Solubility <60

%

CN Solubility*100

HLCH Recovery with CN Solubility >= 60

%

((0.1225 * [Au_ppm]) + 0.4164)*100

CIL Recovery

%

((0.5388 * CN Solubility) + 0.3201)*100

Net revenue gold

Gold price

US$/oz

3,000

Selling costs and penalties***

US$/oz

114

Royalty

Total royalty (simplified)

%

6.00%

Slope angles

Slope Angle

degrees

41

Limits

HLCH

tonnes per year

2,975,000

CIL

tonnes per year

1,050,000

Source: GRE, 2026

1

HLCH and CIL costs include US$1.56/tonne milled for admin costs.

2

Various royalties are applicable at various points throughout the mine life, however for the scope of this

Technical Report, GRE has used a single 6% royalty for the open pit mineral resource.

3

The gold price used for this analysis is the 36-month trailing average gold price as of June 30, 2026.

4

This selling cost is used to apply the 6% royalty

The pit optimization used a royalty of 6%; however, the current royalty is approximately 7.5%. The GRE QP does not believe that the change in

royalty results in a material impact on the pit optimization.

Due to the large ratio of deposit size to block size and method of grade

estimation, the grade model is fully diluted, and the resource is 100% recoverable as estimated.

The gold price selected for the open pit

Mineral Resource has been based on review of the market study presented in Section 16 of this report, which has been compared to market consensus forecast data, review by the GRE QP of industry peers, plus discussion with i-80 technical teams. The

resource gold price of US$3,000/oz reflects the 36-month trailing average as of June 30, 2026. Based on review, it is the GRE QP’s opinion that this price is reasonable over a time period of 10 years, covering both short-term pricing and

future long-term pricing. The point of reference for the reporting of open pit Mineral Resources is the tonnage and grades contained within the Whittle pit shell limits, reported at a 0.20 g/t cut-off grade using variable metallurgical recovery.

The Granite Creek open pit mineral resource constrained by a Whittle pit shell that corresponds to a gold price of US$3,000 per troy ounce

is shown in Table 11-16. The reader is cautioned that the results from the pit optimization are used solely for testing the “reasonable prospects for economic extraction” by an open pit and do not

represent an attempt to estimate Mineral Reserves.

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Table 11-16: Granite Creek Open Pit Mineral Resource Statement

Deposit

Cutoff

Grade

(ppm)

Mass

(‘000s tonnes)

Mass

(‘000s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(‘000s tr oz)

Measured

Pit B

0.2

3,900

4,299

1.06

0.031

133.49

Pit A

0.2

1,098

1,211

0.71

0.021

25.08

CX

0.2

13,493

14,873

1.10

0.032

478.41

Mag

0.2

15,055

16,596

1.03

0.030

499.50

Total

33,546

36,978

1.05

0.031

1,136.47

Indicated

Pit B

0.2

823

907

0.64

0.019

16.86

Pit A

0.2

969

1,069

0.68

0.020

21.06

CX

0.2

4,423

4,876

1.00

0.029

141.96

Mag

0.2

11,509

12,687

0.74

0.022

275.03

Total

17,725

19,538

0.80

0.023

454.91

Measured + Indicated

Pit B

0.2

4,723

5,206

0.99

0.029

150.34

Pit A

0.2

2,068

2,280

0.69

0.020

46.13

CX

0.2

17,916

19,749

1.08

0.031

620.37

Mag

0.2

26,565

29,282

0.91

0.026

774.53

Total

51,271

56,517

0.97

0.028

1,591.38

Inferred

Pit B

0.2

50

55

0.52

0.015

0.84

Pit A

0.2

440

485

0.44

0.013

6.22

CX

0.2

1,997

2,201

0.94

0.028

60.66

Mag

0.2

1,339

1,476

0.75

0.022

32.12

Total

3,826

4,217

0.81

0.024

99.84

Source: GRE, 2026

1) The effective date of the Mineral Resources Estimate is March 31, 2026

2) The Qualified Persons for the estimate are GRE.

3) Mineral Resources are not ore reserves and are not demonstrably economically recoverable.

4) Mineral Resources are reported at a 0.20 g/t (0.006 oz/st) cutoff, an assumed gold price of 3,000 US$/tr. oz, using variable recovery, a

slope angle of 41 degrees, 6% royalty, heap leach processing cost US$9.04 per tonne (includes admin), CIL processing cost of US$17.22 per tonne (includes admin).

11.1.16

Mineral Resource Sensitivity by Domain

Table 11-17 shows the sensitivity of the mineral resource to cut-off grade in each domain.

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Table 11-17: Granite Creek Mineral Resource

Sensitivity to Cut-off Grade – Reported on 100% basis

Deposit

Cutoff Grade

(ppm)

Mass

(000’s tonnes)

Mass

(000’s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(000’s tr oz)

Measured

Pit B

0.1

5,131

5,656

0.84

0.025

139.202

0.15

4,399

4,849

0.96

0.028

136.288

0.2

3,900

4,299

1.06

0.031

133.487

0.25

3,501

3,859

1.16

0.034

130.620

0.3

3,153

3,476

1.26

0.037

127.537

0.35

2,856

3,148

1.36

0.040

124.449

0.4

2,629

2,897

1.44

0.042

121.702

0.45

2,391

2,635

1.54

0.045

118.446

0.5

2,187

2,410

1.64

0.048

115.323

Pit A

0.1

1,489

1,641

0.56

0.016

26.932

0.15

1,270

1,400

0.64

0.019

26.038

0.2

1,098

1,211

0.71

0.021

25.078

0.25

915

1,009

0.81

0.024

23.756

0.3

773

852

0.91

0.026

22.507

0.35

638

703

1.03

0.030

21.103

0.4

526

580

1.17

0.034

19.759

0.45

440

485

1.31

0.038

18.590

0.5

370

408

1.47

0.043

17.520

CX

0.1

18,417

20,301

0.85

0.025

500.907

0.15

15,362

16,933

0.99

0.029

488.857

0.2

13,493

14,873

1.10

0.032

478.406

0.25

12,117

13,356

1.20

0.035

468.522

0.3

11,130

12,268

1.29

0.037

459.863

0.35

10,282

11,334

1.36

0.040

451.026

0.4

9,577

10,557

1.44

0.042

442.531

0.45

8,944

9,859

1.51

0.044

433.891

0.5

8,344

9,198

1.58

0.046

424.760

Mag

0.1

18,446

20,333

0.87

0.025

515.371

0.15

16,524

18,214

0.96

0.028

507.759

0.2

15,055

16,596

1.03

0.030

499.500

0.25

13,683

15,083

1.11

0.032

489.626

0.3

12,603

13,893

1.18

0.035

480.080

0.35

11,672

12,866

1.25

0.037

470.359

0.4

10,721

11,818

1.33

0.039

458.902

0.45

9,852

10,860

1.41

0.041

447.027

0.5

9,114

10,046

1.49

0.043

435.771

Indicated

Pit B

0.1

1,491

1,643

0.42

0.012

19.963

0.15

1,098

1,210

0.52

0.015

18.406

0.2

823

907

0.64

0.019

16.857

0.25

587

647

0.80

0.023

15.168

0.3

433

477

0.99

0.029

13.825

0.35

347

382

1.16

0.034

12.925

0.4

279

307

1.35

0.039

12.113

0.45

243

268

1.49

0.043

11.635

0.5

207

229

1.66

0.049

11.090

Pit A

0.1

1,290

1,422

0.54

0.016

22.538

0.15

1,114

1,228

0.61

0.018

21.854

0.2

969

1,069

0.68

0.020

21.057

0.25

863

952

0.73

0.021

20.288

0.3

782

862

0.78

0.023

19.575

0.35

694

765

0.84

0.024

18.655

0.4

605

667

0.90

0.026

17.579

0.45

524

578

0.98

0.029

16.466

0.5

463

510

1.04

0.030

15.531

CX

0.1

6,949

7,660

0.69

0.020

153.509

0.15

5,406

5,959

0.85

0.025

147.421

0.2

4,423

4,876

1.00

0.029

141.964

0.25

3,755

4,139

1.14

0.033

137.167

0.3

3,364

3,708

1.24

0.036

133.734

0.35

3,102

3,419

1.31

0.038

131.006

0.4

2,875

3,169

1.39

0.040

128.266

0.45

2,652

2,924

1.47

0.043

125.225

0.5

2,482

2,736

1.54

0.045

122.621

Mag

0.1

15,095

16,640

0.60

0.018

291.797

0.15

13,067

14,403

0.68

0.020

283.755

0.2

11,509

12,687

0.74

0.022

275.034

0.25

10,179

11,221

0.81

0.024

265.446

0.3

9,029

9,953

0.88

0.026

255.304

0.35

8,018

8,838

0.95

0.028

244.767

0.4

7,160

7,893

1.02

0.030

234.421

0.45

6,245

6,884

1.11

0.032

221.921

0.5

5,416

5,970

1.20

0.035

209.258

Inferred

Pit B

0.1

72

79

0.40

0.012

0.919

0.15

52

57

0.51

0.015

0.848

0.2

50

55

0.52

0.015

0.837

0.25

45

50

0.55

0.016

0.804

0.3

39

43

0.60

0.017

0.746

0.35

30

33

0.67

0.020

0.655

0.4

24

27

0.75

0.022

0.582

0.45

23

26

0.76

0.022

0.569

0.5

21

23

0.80

0.023

0.528

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Deposit

Cutoff Grade

(ppm)

Mass

(000’s tonnes)

Mass

(000’s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(000’s tr oz)

Pit A

0.1

615

678

0.35

0.010

7.010

0.15

506

558

0.40

0.012

6.583

0.2

440

485

0.44

0.013

6.216

0.25

352

388

0.49

0.014

5.577

0.3

275

304

0.55

0.016

4.910

0.35

229

253

0.60

0.018

4.432

0.4

188

207

0.65

0.019

3.936

0.45

154

170

0.70

0.020

3.480

0.5

124

137

0.76

0.022

3.018

CX

0.1

3,155

3,478

0.65

0.019

65.995

0.15

2,464

2,716

0.80

0.023

63.252

0.2

1,997

2,201

0.94

0.028

60.660

0.25

1,734

1,912

1.05

0.031

58.790

0.3

1,574

1,735

1.13

0.033

57.387

0.35

1,426

1,572

1.22

0.036

55.846

0.4

1,316

1,451

1.29

0.038

54.524

0.45

1,232

1,358

1.35

0.039

53.370

0.5

1,146

1,263

1.41

0.041

52.064

Mag

0.1

1,901

2,095

0.57

0.017

34.750

0.15

1,594

1,757

0.65

0.019

33.550

0.2

1,339

1,476

0.75

0.022

32.123

0.25

1,036

1,142

0.90

0.026

29.916

0.3

783

863

1.10

0.032

27.723

0.35

699

770

1.20

0.035

26.851

0.4

635

700

1.28

0.037

26.090

0.45

612

675

1.31

0.038

25.783

0.5

602

664

1.32

0.039

25.631

Source: GRE, 2026

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11.1.17

Factors that Could Affect Open Pit Mineral Resources

To the best of the QP’s knowledge, there are no known legal, political, environmental, permitting, title, taxation, socio-economic,

marketing, mining, metallurgical, or other factors that would further materially affect the open pit Mineral Resources reported herein.

There are no known significant factors or risks that may affect property access, title, or the right to perform work on the Property.

The open pit MRE could be materially affected negatively by low market prices for gold and by difficulties in material handling and processing

that would affect the recovery and production of gold. Changes in the estimated materials and supply costs, and in labor availability and rates are other factors that could materially affect the open pit MRE. The taxation and political environment

for mining in Nevada is relatively stable. The Project requires infrastructure development and permitting.

11.2

Underground Mineral Resources

SRK completed the MRE process using updated mineralization models. i-80 provided SRK with a comprehensive exploration database with logging

indicating the main geological features and units, including full assays from exploration and production. In addition to the database, SRK worked with i-80 on preliminary geological interpretations, on which, SRK made minor alterations accordingly.

The resource estimation methodology involved the following procedures:

•

Database compilation and verification

•

Construction of wireframe models for the fault networks and centerlines of mining development per vein

•

Definition of resource domains

•

Data conditioning (compositing and capping) for statistical and geostatistical analysis

•

Variography

•

Block modeling and grade interpolation

•

Resource validation and classification

•

Application of reporting CoG using the 2026 inputs

•

Preparation of the mineral resource statement

11.2.1

Drillhole Database

i-80 supplied SRK with extracts from the Central database (acQuire) in both csv and Leapfrog formats. The cut-off for the database was

February 17, 2026. The digital database includes both holes inside the area of interest and the regional drilling over the Granite Creek deposit. The database contains a total of 7,977 boreholes and production samples for 521,332.6 m

(1,710,408.7 ft) of drilling. SRK has limited the analysis for the study to the drilling in the near vicinity to the current underground operations and therefore has applied a spatial limitation of Easting between 8,200 to 12,800, and northing 9,000

to 15,000 in the local grid. The export reduced the total database to 6,800 holes which is a combination of the exploration, production and environmental drillhole data (Table 11-18).

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Table 11-18: Summary of Drilling and Sampling in Target Area

Drilling Type

Count

Sum of Depth (ft)

Sum of Depth (m)

Environmental

22

18,227.0

5,555.6

Exploration

1,669

905,284.2

275,930.6

Core

159

125,636.1

38,293.9

CV

111

29,903.0

9,114.4

CVM

1

1,000.0

304.8

DDH

131

55,777.7

17,001.0

Metallurgical

3

1,535.0

467.9

RC

1,075

456,390.0

139,107.7

RC/CORE

65

92,353.0

28,149.2

RC_CVM

1

1,940.0

591.3

RC_DDT

121

140,084.4

42,697.7

UNK

2

665.0

202.7

Production

5,109

243,178.1

74,120.7

Channel

544

48,624.3

14,820.7

Core

199

67,446.8

20,557.8

Cubex

641

89,857.0

27,388.4

Sludge

3,725

37,250.0

11,353.8

Grand Total

6,800

1,166,689.3

355,606.9

Source: SRK, 2026

The subset databased in Leapfrog used for the estimation is split into a series of tables taken from the central acQuire database is summarized

in Table 11-19, using the above defined spatial restrictions (whereby any holes which have intersections with the geographical area have been included in the export). The assay database contains the complete

multi-element data analysis plus basis for the geo-metallurgical analysis including PregRob (%), Total Organic Carbon (TOC %), Carbonate (CO3 %), and cyanide soluble gold Au_CN (ppm), which are used to aid in the gold recovery calculations.

Table 11-19: Summary of Drilling and Sampling Information Available within Leapfrog

Data Table

Data Type

Number of Columns

Number of Rows

collar

collar

19

6,800

survey

survey

8

35,633

Assay_Combined

interval

68

189,851

alteration

interval

22

54,079

density

interval

20

8,431

geotech

interval

11

60,593

lithology

interval

33

58,466

oxides

interval

20

47,247

structures

interval

12

21,935

sulfides

interval

20

45,239

veins

interval

14

14,175

point_structures

point

16

8,924

Source: SRK, 2026

During the review SRK noted several missing sampling intervals (60), which were reset to zero, and a total of 1,980 missing assays which were

omitted from the analysis. In addition to the absent values, SRK notes there are 10,774 occurrences of zero in the database which have been kept as they define areas which were not sampled and assumed to be waste by the geologist. Figure 11-28 shows the collar locations for all drillholes in the drillhole database and are colored coded by company drilled.

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Figure 11-28: Plan Showing Collar Locations and Borehole Traces within the MRE-Focused Dataset

11.2.2

Topography and Mining Depletion

SRK has been supplied with a series of wireframes from the i-80 Geology team to reflect the current topography and the underground mining

depletion at the Granite Creek deposit. The depletion was based on a combination of both open pit mining and month-end depletion from the underground mining operation as of March 31, 2026. Separate models were produced by SRK for the open pit

and underground models to account for these separately.

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To create the depletion, SRK has imported the information provided by i-80 in the .dxf

formats supplied. The three key files used for the depletion are:

•

Regional Topo – Pre Mine.dxf

•

Mine Topo – Backfilled.dxf

•

GC_Merged_2603_Clean.dxf

SRK created a depletion model using the same prototype as the geological and mineralization models. The open pit depletion has been treated as

an erosional surface in Leapfrog®. SRK generated an “intrusion” model in Leapfrog® to account for the underground depletion

and sterilization around the workings. Overall, SRK considers the depletion model to accurately reflect the information provided. There is a low risk that these depletions do not accurately reflect the final pit or stopes, but SRK would not consider

that material and considers the current process as reasonable for the definition of the current Mineral Resources. An isometric view of the depletion model is seen below in Figure 11-29.

Source: SRK, 2026

Figure 11-29: Isometric View Showing Open Pit Depletion (Surface) and Underground Workings (Grey)

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11.2.3

Geological Model

SRK conducted a site visit to the Granite Creek Project located in Nevada on February 4, 2026. The purpose of the geological site visit

was in part to discuss the latest geological model developed by i-80 geological team which has been reviewed in detail by SRK, to understand the controls on the mineralization and to provide recommendations for adjustments as appropriate. Final work

on the geological model was completed on February 26, 2026.

During the review it was noted that the major structure and control on

fluid movement is associated with a series of key faults striking northeast and northeast-east, located to the north of the historical open pit mining operations.

The primary structural control is the northeast / southwest-striking, steeply dipping normal faults related to extension. The structures are

the main conduits for mineralizing fluids moving along strike and vertically. Lithologic contacts, bedding and localized folds also play an important role for developing mineralized zones. High-grade mineralized zones are moderately continuous along

faults with the most prolific zones occurring at structural intersections along the Rangefront, Otto, Ogee and South Pacific fault corridors.

i-80 geologists produced an updated fault model for the Granite Creek Project. The updated model is based on the review of the following

datasets. A total of 62 faults have been identified by i-80 and modeled of which a total of 57 are considered active as controlling features in the geological model at a regional level. These faults and structures were defined and confirmed by:

•

Onsite core review and geological logging

•

Core photos for verification

•

Geological mapping of the underground workings

Mineralization at Granite Creek is controlled by a combination of structural, lithologic, and intrusive features within a complex

intrusion-related hydrothermal system. In addition to the regional model

i-80 have supplied a mine scale geological model, which contains

more localized scale offsets. The mine-scale model carries a dedicated mine-scale set M1–M25 (plus M1 Offset) that isn’t in the Regional model, while dropping many of the Regional-only structures (the NNW / NS sets, several NE / EW

faults, Felix, Delaney, Mag_West, etc.).

11.2.4

Estimation Domain Analysis

SRK reviewed the current underground operation at Granite Creek as part of the site inspection. The current geological model and mineralization

at Granite Creek has identified three main areas of mineralization defined as the Ogee, Otto, and SPZ Additional mineralization potential exists outside these three areas but is not currently considered in the underground model as further

exploration is needed to better define the continuity. The three areas are controlled by faulting in the local areas which has acted as controls on the mineralization. The key faults used to define the fault blocks which host the mineralization are:

•

CX West Fault

•

Range Front Fault

•

Domain Boundary A (Boundary A)

•

Domain Boundary B (Boundary B)

•

Domain Boundary C (Boundary C)

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The Ogee domain is bound to the south by domain Boundary C, to the north by domain Boundary

B, and the domain Boundary A, cuts the mineralization to the northeast. The Otto domain is located to the north of Ogee and is bound between the Boundary B limit and the Front Range Fault to the west, with the Boundary A surface limiting the

mineralization to the northeast. SRK notes some mineralization remains to the west of the Range Front Fault within a local domain referred to as Osgood.

Further exploration work will be needed to confirm the depth, extent and geotechnical conditions of the mineralization. The current focus for

i-80 during the most recent exploration has been to test the depth extensions of the SPZ at the operation which is limited to a fault block below the CW-West Fault, with the current depth extents limited to the Boundary C surface. Above the CW-West

Fault the mineralization has been defined for the purpose of this study as the SPZ hanging wall Zone (SPZhw). A summary of the controls used to define the broad mineralization domain blocks is shown in Figure

11-30.

A review of the lithology shows the Range Front Fault marks the change between the Ocl

(hanging wall) and Cp (footwall) units, with most of the mineralization occurring within the Ocl unit of the overlying Ocu units. It is therefore assumed these are considered good hosts for the mineralization but that the lithological contacts

between these two units do not act as a lithological control. Additional analysis of the statistics by lithology also indicated higher grades in the vicinity of the faults which supports the observations made during the site inspection and knowledge

from the current operation.

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Source: SRK, 2026

Figure 11-30: Plan Showing Main Defined Mineralization Domain Boundary Areas at the Granite Creek Underground Operation

Using the defined limits of the three main boundary areas the first step has been to identify within these areas the regions of potential

mineralization (Figure 11-31 and Figure 11-32). To undertake this SRK initially defined three low-grade indicator grade shells based on a nominal 0.5 g/t (0.015 oz/st)

Au cut-off. To create the grade shells the sampling has been limited to each domain, with a series of structural trends generated based on selected local faults mapped at the current operation and interpreted from the diamond drilling data. The

structural model is under continual review and updates by the i-80 geological team are reviewed by SRK. Overall, it is opinion of SRK that the structures selected are reasonable for consideration controlling the mineralization, so the trends have

been modeled in Leapfrog using the “strongest along inputs” for Ogee and Otto but using a “blended” approach at SPZ due to most of the interpretation at depth being based on diamond drilling interpretation vs. underground

operations. Using this method each structure is assigned a strength and range with the range of influence being between 9 m and 14 m (30 to 45 ft. )The interpretations were then finalized using an ISO Value of 0.45 (which is defined by Leapfrog as

the confidence that the probability of the indicator shell will contain the cut-off value), which upon review by the QP was deemed to provide a reasonable limit to the edge of the mineralization for which the data could be coded to undertake the

estimation domain analysis.

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Source: Source: SRK, 2026

Figure 11-31: Plan Showing Main Defined Mineralization Areas at the Granite Creek Underground Operation

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Source: SRK, 2026

Figure 11-32: Long Section (Looking Northwest) Showing Main Defined Mineralization Areas at the

Granite Creek Underground Operation

To undertake the estimation domain analysis SRK used a combination of Phinar X-10 (version

2.1.42.14) and Snowden Supervisor (v.9.2.5). Using the 0.5 g/t (0.015 oz/st) Au limits as a lower boundary of potentially economic mineralization SRK undertook a statistical analysis of the grade distribution within each domain using a combination

of log histograms, log-probability plots and noted in general there are still mixed populations within the domains defined. A review of the log-probability plots shows increases in the gradient of the curves above values in the order of 1.5 g/t

(0.044 oz/st) to 3.0 g/t (0.088 oz/st) Au within each of the domains. The changes in the population are noted to be subtle, and therefore further work on sensitivity was completed in Leapfrog to test a range of values. An example of the analysis is

shown in Figure 11-33 for the Otto domain with the interpreted possible trend breaks shown on the log-probability chart. Figure 11-34 shows the equivalent analysis using

Snowden supervisor for all domains on the log-probability highlighting potential changes in trend based on the QP’s opinion and analysis.

It is noted there remains a high-grade component (shown by the red line) for the mineralization above values of 10 to 15 g/t (0.438 oz/st) in

the example shown but based on the grade continuity and that typically < 5% of the data population within the boundary domains are above these values they have been combined with the >1.5 g/t (0.044 oz/st) grade domains.

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Source: SRK, 2026

Figure 11-33: Example of EDA Analysis in X-10 to Identify Potential Changes in Grade Populations

within the Ogee Domain

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Figure 11-34: Example of Log-Probability Plots Showing Polynomial Model Fitting (Pink) and Potential

Trend Breaks (Yellow)

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For each domain SRK completed a sensitivity analysis testing the impact of grade limits, the

use of structural trends, and variable search ranges during the indicator grade shell analysis. The final sensitivity test was completed initially on eight different cut-off grades 0.5, 1.0, 1.5, 2.0, 3.0, 3.5, 4.0, 10.0 g/t (0.292 oz/st). Based on

visual review and comparison to the i-80 mine production model it was then decided to focus on the grade ranges of 1.5 to 2.5 g/t (0.073 oz/st) for more detailed analysis. A total of 14 different models were generated on selected confidence levels

using various ISO values to understand the variable ranges in sample statistics for each scenario.

The results of each analysis were

captured with the wireframe volume, average grades, and the percentage of samples above and below cut-off tracked to determine the portion of internal waste for any given scenario. The results were then charted (Figure

11-35) and the final scenarios selected based on the QP’s opinion of an optimized search in terms of the proportion of internal waste vs. maintaining a reasonable volume and potential loss of metal

outside the indicator shells. These values tend to fall in the range of 0.44 to 0.45 ISO values based on the analysis.

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Source: SRK, 2026

Figure 11-35: Example of Sensitivity Analysis on ISO Values Completed by SRK for the Ogee Domain

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During the review SRK has considered the statistical support for the domains but has also

reviewed the continuity of the mineralization in 3D in Leapfrog to ensure grade continuity is maintained (Figure 11-36). Where available in the upper portions of the active mine the results have been compared

to the grade control sampling (muck and sludge sampling) to visually check the preferred parameters.

The review of the geological

boundaries to the grade control has also aided in the assessment of the key faults and orientation controls used for the interpretation. Figure 14 36 provides an example of the review, with the key faults shown in gray on. It is the QP’s

opinion that the grade boundaries of the high-grade domains closely follow the limits of the high-grade data as defined by the muck sampling. In the case of the Ogee domain in the example shown, the two key orientations of a north-northeast limb and

a northeast limb are clearly shown which is supported by the mining activities in these areas of the mine.

Source: SRK, 2026

Figure 11-36: Level Plan 4230 Showing Selected Ogee and Otto Domain Boundaries vs. Sampling

(Including Grade Control Muck Sampling)

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The final domains were initially exported to dxf format and provided to the i-80 Geology team

for review. Several small areas were noted where some manual adjustments to the interpretation were needed to improve the geological interpretation (namely in areas of lower sampling volumes), which has been completed by SRK and then final review

and agreement reached with the i-80 Geology team that the selected grade shells are representative of the underlying geological controls.

Based on the review and in summary the following domains have been defined by SRK for use in the underground Mineral Resource (Table 11-20), with the domains shown in Figure 11-37.

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Table 11-20: Summary of Domains and Criteria Used

Domain

Wireframe

Description/Criteria

OgeeHG

GM_EstDomain – OgeeHG.dxf

Indicator 2.0 g/t Au, Trend = Ogee FS 2025_12, Range =300 ft, ISOVALUE 0.45

OgeeLG

GM_EstDomain – OgeeLG.dxf

Indicator 0.5 g/t Au, Trend = Ogee FS 2025_12, , Range =300 ft, ISOVALUE 0.45

OttoHG

GM_EstDomain – OttoHG.dxf

Indicator 2.0 g/t Au, Trend = OT FS 2025_12, Range =300 ft, ISOVALUE 0.45

OttoLG

GM_EstDomain – OttoLG.dxf

Indicator 0.5 g/t Au, Trend = OT FS 2025_12, , Range =300 ft, ISOVALUE 0.45

SPZHG

GM_EstDomain – SPZHG.dxf

Indicator 1.5 g/t Au, Trend = SPZ FS 2025_12, Range =300 ft, ISOVALUE 0.45

SPZLG

GM_EstDomain – SPZLG.dxf

Indicator 0.5 g/t Au, Trend = SPZ FS 2025_12, Range =300 ft, ISOVALUE 0.45

SPZhwHG

GM_EstDomain – SPZhwHG.dxf

Indicator 2.5 g/t Au, Trend = SPZ FS 2025_12, Range =175 ft, ISOVALUE 0.45

SPZhwLG

GM_EstDomain – SPZhwLG.dxf

Indicator 0.5 g/t Au, Trend = SPZ FS 2025_12, Range =175 ft, ISOVALUE 0.45

Source: SRK, 2026

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Source: SRK, 2026

Figure 11-37: SRK Domain Model Used for Grade Estimation

As a final step, SRK has cross-checked statistically to ensure the domains are reasonable for the purpose of the grade estimation process and a

box-whisker plot of the final domained gold statistics is shown in Figure 11-38. It is the QP’s opinion that these represent valid domains for the purpose of estimation.

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Source: SRK, 2026

Figure 11-38: Box-Whisker Plot Showing Gold Grade Distributions within the SRK Domained Models,

showing mean grades.

11.2.5 Compositing

The composite length review indicates that approximately 90% of the samples taken are less than 1.53 m (5 ft) in length and that 98% of the

samples are less than 3.05 m (10 ft) in length (Figure 11-39). SRK completed this review on both the combined dataset and within the individual domains. Figure 11-40

shows the results of the analysis within the high-grade domains. The QP has opted to use a single composite length for all domains.

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Source: SRK, 2026

Figure 11-39: Log Probability Plot Showing Sampling Length in Diamond and RC Sampling at Granite

Creek

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Figure 11-40: Summary of Sampling Lengths Inside the High-Grade Domains within Diamond and RC

Drillholes

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Based on this review SRK selected 1.53 m (5 ft) for the composite lengths for all estimation

domains, as in the QP’s opinion this better reflected the downhole grade variations noted in the drilling than the larger 10 or 5 m (15 ft) composites as shown in Figure 11-39 and Figure 11-40. SRK also noted the difference in the mean grades between the raw and composited data reported closer in the 1.53 m (5 ft) composites than the longer 3.05 (10 ft) composites.

SRK completed a statistical comparison of the composite lengths and the mean grades for Au (g/t) before capping, which reported within < 1%

difference in the mean grades for the mineralized data. However, on a zone-by-zone basis the QP noted there is a slight increase in the mean grade within the OgeeHG domain (approximately 2.5%), which in the QP’s opinion is acceptable. Overall,

in the QP’s opinion the 2 m (5 ft) composites were preferred.

11.2.6

Assay Capping

High-grade capping is undertaken where data is no longer considered to be part of the main population. The QP completed the analysis based on

log and raw probability plots and histograms which can be used to distinguish the grades at which samples have significant impacts on the local estimation and affect is considered extreme. The QP notes that the mean grades within the different

domains are sensitive to changes in the capping values.

SRK completed a statistical analysis of the impact of grade capping by importing

the geologically domained coded composites into Phinar’s X10 Geo (X10) statistical software packages for review. During the analysis, SRK reviewed the percentage of metal loss and the reduction in the coefficient of variation (CoV). Examples

of the capping showing a review of the data distribution at the highest grades completed in X10 are shown in Figure 11-41.

The high-grade domains which account for the majority of the Mineral Resources are shown in Figure

11-41 and Figure 11-42. The selected capping values have also been cross-checked to the impact on the distributions using Snowden Supervisor (v9.0.4) to ensure the

selected capping is appropriate relative to the log-histograms, log probability, mean and variance cumulative frequency plots, and the distribution of the cumulative metal above the selected cap (Figure

11-43).

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Source: SRK, 2026

Figure 11-41: Examples of X-10 Capping Analysis Showing Capped Values in Yellow

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Source: SRK, 2026

Figure 11-42: Capping Analysis Showing Breaks in Data Population (Ogee HG Domain)

Source: SRK, 2026

Figure 11-43: Capping Analysis Showing Breaks in Data Population (Otto HG Domain)

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Figure 11-44: Capping Analysis Showing Breaks in Data Population (Ogee SPZ Domain)

The analysis process has been completed for all the key elements used in the estimation process covering the gold and the geo-metallurgical

parameters discussed in more detail in Section 14.3.10 of this report. The capping levels have been selected (Table 11-21) where the populations were identified by the QP to be breaking down, but also

with the goal of reducing the coefficient of variation (CoV) to acceptable levels for the application of geostatistical methodology (1.5 to 2.0).

The capping levels typically reduced the mean grades in the order of 2.5% to 5.0% for most domains, except OgeeLG which has the largest

significant differences noted (-15.8%). The basis for the more significant capping at this domain was the relatively high coefficient of variation (2.42 in the raw data), which needed more aggressive capping to reduce the CoV compared to the other

domains. The QP notes that continuing to monitor this domain during the mining process should be considered to either identify if additional high-grade materials have been missed during the geological modeling, or if the grade capping needs further

adjustment based on reconciliation data as mining progresses. In all the domains the CoV has been reduced from high variability >2.0 to approximately 1.5 to 1.7 which in the QP’s opinion is reasonable for the application of geostatistics

and estimation and the results are summarized in Table 11-21.

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Table 11-21: Summary of Raw vs. Capped Composite

Statistics per Domain

Name

Parameter

Count

Minimum

Maximum

Mean

Variance

Std Dev

CoV

%

Change

OgeeHG

Au (g/t)

1668

0.00

134.85

12.37

283.70

16.84

1.36

-2.4%

AuCap

1668

0.00

70.00

12.07

239.67

15.48

1.28

OgeeLG

Au (g/t)

1768

0.00

82.43

3.44

69.00

8.31

2.42

-15.8%

AuCap

1768

0.00

24.00

2.89

31.25

5.59

1.93

OttoHG

Au (g/t)

1150

0.00

95.50

8.32

136.07

11.66

1.40

-1.4%

AuCap

1150

0.00

62.00

8.20

120.55

10.98

1.34

OttoLG

Au (g/t)

2030

0.00

67.20

2.58

29.39

5.42

2.10

-2.5%

AuCap

2030

0.00

32.00

2.52

24.50

4.95

1.97

SPZHG

Au (g/t)

940

0.00

140.91

9.43

159.62

12.63

1.34

-2.3%

AuCap

940

0.00

52.50

9.22

129.60

11.38

1.23

SPZLG

Au (g/t)

568

0.00

50.40

2.05

21.81

4.67

2.28

-6.3%

AuCap

568

0.00

20.00

1.92

14.87

3.86

2.01

SPZhwHG

Au (g/t)

44

0.02

64.50

11.91

199.53

14.13

1.19

0.0%

AuCap

44

0.02

64.50

11.91

199.53

14.13

1.19

SPZhwLG

Au (g/t)

194

0.00

10.90

1.41

2.66

1.63

1.15

-2.4%

AuCap

194

0.00

7.50

1.38

2.13

1.46

1.06

Source: SRK, 2026

11.2.7

Variogram Analysis and Modeling

The QP reviewed the geostatistical properties of the domains using Snowden Supervisor (v9.0.4) variogram analysis on the capped 1.53 m (5.02

ft) composite dataset for the Otto, Ogee, and SPZ domains (LG and HG domains for each). Based on the size of the data populations, initial test work for semi-variograms in the SPZhw (LG and HG) returned poor structures for directional variogram;

therefore, the QP elected to use omni- variograms within these domains.

The SRK workflow included review of the radial plot (to define the

general orientation), then definition of the Major, Semi-Major, and minor axis variograms, which in most cases have been orientated to a dip of 55° to 80°, dip azimuth of 130° to 155° and pitches ranging from of 65° to 120°.

The dip and orientations have been selected based on discussions with the Geology team’s analysis of key structures localized to each domain and visual trends noted in the mineralization. Variograms have been modeled using a combination of

variograms and correlograms depending on the quality of the experimental variogram data, which have been transformed and normalized in Supervisor and the parameters transferred to Leapfrog®

for the estimation process. Summary of gold variogram parameters are shown in Table 11-22 and Figure 11-45 through Figure 11-47.

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Table 11-22: Summary of Gold Variogram Parameters

Domain

Nugget

s1_sill

s1_maj

s1_semi

s1_min

s2_sill

s2_maj

s2_semi

s2_min

Dip

dip_azi

Pitch

Variance

OgeeHG

0.20

0.44

20

15

15

0.37

130

55

50

80

165

100

214.10

OgeeLG

0.20

0.44

20

15

15

0.37

130

55

50

80

155

65

25.80

OttoHG

0.22

0.55

20

10

10

0.23

70

60

30

80

155

65

107.90

OttoLG

0.27

0.58

10

10

10

0.15

115

55

35

80

155

65

23.80

SPZHG

0.24

0.44

40

30

15

0.32

230

175

50

55

130

120

125.60

SPZLG

0.22

0.55

10

10

10

0.23

70

60

30

80

155

65

13.90

SPZhwHG

0.21

0.39

10

10

10

0.40

65

65

65

80

155

65

269.60

SPZhwLG

0.21

0.51

10

10

10

0.28

60

60

60

80

155

65

2.20

Source: SRK, 2026

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Source: SRK, 2026

Figure 11-45: Modeled Normal Score Downhole Variograms and Directional Correlograms for OgeeHG

(Gold)

Source: SRK, 2026

Figure 11-46: Modeled Normal Score Downhole and Directional Variograms for OttoHG (Gold)

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Source: SRK, 2026

Figure 11-47: Modeled Normal Score Downhole and Directional Variograms for SPZHG (Gold)

11.2.8

Block Model Parameters

SRK has produced the block models using Leapfrog® Edge (version 2025.3.1). The

procedure involved construction of wireframe models for the fault networks, lithology domains, definition of resource domains (Ogee, Otto, SPZ and the SPZhw), split into low and high-grade subdomains. SRK produced a single block model for

underground portion of the Project to cover the main areas of mineralization. A single model was generated to cover the estimation of all the key domains. Estimation was completed based on coded 1.53 m (5 ft) composites (based on the updated defined

wireframes), which have been capped to appropriate levels.

The resource evaluation work was completed by SRK. Grade estimation has been

based on block dimensions of 4.57 m x 4.57 m x 4.57 m or (15 ft x 15 ft x 15 ft). The block size reflects potential size variations for the geological units plus consideration for any underground smallest mining units (SMU). The block size is

consistent with the current grade underground production models to aid future reconciliation. SRK utilized sub-blocking methodology in Leapfrog®

(Octree), which allows subdivision of the parent block by division of (4, 8, 16, and 32) to accurately reflect the defined mineralization and lithological models. SRK has utilized sub-blocking to accurately

reflect the defined geological models, with a sub-block size of 0.57 m x 0.57 m x 1.14 m (1.875 ft x 1.875 ft x 3.75 ft) used to reflect the thickness and orientation of the domain wireframes (Table 11-23). It was also decided through discussions with the mining engineer to use an orientated grid which aligns to the strike of the mineralization. SRK used an azimuth of 60 degrees to align to the primary

mineralization orientation.

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Table 11-23: Summary of Rotated Prototype

Parameters Used in Granite Creek Model

Item

Origin

Extent

Block Size

Block

Count

Subblock

(ft)

(m)

(ft)

(m)

X

7,775.500

2,295

15

4.75

153

1.875

0.57

Y

11,743.045

4,200

15

4.75

280

1.875

0.57

Z

5,960.000

3,375

15

4.75

225

3.75

1.14

Rotation

60º (azimuth)

Source: SRK, 2026

11.2.9

Estimation Methodology

Grades have been interpolated for Au (g/t), using a combination of Ordinary Kriging and Inverse Distance Weighting (Power 3). SRK tested

several estimation scenarios for the Ogee, Otto, and SPZ domains. The primary focus for the parameter selection has been the OgeeHG and SPZHG (based on Ogee HG representing the current mining areas and SPZ HG providing the longer-term potential

within the main mineralization domains) to identify the sensitivity of the estimates to the maximum number of samples being used and the assumption to use a maximum of three composites per drillhole during the estimation process. The QP did not

lower performance of the estimators based on slope of regression and negative weights within the Otto HG domain compared to Ogee and SPZ HG domains, which is a function of the variogram, however it is the QP view that keeping the estimation

parameters relatively consistent in terms of block size and number of samples is preferred.

SRK has utilized variable orientation searches

to control the orientation of the estimates for the Granite Creek deposit. SRK tested both linear and variable orientation (default dip and strike orientations) for each zone.

The principal direction of mineralization can change across a domain, such as when the domain features an undulating, gently folded structure.

Using a single fixed orientation for the sample search and the variogram can result in poor sample selection and weighting locally. Using a variable orientation makes it possible to re-orient the search and

variogram according to local characteristics, which results in improved local value estimates. Variable orientations can have multiple inputs, including veins, bedding or faults, and practically are based on any selected mesh in the project. The

variable orientation can be applied to Kriging and inverse distance estimators.

To set up the variable orientations in Leapfrog the QP has

worked with the i-80 Geology team to identify the key features within each domain that impact the local grade distribution. Given the predominance of the higher-grade mineralization to be controlled by

faulting, selected fault surfaces formed the majority of the controls on the orientation. The i-80 team have been working during 2025 to 2026 to integrate variable search orientation into the operational level

grade estimates with success. To determine the final orientation parameters for each domain the QP used the initial models as presented by i-80 and then refined to the controls based on visual inspection of

rotations within the block model to avoid any sharp changes in the grade estimates outside the areas of influence of the faults. The QP noted a few occasions where the search ranges were being impacted by some mine scale cross faulting outside the

area of the mineralization and has therefore removed these from the analysis. A summary of the final fault structures used per vein is included in Table 11-24.

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Table 11-24: Summary of Faults and Key Structures

Used to Define Variable Orientations for Estimation

Azimuth: 0

Plunge: 0

VO, Ogee 2026_02

VO, Otto_RF 2026_02

VO, SPZ 2026_02 Exclude NE_12_Offset

GC - Mine: EW_4

GC - Mine: Adam Peak Offset Fault

GC - Mine: M9

GC - Mine: LH Offset Fault

GC - Mine: Adam_Peak_Fault

GC - Mine: M14

GC - Mine: LH_Fault

GC - Mine: EW_5

GC - Mine: M19

GC - Mine: M1

GC - Mine: M10

GC - Mine: M20

GC - Mine: M1 Offset

GC - Mine: M11

GC - Mine: M21

GC - Mine: M2

GC - Mine: M12

GC - Mine: M23

GC - Mine: M4

GC - Mine: M13

GC - Mine: M24

GC - Mine: M5

GC - Mine: M14

GC - Mine: M25

GC - Mine: M7

GC - Mine: M15

GC - Regional: Adam Peak Offset Fault

GC - Mine: M8

GC - Mine: NE_1

GC - Regional: Otto Offset Fault

GC - Mine: M9

GC - Mine: NE_10

GC - Regional: Otto Offset Splay

GC - Mine: M16

GC - Mine: NW_1

GC - Regional: Otto_Zone

GC - Mine: M17

GC - Mine: NW_2

GC - Regional: Otto_Zone_2

GC - Mine: M18

GC - Mine: Ogee Fault

GC - Regional: Otto_Zone_3

GC - Mine: NE_3

GC - Mine: Otto Offset Fault

GC - Regional: SPZ

GC - Mine: NE_13

GC - Mine: Otto_Fault

GC - Mine: NE_14

GC - Mine: Otto_Zone

GC - Mine: NNE 1

GC - Mine: Range_Front_Fault

GC - Mine: Ogee Fault

GC - Mine: Splat_Fault

GC - Mine: Otto Offset Splay

GC - Mine: Otto_Zone_2

GC - Mine: SPZ

Source: SRK, 2026

SRK also undertook a review of the use of a hard and soft boundary between the low-grade / high-grade

and boundaries to test if a hard or soft boundary was more appropriate. Given the use of 2 m(5 ft) composites, the analysis focused on 2 m (5 ft) and 3 m (10 ft) boundaries to allow the influence of either single or two samples across the various

boundaries. Based on visual review of the models and from boundary analysis the 2 m (5 ft) soft boundary was selected for the HG and LG domains.

SRK undertook a Kriging Neighborhood Analysis (KNA) in Snowden Supervisor to aid in the definition of the minimum and a maximum number of

composites to be used. The analysis looked at a range of block sizes (Figure 11-48) with 15 ft x 15 ft x 15 ft selected, and composites numbers ranging from 2 to 60 samples to review the impact on potential

smoothing on the grades. SRK noted for both domains through the KNA in Snowden that the number of negative weights increased significantly above 26 to 30 composites. A review of the statistics indicates that the sum of the negative weights first

impacts the estimates in the 8 to 12 sample range, and that within this range the estimation variance of the blocks has already been reduced by 52% to 59% with the reduction levelling out in the 16 to 20 composite range and shown in. Figure 11-49 and Figure 11-50.

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Source: SRK, 2026

Figure 11-48: Block Size Analysis, Ogee HG Domain

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Source: SRK, 2026

Figure 11-49: KNA Analysis on Number of Samples vs. Slope of Regression (Top) and Impact of

Negative Weights (Bottom) — Ogee HG Domain

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Source: SRK, 2026

Figure 11-50: KNA Analysis on Number of Samples vs. Slope of Regression (Top) and Impact of

Negative Weights (Bottom) — SPZ HG Domain

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Based on the KNA analysis the QP completed a number of test estimates with various minimum

and maximum number of composites with the view of practically testing the selection ranges presented. The QP then reviewed the results and considered the impact of potential grade smoothing within each of the scenarios and concluded that the

preferred number of minimum and maximum samples in the first search should be in the order of 9 (minimum) and 12 (maximum) composites. A maximum of 4 composites per hole were used in the first and second pass which was reduced to 3 composites in the

final pass.

The decision to use a maximum of 12 composites vs. 16 or 20 composites (3 holes vs. 4 or 5 holes) was selected based on visual

validation and discussions with the i-80 technical team based on the current mining assumptions and model performance. A summary of the final gold parameters is contained in Table 11-25..

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Table 11-25: Summary of Estimation Parameters for

Gold Estimates

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier Restrictions

Drillhole Limit

Estimator

Name

Domain

Numeric Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples per

Hole

Apply Drillhole Limit

per Sector

Kr, AuCap: OgeeHG Trend SftP1

OgeeHG

AuCap

55

45

30

VO, Ogee 2026_02

9

12

None

4

TRUE

Kr, AuCap: OgeeHG Trend SftP2

OgeeHG

AuCap

125

90

60

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCap: OgeeHG Trend SftP3

OgeeHG

AuCap

210

180

90

VO, Ogee 2026_02

5

12

None

3

TRUE

Kr, AuCap: OgeeLG Trend SftP1

OgeeLG

AuCap

65

65

30

VO, Ogee 2026_02

9

12

None

4

TRUE

Kr, AuCap: OgeeLG Trend SftP2

OgeeLG

AuCap

130

130

60

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCap: OgeeLG Trend SftP3

OgeeLG

AuCap

210

210

100

VO, Ogee 2026_02

4

12

None

3

TRUE

Kr, AuCap: OttoHG Trend SftP1

OttoHG

AuCap

55

45

30

VO, Otto_RF 2026_02

9

12

None

4

TRUE

Kr, AuCap: OttoHG Trend SftP2

OttoHG

AuCap

105

90

60

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCap: OttoHG Trend SftP3

OttoHG

AuCap

330

300

125

VO, Otto_RF 2026_02

4

12

None

3

TRUE

Kr, AuCap: OttoLG Trend SftP1

OttoLG

AuCap

55

30

20

VO, Otto_RF 2026_02

9

12

None

4

TRUE

Kr, AuCap: OttoLG Trend SftP2

OttoLG

AuCap

115

55

40

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCap: OttoLG Trend SftP3

OttoLG

AuCap

172.5

110

70

VO, Otto_RF 2026_02

2

12

None

3

TRUE

Kr, AuCap: SPZHG Trend SftP1

SPZHG

AuCap

115

90

30

VO, SPZ 2026_02 Exclude NE_12_Offset

9

12

None

4

TRUE

Kr, AuCap: SPZHG Trend SftP2

SPZHG

AuCap

175

135

40

VO, SPZ 2026_02 Exclude NE_12_Offset

5

16

None

4

TRUE

Kr, AuCap: SPZHG Trend SftP3

SPZHG

AuCap

230

175

50

VO, SPZ 2026_02 Exclude NE_12_Offset

3

12

None

3

TRUE

Kr, AuCap: SPZhwHG Trend SftP1

SPZhwHG

AuCap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

Kr, AuCap: SPZhwHG Trend SftP2

SPZhwHG

AuCap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, AuCap: SPZhwHG Trend SftP3

SPZhwHG

AuCap

180

180

60

VO, SPZhw 2026_02

2

12

None

3

TRUE

Kr, AuCap: SPZhwLG Trend SftP1

SPZhwLG

AuCap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

Kr, AuCap: SPZhwLG Trend SftP2

SPZhwLG

AuCap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, AuCap: SPZhwLG Trend SftP3

SPZhwLG

AuCap

180

180

60

VO, SPZhw 2026_02

2

12

None

3

TRUE

Kr, AuCap: SPZLG Trend SftP1

SPZLG

AuCap

85

65

30

VO, SPZ 2026_02 Exclude NE_12_Offset

9

12

None

4

TRUE

Kr, AuCap: SPZLG Trend SftP2

SPZLG

AuCap

120

100

40

VO, SPZ 2026_02 Exclude NE_12_Offset

5

16

None

4

TRUE

Kr, AuCap: SPZLG Trend SftP3

SPZLG

AuCap

240

200

75

VO, SPZ 2026_02 Exclude NE_12_Offset

2

12

None

3

TRUE

Source: SRK, 2026

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11.2.10 Geo-Metallurgical

Parameters

The mineralization at Granite includes refractory mineralization which requires different processing routes to maximize the recovery and value

for the material. The decision to direct material heap leaching or CIL methods and those that require autoclave treatment .is based on metallurgical indicators. The final decisions for material routing are expected to happen during the short-term

modeling and production stage of the Project. However, to provide a more accurate reflection of the mine plan within the long-term models and Mineral Reserves calculations, SRK has estimated the key metallurgical factors into the block model. The

factors are based on LECO assay analysis of selected samples taken by an i-80 geologist during the recent i-80 infill drilling program.

SRK completed a review of the key parameters which can impact the potential decision for processing routes and expected recoveries which

included:

•

Cyanide Soluble Gold (ppm) – AuCN_BestValue_ppm

•

Total Organic Carbon (%) – TOC_pct

•

PregRob (%) – PregRob_pct

•

Sulfide Sulfur (%) – SulfideSulfur_pct

•

CO3 (%) – CO3_LT_pct

Prior to the estimation of these elements the QP completed a review of the sampling available as the LECO test work has not been completed

routinely during the exploration phases. The results of the analysis (Table 11-26) indicated that there is limited coverage in the additional elements needed with limited datasets in the upper portions of the

mine in Ogee and Otto. It is the QP’s opinion that estimates within these regions would not be representative if the same restriction used in the grade estimates on only using Diamond and RC drilling sampling was applied. It is noted that the

coverage with the addition of the grade control sampling from within the current underground operation would significantly improve the sampling coverage for use in the metallurgical assessment.

The additional sampling consists of sampling defined as Muck Sampling and Sludge Sampling. The sludge sampling has not been included in the

analysis as it is the QP’s opinion that the samples are not representative as all sampling is assigned a default length. In comparison, the muck samples are taken from each round as the mining advances, and while these are not considered as

representative across the width of the advance, they are representative of each advance. The QP has therefore taken the approach to supplement the exploration database as used to define the gold estimates, with the aim to provide a more

representative local scale variability in the metallurgical parameters.

In addition to the parameters defined above and that the ratio of

gold vs. cyanide soluble gold ratio may impact the recovery curves to ensure accurate comparisons between the two values, SRK has also completed an additional estimate for gold using the muck samples defined in the block model as AuGC (g/t). In the

cases of the low-grade domains where there remains gaps in the paired data for Au and AuCN only the paired datasets have been used for OgeeLG, OttoLG and SPZLG.

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Table 11-26: Summary Statistics and Coverage of

Raw Sampling per Metallurgical Parameter vs. Gold Coverage

Name

Count

Length

Mean

Std Dev

CoV

Variance

Minimum

Maximum

% Coverage

Au_BestValue_ppm

OgeeHG

1,970.0

8,097.9

12.12

18.02

1.49

324.72

0.000

143.07

OgeeLG

1,912.0

8,619.6

3.40

8.96

2.63

80.27

0.000

114.67

OttoHG

1,222.0

5,582.1

8.21

12.27

1.49

150.57

0.000

130.42

OttoLG

2,061.0

9,870.9

2.58

5.66

2.20

32.06

0.000

67.20

SPZHG

1,063.0

4,554.6

9.31

13.98

1.50

195.58

0.000

140.91

SPZhwHG

53.0

215.1

11.93

15.80

1.32

249.60

0.008

79.00

SPZhwLG

193.0

947.4

1.41

1.92

1.36

3.67

0.000

15.15

SPZLG

601.0

2,704.1

2.03

5.14

2.54

26.38

0.000

51.50

AuCN_BestValue_ppm

OgeeHG

972.0

4,242.4

11.60

13.88

1.20

192.65

0.001

125.73

52.4

OgeeLG

577.0

2,666.0

5.38

9.10

1.69

82.72

0.001

104.40

30.9

OttoHG

578.0

2,630.3

4.13

8.69

2.11

75.59

0.015

72.37

47.1

OttoLG

677.0

3,189.9

1.99

4.57

2.30

20.86

0.001

34.48

32.3

SPZHG

748.0

3,135.2

2.05

5.24

2.55

27.42

0.001

47.74

68.8

SPZhwHG

25.0

106.6

1.53

1.76

1.15

3.11

0.034

6.10

49.6

SPZhwLG

113.0

553.9

0.46

0.65

1.42

0.42

0.014

3.27

58.5

SPZLG

244.0

1,058.2

1.13

2.91

2.58

8.47

0.001

19.49

39.1

CO3_LT_pct

OgeeHG

651.0

2,631.7

11.55

13.43

1.16

180.41

0.000

62.26

32.5

OgeeLG

428.0

1,912.3

15.99

13.90

0.87

193.21

0.000

54.88

22.2

OttoHG

493.0

2,187.1

19.30

15.15

0.79

229.59

0.000

61.20

39.2

OttoLG

444.0

2,051.7

14.47

14.54

1.00

211.40

0.000

66.06

20.8

SPZHG

789.0

3,344.9

12.66

13.20

1.04

174.17

0.000

64.66

73.4

SPZhwHG

17.0

62.7

6.48

7.91

1.22

62.62

0.200

29.35

29.1

SPZhwLG

44.0

208.0

13.58

16.19

1.19

262.08

0.864

54.60

22.0

SPZLG

327.0

1,446.4

17.14

16.40

0.96

269.09

0.000

57.10

53.5

PregRob_Pct

OgeeHG

144.0

607.8

6.87

15.16

2.21

229.87

0.000

95.00

7.5

OgeeLG

71.0

295.0

2.56

6.29

2.46

39.59

0.000

38.80

3.4

OttoHG

181.0

821.6

3.93

8.83

2.25

78.03

0.000

75.50

14.7

OttoLG

77.0

335.6

2.67

6.47

2.42

41.91

0.000

47.10

3.4

SPZHG

667.0

2,796.9

3.06

9.17

3.00

84.16

0.000

164.10

61.4

SPZhwHG

14.0

52.2

3.08

5.32

1.73

28.35

0.000

16.70

24.3

SPZhwLG

26.0

128.0

2.17

5.25

2.41

27.51

0.000

19.40

13.5

SPZLG

201.0

861.0

2.14

4.21

1.97

17.76

0.000

26.00

31.8

SulfideSulfur_pct

OgeeHG

651.0

2,631.7

0.50

0.94

1.87

0.88

0.000

6.11

32.5

OgeeLG

428.0

1,912.3

0.50

0.91

1.84

0.83

0.000

5.47

22.2

OttoHG

493.0

2,187.1

1.18

1.28

1.09

1.63

0.000

6.44

39.2

OttoLG

444.0

2,051.7

1.10

1.35

1.23

1.83

0.000

6.72

20.8

SPZHG

789.0

3,344.9

1.76

1.26

0.71

1.59

0.000

8.63

73.4

SPZhwHG

17.0

62.7

2.69

0.84

0.31

0.71

1.803

4.85

29.1

SPZhwLG

44.0

208.0

1.43

1.17

0.82

1.37

0.000

4.20

22.0

SPZLG

327.0

1,446.4

1.10

1.12

1.02

1.25

0.000

9.95

53.5

TOC_pct_D

OgeeHG

651.0

2,631.7

0.27

0.25

0.96

0.06

0.017

2.83

32.5

OgeeLG

428.0

1,912.3

0.25

0.23

0.92

0.05

0.000

2.02

22.2

OttoHG

493.0

2,187.1

0.19

0.18

0.94

0.03

0.002

1.36

39.2

OttoLG

444.0

2,051.7

0.21

0.24

1.17

0.06

0.000

5.17

20.8

SPZHG

789.0

3,344.9

0.47

0.60

1.26

0.35

0.008

6.79

73.4

SPZhwHG

17.0

62.7

1.64

2.00

1.22

3.99

0.082

5.53

29.1

SPZhwLG

44.0

208.0

1.11

1.14

1.03

1.30

0.019

4.07

22.0

SPZLG

327.0

1,446.4

0.62

0.85

1.38

0.73

0.000

5.10

53.5

Source: SRK, 2026

Note: % Coverage is based on the total length per domain of each element versus the Au BestValue ppm coverage.

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Based on the revised database using the muck samples, the process workflow used by SRK has

followed the same procedures as presented in the gold estimation process (Section 14.3.9). Capping has been completed per element with the AuGC values using the same capping parameters as presented in the gold estimates, with all other elements

considered on a case-by-case basis per domain. The capped statistics are presented in Table 11-27.

Table 11-27: Summary Statistics of Capped Composites (5 ft), Showing Capping (Maximum) Levels per

Metallurgical Elements

Name

Domain

Count

Length

Mean

Std

Dev

CoV

Variance

Minimum

Maximum

AuCn_cap

OgeeHG

938.0

4,667.3

11.22

11.91

1.06

141.87

0.00

52.00

OgeeLG

594.0

2,944.3

4.80

6.58

1.37

43.26

0.00

22.50

OttoHG

583.0

2,887.1

3.74

6.59

1.76

43.40

0.02

35.00

OttoLG

725.0

3,566.0

1.91

4.13

2.16

17.07

0.00

25.00

SPZHG

680.0

3,349.7

1.93

4.06

2.10

16.48

0.00

25.00

SPZhwHG

22.0

110.1

1.60

1.69

1.06

2.85

0.03

5.50

SPZhwLG

120.0

594.0

0.48

0.66

1.38

0.44

0.02

3.25

SPZLG

247.0

1,200.2

1.11

2.50

2.25

6.25

0.00

12.50

PregRob_Cap

OgeeHG

147.0

730.5

6.01

9.81

1.63

96.20

0.00

47.50

OgeeLG

73.0

355.5

2.24

4.54

2.02

20.59

0.00

19.50

OttoHG

188.0

938.4

3.25

4.74

1.45

22.42

0.00

21.50

OttoLG

85.0

413.9

2.26

3.62

1.60

13.12

0.00

17.50

SPZHG

613.0

3,026.2

2.51

4.28

1.70

18.31

0.00

20.00

SPZhwHG

11.0

55.2

3.24

4.55

1.40

20.72

0.00

15.00

SPZhwLG

32.0

157.0

2.03

4.47

2.20

19.96

0.00

15.00

SPZLG

205.0

999.8

2.24

3.83

1.71

14.70

0.00

15.50

TOCcap

OgeeHG

562.0

2,760.5

0.27

0.23

0.86

0.05

0.03

1.60

OgeeLG

403.0

1,959.5

0.25

0.21

0.84

0.04

0.00

1.35

OttoHG

469.0

2,317.5

0.20

0.18

0.93

0.03

0.00

1.36

OttoLG

452.0

2,181.5

0.21

0.20

0.92

0.04

0.00

1.95

SPZHG

711.0

3,506.7

0.46

0.52

1.12

0.27

0.01

3.30

SPZhwHG

13.0

65.2

1.60

1.96

1.22

3.82

0.10

5.00

SPZhwLG

45.0

217.5

1.49

2.67

1.79

7.13

0.02

13.30

SPZLG

316.0

1,540.4

0.61

0.79

1.30

0.62

0.01

3.50

SulfSulf_Cap

OgeeHG

562.0

2,760.5

0.49

0.82

1.67

0.67

0.00

3.65

OgeeLG

403.0

1,959.5

0.48

0.83

1.75

0.70

0.00

3.20

OttoHG

469.0

2,317.5

1.21

1.22

1.01

1.49

0.00

4.75

OttoLG

452.0

2,181.5

1.11

1.25

1.13

1.56

0.00

4.75

SPZHG

711.0

3,506.7

1.75

1.11

0.63

1.24

0.00

5.00

SPZhwHG

13.0

65.2

2.68

0.65

0.24

0.42

1.80

3.60

SPZhwLG

45.0

217.5

1.44

1.12

0.78

1.25

0.00

3.75

SPZLG

316.0

1,540.4

1.09

1.02

0.94

1.05

0.00

4.15

CO3cap

OgeeHG

562.0

2,760.5

11.30

12.32

1.09

151.80

0.00

50.00

OgeeLG

403.0

1,959.5

15.96

13.11

0.82

171.80

0.00

50.00

OttoHG

469.0

2,317.5

18.76

13.78

0.73

189.94

0.00

47.00

OttoLG

452.0

2,181.5

14.24

13.79

0.97

190.13

0.00

48.00

SPZHG

711.0

3,506.7

12.85

12.60

0.98

158.67

0.00

55.00

SPZhwHG

13.0

65.2

6.34

6.60

1.04

43.56

0.26

25.00

SPZhwLG

45.0

217.5

13.72

15.28

1.11

233.38

1.04

50.00

SPZLG

316.0

1,540.4

17.13

15.58

0.91

242.68

0.09

52.00

Source: SRK, 2026

SRK reviewed the geostatistical properties of the domains using Snowden Supervisor (version 9.0.4.0). Variography was completed on each of the

parameters, which have been modeled using nested double structured spherical variograms.

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

243

Variograms have been modeled using a combination of variograms and correlograms depending on

the quality of the experimental variogram data, which have been transformed and normalized in Supervisor and the parameters transferred to Leapfrog® for the estimation process. A summary of

the parameters are shown in Table 11-28.

Estimation has been completed using a combination of

either Ordinary Kriging or Inverse Distance weighting using the capped 2 m (5 ft) composites and the same variable orientation as used for the primary elements. The estimation has been completed using a three-pass process using the parameters as

presented in Table 11-28 to Table 11-33.

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

244

Table 11-28: Summary of Variogram Parameters Used

for Metallurgical Elements

Variable

Domain

Nugget

s1_sill

s1_major

s1_semi_major

s1_minor

s2_sill

s2_major

s2_semi_major

s2_minor

dip

dip_azimuth

Pitch

Variance

AuCN (cyanide-soluble gold)

OgeeHG

0.091

0.678

24

26

17

0.23

65

97

60

80

165

100

65.6

AuCN (cyanide-soluble gold)

OgeeLG

0.316

0.39

20

25

16

0.294

95

117

70

80

155

95

20.73

AuCN (cyanide-soluble gold)

OttoHG

0.159

0.488

22

14

20

0.353

300

265

80

70

140

45

24.68

AuCN (cyanide-soluble gold)

OttoLG

0.297

0.411

28

19

20

0.293

300

80

100

55

140

90

9.68

AuCN (cyanide-soluble gold)

SPZHG

0.345

0.329

30

30

15

0.326

170

120

50

55

130

120

15.1

AuCN (cyanide-soluble gold)

SPZLG

0.343

0.427

40

17

10

0.23

160

135

50

0

0

90

5.03

PregRob (preg-robbing)

OgeeHG

0.274

0.596

20

25

20

0.129

100

115

95

70

165

100

52.98

PregRob (preg-robbing)

OgeeLG

0.239

0.609

20

20

12

0.152

100

95

60

75

165

100

14.63

PregRob (preg-robbing)

OttoHG

0.174

0.64

30

30

17

0.187

310

300

58

80

150

45

15.19

PregRob (preg-robbing)

OttoLG

0.184

0.646

50

50

20

0.171

208

250

60

80

160

50

7.31

PregRob (preg-robbing)

SPZHG

0.222

0.427

50

50

27

0.351

260

260

73

55

130

120

16.95

PregRob (preg-robbing)

SPZLG

0.19

0.32

36

36

36

0.49

330

330

330

0

0

90

12.18

CO3 (carbonate)

OgeeHG

0.142

0.645

30

30

19

0.214

150

160

105

70

165

100

72.87

CO3 (carbonate)

OgeeLG

0.096

0.322

25

22

27

0.581

390

170

150

70

140

45

112.36

CO3 (carbonate)

OttoHG

0.095

0.42

26

20

15

0.485

253

210

117

70

140

45

88.66

CO3 (carbonate)

OttoLG

0.096

0.322

25

22

27

0.581

390

170

150

70

140

45

112.36

CO3 (carbonate)

SPZHG

0.122

0.413

60

26

17

0.465

300

285

90

55

130

120

148.29

CO3 (carbonate)

SPZLG

0.14

0.52

74

74

74

0.34

267

267

267

0

0

90

213.8

TOC (total organic carbon)

OgeeHG

0.114

0.503

25

30

20

0.383

248

204

215

55

345

80

0.04

TOC (total organic carbon)

OgeeLG

0.115

0.368

16

16

15

0.517

160

160

125

55

345

80

0.05

TOC (total organic carbon)

OttoHG

0.119

0.604

50

25

15

0.277

280

140

65

70

140

45

0.03

TOC (total organic carbon)

OttoLG

0.133

0.532

25

25

30

0.336

179

110

110

50

140

50

0.06

TOC (total organic carbon)

SPZHG

0.169

0.425

50

40

30

0.406

280

260

80

55

130

120

0.26

TOC (total organic carbon)

SPZLG

0.1

0.9

370

370

370

SulfS (sulfide sulfur)

OgeeHG

0.153

0.488

23

7

16

0.359

230

112

190

70

140

40

0.31

SulfS (sulfide sulfur)

OgeeLG

0.313

0.337

20

20

25

0.35

275

150

145

70

320

140

0.22

SulfS (sulfide sulfur)

OttoHG

0.211

0.432

30

41

20

0.357

360

260

62

70

140

45

0.76

SulfS (sulfide sulfur)

OttoLG

0.15

0.363

21

25

25

0.487

150

228

64

55

140

90

0.8

SulfS (sulfide sulfur)

SPZHG

0.21

0.301

40

40

15

0.489

330

210

60

55

130

120

1.15

SulfS (sulfide sulfur)

SPZLG

0.07

0.21

15

15

15

0.72

170

170

170

0

0

90

0.9256

Source: SRK, 2026

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

245

Table 11-29: Summary of Search Parameters Used for

AuCN

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier

Restrictions

Drillhole Limit

Estimator Name

Domain

Numeric Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples

per Hole

Apply Drillhole Limit

per Sector

Kr, AuCN_Cap: Domain Trend P1

OgeeHGtrend

AuCn_cap

45

75

30

VO, Ogee 2026_02

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

OgeeHGtrend

AuCn_cap

65

97

60

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

OgeeHGtrend

AuCn_cap

210

210

90

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P1

OgeeLGtrend

AuCn_cap

50

60

35

VO, Ogee 2026_02

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

OgeeLGtrend

AuCn_cap

100

120

70

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

OgeeLGtrend

AuCn_cap

210

210

100

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P1

OttoHGtrend

AuCn_cap

75

65

40

VO, Otto_RF 2026_02

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

OttoHGtrend

AuCn_cap

150

135

60

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

OttoHGtrend

AuCn_cap

330

300

125

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P1

OttoLGtrend

AuCn_cap

55

30

20

VO, Otto_RF 2026_02

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

OttoLGtrend

AuCn_cap

115

55

40

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

OttoLGtrend

AuCn_cap

172.5

110

70

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P1

SPZHGtrend

AuCn_cap

85

60

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

SPZHGtrend

AuCn_cap

125

85

50

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

SPZHGtrend

AuCn_cap

230

175

60

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

ID, AuCN_BestValue_ppm_D: Domain Trend P1

SPZhwHGtrend

AuCn_cap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

ID, AuCN_BestValue_ppm_D: Domain Trend P2

SPZhwHGtrend

AuCn_cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, AuCN_BestValue_ppm_D: Domain Trend P3

SPZhwHGtrend

AuCn_cap

180

180

60

VO, SPZhw 2026_02

4

12

None

4

TRUE

ID, AuCN_Cap: Domain Trend P1

SPZhwLGtrend

AuCn_cap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

ID, AuCN_Cap: Domain Trend P2

SPZhwLGtrend

AuCn_cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, AuCN_Cap: Domain Trend P3

SPZhwLGtrend

AuCn_cap

180

180

60

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P1

SPZLGtrend

AuCn_cap

80

75

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

12

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P2

SPZLGtrend

AuCn_cap

160

135

50

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

Kr, AuCN_Cap: Domain Trend P3

SPZLGtrend

AuCn_cap

200

180

60

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

246

Table 11-30: Summary Estimation Parameters (CO3 %)

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier

Restrictions

Drillhole Limit

Estimator Name

Domain

Numeric Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples

per Hole

Apply Drillhole

Limit per Sector

Kr, CO3cap: Domain Trend P1

OgeeHGtrend

CO3cap

75

80

30

VO, Ogee 2026_02

9

20

None

4

TRUE

Kr, CO3cap: Domain Trend P2

OgeeHGtrend

CO3cap

150

160

60

VO, Ogee 2026_02

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

OgeeHGtrend

CO3cap

200

210

105

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P1

OgeeLGtrend

CO3cap

65

100

45

VO, Ogee 2026_02

9

20

None

4

TRUE

Kr, CO3cap: Domain Trend P2

OgeeLGtrend

CO3cap

129

200

90

VO, Ogee 2026_02

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

OgeeLGtrend

CO3cap

195

300

135

VO, Ogee 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P1

OttoHGtrend

CO3cap

125

105

60

VO, Otto_RF 2026_02

9

20

None

4

TRUE

Kr, CO3cap: Domain Trend P2

OttoHGtrend

CO3cap

190

160

90

VO, Otto_RF 2026_02

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

OttoHGtrend

CO3cap

250

210

120

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P1

OttoLGtrend

CO3cap

145

65

55

VO, Otto_RF 2026_02

9

20

None

4

TRUE

Kr, CO3cap: Domain Trend P2

OttoLGtrend

CO3cap

290

130

110

VO, Otto_RF 2026_02

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

OttoLGtrend

CO3cap

390

170

150

VO, Otto_RF 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P1

SPZHGtrend

CO3cap

85

65

40

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

20

None

4

TRUE

Kr, CO3cap: Domain Trend P2

SPZHGtrend

CO3cap

150

145

45

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

SPZHGtrend

CO3cap

300

285

90

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P1

SPZhwHGtrend

CO3cap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P2

SPZhwHGtrend

CO3cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P3

SPZhwHGtrend

CO3cap

180

180

60

VO, SPZhw 2026_02

4

12

None

4

TRUE

Kr, CO3cap: Domain Trend P1

SPZhwLGtrend

CO3cap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P2

SPZhwLGtrend

CO3cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P3

SPZhwLGtrend

CO3cap

180

180

60

VO, SPZhw 2026_02

4

12

None

4

TRUE

Kr, CO3cap: Domain Trend P1

SPZLGtrend

CO3cap

70

70

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

Kr, CO3cap: Domain Trend P2

SPZLGtrend

CO3cap

135

135

60

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

Kr, CO3cap: Domain Trend P3

SPZLGtrend

CO3cap

270

270

100

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

247

Table 11-31: Summary Estimation Parameters (Preg

Rob %)

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier

Restrictions

Drillhole Limit

Estimator

Name

Domain

Numeric

Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples

per Hole

Apply Drillhole

Limit per Sector

ID, PregRob_Cap: Domain Trend P1

OgeeHG trend

PregRob_Cap

50

60

30

VO, Ogee 2026_02

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

OgeeHG trend

PregRob_Cap

100

115

60

VO, Ogee 2026_02

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

OgeeHG trend

PregRob_Cap

210

210

90

VO, Ogee 2026_02

5

16

None

4

TRUE

ID, PregRob_Cap: Domain Trend P1

OgeeLG trend

PregRob_Cap

50

45

30

VO, Ogee 2026_02

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

OgeeLG trend

PregRob_Cap

100

95

60

VO, Ogee 2026_02

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

OgeeLG trend

PregRob_Cap

200

190

120

VO, Ogee 2026_02

5

16

None

4

TRUE

ID, PregRob_Cap: Domain Trend P1

OttoHG trend

PregRob_Cap

80

75

30

VO, Otto_RF 2026_02

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

OttoHG trend

PregRob_Cap

155

150

60

VO, Otto_RF 2026_02

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

OttoHG trend

PregRob_Cap

235

225

90

VO, Otto_RF 2026_02

5

16

None

4

TRUE

ID, PregRob_Cap: Domain Trend P1

OttoLG trend

PregRob_Cap

75

85

30

VO, Otto_RF 2026_02

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

OttoLG trend

PregRob_Cap

156

195

60

VO, Otto_RF 2026_02

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

OttoLG trend

PregRob_Cap

210

250

90

VO, Otto_RF 2026_02

5

16

None

4

TRUE

ID, PregRob_Cap: Domain Trend P1

SPZHG trend

PregRob_Cap

65

65

35

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

SPZHG trend

PregRob_Cap

130

130

55

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

SPZHG trend

PregRob_Cap

260

260

90

VO, SPZ 2026_02

Exclude

NE_12_Offset

5

16

None

4

TRUE

ID, PregRob_Cap: Domain Trend P1

SPZhwHG trend

PregRob_Cap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

SPZhwHG trend

PregRob_Cap

90

90

30

VO, SPZhw 2026_02

4

12

None

3

TRUE

ID, PregRob_Cap: Domain Trend P3

SPZhwHG trend

PregRob_Cap

180

180

60

VO, SPZhw 2026_02

4

12

None

3

TRUE

ID, PregRob_Cap: Domain Trend P1

SPZhwLG trend

PregRob_Cap

60

60

30

VO, SPZhw 2026_02

5

12

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

SPZhwLG trend

PregRob_Cap

90

90

30

VO, SPZhw 2026_02

4

12

None

3

TRUE

ID, PregRob_Cap: Domain Trend P3

SPZhwLG trend

PregRob_Cap

180

180

60

VO, SPZhw 2026_02

2

12

None

3

TRUE

ID, PregRob_Cap: Domain Trend P1

SPZLG trend

PregRob_Cap

80

65

20

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

20

None

4

TRUE

ID, PregRob_Cap: Domain Trend P2

SPZLG trend

PregRob_Cap

120

100

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

ID, PregRob_Cap: Domain Trend P3

SPZLG trend

PregRob_Cap

160

130

60

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

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Table 11-32: Summary Estimation Parameters (Total

Organic Carbon, TOC %)

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier

Restrictions

Drillhole Limit

Estimator Name

Domain

Numeric Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples

per Hole

Apply Drillhole

Limit per Sector

Kr, TOC_pct_D: Domain Trend P1

OgeeHG trend

TOCcap

75

50

30

VO, Ogee 2026_02

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

OgeeHG trend

TOCcap

150

75

50

VO, Ogee 2026_02

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

OgeeHG trend

TOCcap

200

100

70

VO, Ogee 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

OgeeLG trend

TOCcap

80

80

60

VO, Ogee 2026_02

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

OgeeLG trend

TOCcap

160

160

90

VO, Ogee 2026_02

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

OgeeLG trend

TOCcap

200

200

155

VO, Ogee 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

OttoHG trend

TOCcap

70

50

30

VO, Otto_RF 2026_02

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

OttoHG trend

TOCcap

140

105

65

VO, Otto_RF 2026_02

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

OttoHG trend

TOCcap

280

175

80

VO, Otto_RF 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

OttoLG trend

TOCcap

55

30

20

VO, Otto_RF 2026_02

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

OttoLG trend

TOCcap

135

85

85

VO, Otto_RF 2026_02

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

OttoLG trend

TOCcap

180

110

110

VO, Otto_RF 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

SPZHG trend

TOCcap

85

55

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

SPZHG trend

TOCcap

165

105

45

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

SPZHG trend

TOCcap

330

210

60

VO, SPZ 2026_02 Exclude

NE_12_Offset

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

SPZhwHG trend

TOCcap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

SPZhwHG trend

TOCcap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

SPZhwHG trend

TOCcap

180

180

60

VO, SPZhw 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

SPZhwLG trend

TOCcap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

SPZhwLG trend

TOCcap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

SPZhwLG trend

TOCcap

180

180

60

VO, SPZhw 2026_02

4

16

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P1

SPZLG trend

TOCcap

75

75

75

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P2

SPZLG trend

TOCcap

170

170

170

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

Kr, TOC_pct_D: Domain Trend P3

SPZLG trend

TOCcap

370

370

370

VO, SPZ 2026_02 Exclude

NE_12_Offset

4

16

None

4

TRUE

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Table 11-33: Summary Estimation Parameters

(Sulfide Sulfur %)

General

Ellipsoid Ranges

Ellipsoid Orientation

Number of Samples

Outlier

Restrictions

Drillhole Limit

Estimator Name

Domain

Numeric Values

Maximum

Intermediate

Minimum

Variable Orientation

Minimum

Maximum

Method

Max Samples

per Hole

Apply Drillhole

Limit per Sector

ID, SulfideSulfur_pct_D: Domain Trend P1

OgeeHGtrend

SulfSulf_Cap

60

30

50

VO, Ogee 2026_02

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

OgeeHGtrend

SulfSulf_Cap

115

60

95

VO, Ogee 2026_02

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

OgeeHGtrend

SulfSulf_Cap

230

115

190

VO, Ogee 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

OgeeLGtrend

SulfSulf_Cap

75

35

35

VO, Ogee 2026_02

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

OgeeLGtrend

SulfSulf_Cap

130

75

75

VO, Ogee 2026_02

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

OgeeLGtrend

SulfSulf_Cap

275

150

145

VO, Ogee 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

OttoHGtrend

SulfSulf_Cap

90

80

30

VO, Otto_RF 2026_02

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

OttoHGtrend

SulfSulf_Cap

180

130

60

VO, Otto_RF 2026_02

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

OttoHGtrend

SulfSulf_Cap

270

195

75

VO, Otto_RF 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

OttoLGtrend

SulfSulf_Cap

60

85

30

VO, Otto_RF 2026_02

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

OttoLGtrend

SulfSulf_Cap

115

170

50

VO, Otto_RF 2026_02

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

OttoLGtrend

SulfSulf_Cap

150

230

65

VO, Otto_RF 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

SPZHGtrend

SulfSulf_Cap

80

50

30

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

SPZHGtrend

SulfSulf_Cap

165

105

50

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

SPZHGtrend

SulfSulf_Cap

250

160

75

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

SPZhwHGtrend

SulfSulf_Cap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

SPZhwHGtrend

SulfSulf_Cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

SPZhwHGtrend

SulfSulf_Cap

180

180

60

VO, SPZhw 2026_02

4

12

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

SPZhwLGtrend

SulfSulf_Cap

60

60

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

SPZhwLGtrend

SulfSulf_Cap

90

90

30

VO, SPZhw 2026_02

5

16

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

SPZhwLGtrend

SulfSulf_Cap

180

180

60

VO, SPZhw 2026_02

4

12

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P1

SPZLGtrend

SulfSulf_Cap

45

45

45

VO, SPZ 2026_02 Exclude

NE_12_Offset

9

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P2

SPZLGtrend

SulfSulf_Cap

170

170

170

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

24

None

4

TRUE

ID, SulfideSulfur_pct_D: Domain Trend P3

SPZLGtrend

SulfSulf_Cap

240

240

240

VO, SPZ 2026_02 Exclude

NE_12_Offset

5

16

None

4

TRUE

Source: SRK, 2026

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11.2.11  Density

The drill database contains 419 density measurements collected from drilling and by the mine operations of the Granite Creek underground

Project. Density values are assigned to blocks based on the lithological unit. These values are summarized by lithology in Table 11-34.

Table 11-34: Density Values Used in the Underground Model

Lithologic Unit

Density

(tons/

ft3)

Density

(g/cm3)

Qal

0.0578

1.852

Cp

0.0826

2.646

Ocl

0.0820

2.627

Ocu

0.0814

2.608

Kgd

0.0819

2.624

Other

0.0820

2.628

Source: SRK, 2026

11.2.12  Model Validation

SRK has

validated the block model using the following techniques:

•

Visual inspection of block grades in comparison with informing drillhole sample data

•

Statistical comparison of estimation methods to raw sample data and Nearest Neighbor (NN) estimates

•

Sectional validation of the mean sample grades in comparison to the mean model grades

Visual Comparison

Visual inspection of block grades in comparison with informing drillhole sample data are shown in Figure

11-51 to Figure 11-56, for each domain. Overall, SRK considers the visual validation to be reasonable. The visual comparison of the 1.53 m (5 ft) composite to the

estimated block grades shows no obvious bias with a fair reflection of both the high and low-grade samples. Given the variable sample coverage SRK also inspected the grade continuity compared to the grade

control data which showed reasonable correlations for both high and low-grade areas. SRK completed the analysis on a

domain-by-domain level with additional focus being made to the boundary between the HG and LG domains, which when estimated using a hard contact was felt to produce a

sharp transition when a more gradational transition was assumed, thus 1.53 m (5 ft) soft boundaries were selected for the HG and LG domains.

September 2026

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Source: SRK, 2026

Figure 11-51: Plan Section Showing Block Grades vs. Composites (Ogee)

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Source: SRK, 2026

Figure 11-52: Cross Section Showing Block Grades vs. Composites (Ogee)

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Source: SRK, 2026

Figure 11-53: Level Plan (Map) Showing Block Grades vs. Composites (Otto)

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Source: SRK, 2026

Figure 11-54: Cross Section Showing Block Grades vs. Composites (Otto)

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Source: SRK, 2026

Figure 11-55: Plan Showing Block Grades vs. Composites (SPZ)

September 2026

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Source: SRK, 2026

Figure 11-56: Cross Section Showing Block Grades vs. Composites (SPZ)

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Statistical Analysis Comparison

Statistical analysis has been completed using a zero g/t cut-off as a global bias check. The results

show in general strong correlations in the mean grades between the Ordinary Kriging estimates, the declustered estimates and to some extent the NN estimates (Table 11-35). The majority of the mineralization is

contained within the high-grade domains, so SRK has focused its review on the high-grade domains with the results reporting <5% difference between the declustered (30 ft x 30 ft x 15 ft) composites and the block grade estimates. SRK noted a

slight improvement in the soft boundary results in the low-grade domains at Ogee and Otto, with the largest change in the estimates from the selection of the soft boundary occurring within the SPZ. SRK

highlights to the reader that the majority of the low-grade domains are in the order of 1.8 to 2.5 g/t (0.073 oz/st)which is significantly lower than the mining cut-off

grade and therefore has only been considered during the assignment of dilution, and thus, SRK does not consider these differences to be material.

SRK highlights there are some differences in the SPZhw domain which are notable, but this domain has a relatively low tonnage and high-grades,

therefore SRK has considered the lower confidence in these estimates during the classification process.

September 2026

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Table 11-35: Summary Statistics of Raw and

Declustered Composite Grades vs. Block Estimates and NN-Assigned Values (0 g/t Cut-off)

Domain

Statistic

Composite

Au

(g/t)

Declustered Sample Data

Au

(g/t)

Block Estimate

Au(g/t)

Soft Boundary

Block

Estimate

Au(g/t)

Hard

Boundary

Block Estimate

NN Au (g/t)

Hard

Boundary

OgeeHG

Points

2,482.00

2,482.00

299,379.00

299,907.00

300,100.00

Mean

10.66

10.00

9.41

9.39

10.25

Std Dev

14.63

13.71

6.57

6.58

12.35

Variance

214.10

188.06

43.18

43.24

152.49

CV

1.37

1.37

0.70

0.70

1.21

Maximum

70.00

70.00

44.05

44.05

70.00

Minimum

0.00

0.00

0.02

0.00

0.00

Diff % (Raw vs. Estimate)

-11.8

%

-11.9

%

-3.9

%

Diff % (Declust vs. Estimate)

-6.0

%

-6.1

%

2.4

%

OgeeLG

Points

6,095.00

6,095.00

545,613.00

547,552.00

547,175.00

Mean

2.95

2.56

2.32

2.02

2.08

Std Dev

7.33

6.46

2.59

1.94

3.85

Variance

53.80

41.78

6.72

3.77

14.83

CV

2.49

2.52

1.12

0.96

1.85

Maximum

70.00

70.00

36.81

16.87

24.00

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-21.2

%

-31.5

%

-29.5

%

Diff % (Declust vs. Estimate)

-9.5

%

-21.3

%

-19.0

%

OttoHG

Points

1,982.00

1,982.00

438,647.00

439,346.00

438,771.00

Mean

8.12

7.69

7.71

7.71

7.74

Std Dev

10.39

10.20

4.24

4.24

7.61

Variance

108.03

103.95

17.95

17.99

57.94

CV

1.28

1.33

0.55

0.55

0.98

Maximum

62.00

62.00

35.17

35.17

62.00

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-5.0

%

-5.1

%

-4.7

%

Diff % (Declust vs. Estimate)

0.3

%

0.2

%

0.6

%

OttoLG

Points

6,214.00

6,214.00

998,700.00

1,011,333.00

1,011,110.00

Mean

2.43

2.19

2.26

1.81

1.87

Std Dev

5.38

5.08

2.23

1.75

3.06

Variance

28.91

25.80

4.98

3.06

9.35

CV

2.21

2.32

0.99

0.97

1.64

Maximum

62.00

62.00

42.86

18.62

32.00

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-7.2

%

-25.6

%

-23.3

%

Diff % (Declust vs. Estimate)

3.0

%

-17.4

%

-14.9

%

SPZHG

Points

977.00

977.00

564,489.00

564,489.00

564,489.00

Mean

8.96

8.66

8.56

8.56

8.18

Std Dev

11.23

11.11

4.93

4.93

8.49

Variance

126.02

123.45

24.30

24.30

72.09

CV

1.25

1.28

0.58

0.58

1.04

Maximum

52.50

52.50

40.29

40.29

52.50

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-4.4

%

-4.4

%

-8.7

%

Diff % (Declust vs. Estimate)

-1.1

%

-1.1

%

-5.5

%

SPZLG

Points

1,580.00

1,580.00

591,655.00

591,655.00

587,689.00

Mean

2.67

2.49

2.31

1.30

1.41

Std Dev

6.24

5.90

2.34

1.33

2.61

Variance

38.94

34.81

5.47

1.76

6.82

CV

2.34

2.37

1.01

1.02

1.85

Maximum

52.50

52.50

22.75

13.81

20.00

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-13.4

%

-51.2

%

-47.2

%

Diff % (Declust vs. Estimate)

-7.1

%

-47.6

%

-43.3

%

SPZhwHG

Points

205.00

205.00

50,323.00

50,323.00

50,323.00

Mean

11.85

9.08

9.69

11.49

12.01

Std Dev

15.55

13.95

7.55

10.25

11.37

Variance

241.81

194.65

56.95

105.03

129.25

CV

1.31

1.54

0.78

0.89

0.95

Maximum

65.00

65.00

42.96

54.86

65.00

Minimum

0.00

0.00

0.00

0.00

0.00

Diff % (Raw vs. Estimate)

-18.2

%

-3.0

%

1.3

%

Diff % (Declust vs. Estimate)

6.7

%

26.5

%

32.2

%

SPZhwLG

Points

411.00

411.00

123,673.00

123,399.00

123,673.00

Mean

1.54

1.40

1.65

1.34

1.23

Std Dev

2.88

2.89

1.11

0.89

0.95

Variance

8.30

8.34

1.23

0.79

0.89

CV

1.88

2.06

0.67

0.66

0.77

Maximum

26.06

26.06

10.74

6.32

5.16

Minimum

0.00

0.00

0.00

0.00

0.02

Diff % (Raw vs. Estimate)

7.3

%

-13.0

%

-20.1

%

Diff % (Declust vs. Estimate)

17.4

%

-4.9

%

-12.6

%

Source: SRK, 2026

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Swath Plots

A more local comparison between the blocks and the informing composites is made using swath plots. The comparisons show both the varying means

of the block and composites (declustered) along swaths or slices through the model, as well as the amount of data supporting the estimate in each swath. The swath plots show that there are no significant local biases in the estimation. Analysis was

completed on a domain-by-domain level for gold and all the other elements. Only gold results have been shown in Figure 11-57

through Figure 11-59 for the high-grade domains which carry the majority of the value for the Project. Overall, it is the QP’s opinion that the swath plots show similar trends for the three different

estimation methods with a close correlation between the NN and the estimated grades per swath. The declustered composite means show more variation which is expected due to a difference in sample support, but the peaks and troughs of the charts

display a reasonable correlation and therefore in the QP’s opinion are acceptable for the purpose of validation.

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Source: SRK, 2026

Figure 11-57: Swath Plot Analysis of Au (g/t), Ogee HG Domain

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Source: SRK, 2026

Figure 11-58: Swath Plot Analysis of Au (g/t), Otto HG Domain

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Source: SRK, 2026

Figure 11-59: Swath Plot Analysis of Au (g/t), SPZ HG Domain

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11.2.13  Classification

Block model quantities and grade estimates for the Project were classified according to the Society for Mining, Metallurgy and Exploration

(2017).

Mineral Resource classification is typically a subjective concept. Industry best practices suggest that classification should

consider the confidence in the geological continuity of the mineralized structures, the quality and quantity of data supporting the estimates, and the geostatistical confidence in the tonnage and grade estimates. Appropriate classification criteria

should aim to integrate both concepts to delineate regular areas at similar resource classification.

Data quality, drillhole spacing, and

the interpreted continuity of grades controlled by the structures and high-grade domains allowed SRK to classify portions of the Project into Measured, Indicated, and Inferred Mineral Resources categories. SRK has based the current classification on

a review of the variograms, statistical support to the confidence of the estimates (Estimation Variance and Slope of Regression), number of composites and number of holes used to define each estimate:

•

Measured Mineral Resources: Measured blocks have been focused on drilling coverage and proximity to the

current workings within the drilling completed by i-80 averaging a 9.1 m x 9.1 m (30 ft x 30 ft) grid, with three drillholes and three composites. Using the equation:

•

[AuOK_trend: NDh]>=3 and [3 Holes: AvgD]<30 = Measured

•

Indicated Mineral Resources: Limited to drilling coverage within a 27.4 m x 27.4 m (90 ft x 90 ft) grid

completed and with a minimum three holes. Using the equation:

•

[AuOK_trend: NDh]>=3 and [3 Holes: AvgD]<90 = Indicated

•

Inferred Mineral Resources: All other material within the key modeled domains (which limits of 91 m (300 ft) were

applied to the range during definition).

The classification criteria was initially applied and based on study of the

variogram ranges, the QP then undertook a visual review of the grade and geological continuity to assess the level of consistency before applying the final assessment of confidence.

To achieve this the initial mathematical definitions have been completed, they are stored in the block model as Class1_Pass1, which have then

been converted into point data to allow for smoothing of the shapes to avoid block by block assignment. The generated shapes were created in Leapfrog and used a buffer of 2.3 m (7.5 ft) (half the block size) for Measured and 3 m (10 ft) (for

Indicated). The generated shapes were then reviewed by the QP and only areas deemed to demonstrate continuity were defined in the final classification process. Areas of low continuity were down-graded appropriately.

The final block model was generated by creating a stacked classification “geological” model in Leapfrog, which is overlain on the

grade estimation model for reporting.

A summary of the final classification for the key domains is shown in Figure 11-60 to Figure 11-63.

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Source: SRK, 2026

Figure 11-60: Long Section Showing Final Measured (red/brown) and Indicated Shapes (green) Selected

for Granite Creek

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Source: SRK, 2026

Figure 11-61: Cross Section Showing Final Measured and Indicated Shapes Selected for Granite Creek

(Ogee and Otto Domains)

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Source: SRK, 2026

Figure 11-62: Cross Section Showing Final

Measured and Indicated Shapes Selected for Granite Creek (SPZ Domains)

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Source: SRK, 2026

Figure 11-63: Final Classification Shown

Against Filtered Mineral Resource Blocks

11.2.14

Underground Mineral Resource Statement

SME defines a Mineral Resource as:

“a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality,

and quantity that there are reasonable prospects for economic extraction.”

The reasonable prospects for economic extraction (RPEE)

requirement generally implies that the quantity and grade estimates meet certain economic thresholds and that the Mineral Resources are reported at an appropriate cut-off grade (CoG) taking into account

extraction scenarios and processing recoveries. “In order to meet reasonable prospects for economic extraction” requirement, Granite Creek has been deemed amenable to both open pit (discussed in Section 14.2) and underground mining

(based on a cut and fill mining methodology). Metallurgical test work programs have been used to define dynamic recoveries in the mining models which have also been applied to the RPEE assessment. The current underground mined material is currently

toll treated in an offsite autoclave process.

To determine the potential for RPEE, SRK has used the following key assumptions for costing,

and metallurgical recoveries. Metallurgical recoveries are based on the outcomes of both the historical and

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the most recent metallurgical test work completed by i-80. To define the limits for processing route selection and estimated recoveries, SRK estimated AuCn

(ppm), TOC% (percent), PregRob% (percent), and Sulfide Sulfur % (percent), which were reviewed and used by the mining team in relation to the metallurgical recovery.

SRK worked with i-80 to review the findings of the metallurgical studies which indicated that the

metallurgical recoveries range between 77.6% to 93.4% (average 87.6%) for the autoclave and 60.0% to 87.4% (average 68.9%) for the oxide material. SRK has used an MSO, using Deswick Mining Software to determine the limits for the underground RPEE

(Table 11-36).

The gold price selected for Mineral Resources has been based on review of market

study as presented in section 16 of this report which has been compared to market consensus forecast data through a database to which SRK subscribes, review by the QP of industry peers, plus discussion with

i-80 technical teams. Based on the review SRK is considering the price reasonable over a time period of 10 years. It is the QP’s opinion that this is a reasonable forecast for the time considered and the

current time of the mine being in operation (ramping up production), to cover short-term pricing and future long-term pricing. The point of reference for the reporting of mineral resources is the in-situ

diluted tonnage and grades contained with the defined MSO limits.

Table 11-36: Summary of

Mineral Resource Cut-off and MSO Assumptions

Metric (costs)

Imperial (costs)

Process Method1

Parameter

Value

Unit

Value

Unit

Gold Price

3,000

US$/oz

3,000

US$/oz

TC/RC

1.85

US$/oz

1.85

US$/oz

NSR Royalty

6%

6%

Nevada Excise Tax

0.75%

0.75%

Oxides

Metallurgical Recovery

30.0%-86.5%

30.0%-86.5%

Mining Cost

190.18

US$/t

172.53

US$/st

Process Cost

59.94

US$/t

54.38

US$/st

Shipping Cost

16.04

US$/t

14.55

US$/st

G&A Cost

22.29

US$/t

20.22

US$/st

Total Cost

288.45

US$/t

261.68

US$/st

Stope Cut-off Grade

0.108 - 0.312

oz/st Au

0.108 - 0.312

oz/st Au

Cut-off Grade

3.38 - 9.75

g/t Au

3.38 - 9.75

g/t Au

Sulfide

Metallurgical Recovery

37.2%-97.0%

37.2%-97.0%

Mining Cost

190.18

US$/t

172.53

US$/st

Process Cost

131.13

US$/t

118.96

US$/st

Shipping Cost

16.04

US$/t

14.55

US$/st

G&A Cost

22.29

US$/t

20.22

US$/st

Total Cost

359.64

US$/t

326.26

US$/st

Cut-off Grade

0.120 - 0.314

oz/t Au

0.120 - 0.314

oz/st Au

Cut-off Grade

3.75 - 9.81

g/t Au

3.75 - 9.81

g/t Au

Source: SRK, 2026

$:US dollars

oz/st = troy ounce

per short ton

g/t = grams per metric tonne

1. Dynamic cut-off grade was used for the reserves design based on whether the material is oxide

or sulfide.

The final Mineral Resource for the underground portion of Granite Creek has been reported on a diluted basis within the MSO

defined shapes using the following conditions:

•

GM MSO Final = Resource

•

GM Classification = Measured, Indicated or Inferred

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•

Au (g/t) > 0.00 g/t (0.000 oz/st)

The Mineral Resource statement for the underground portion of the Granite Creek Project is shown in Table

11-37 on an exclusive basis and as per SK-1300 the QP has reported the Mineral Resources on an Exclusive basis which has been defined by removing all material within the

mine design which was converted to Reserves, including a marginal small portion of Inferred from within the design which was considered as waste within the reserves. Note, for reporting purposes given the proximity and similar mineralization, the

SPZhw has been combined with the SPZ MSO domains for the final reporting.

Table 11-37:

Inclusive Mineral Resource Statement for Granite Creek Underground—Effective Date March 31, 2026

Classification

Domain

Mass

(000’s

tonnes)

Mass

Attributable

(000’s

tonnes)

Au

Grade

(g/t)

Material

Content Au

(000’s t. oz)

Attributable

Material Content

Au

(000’s t. oz)

Measured

Ogee

231

231

7.88

58.6

58.5

Otto

310

310

6.69

66.6

66.6

SPZ

65

65

9.85

20.7

20.7

Total

606

606

7.49

145.9

145.8

Indicated

Ogee

322

321

7.06

73.0

72.9

Otto

746

746

6.26

150.2

150.2

SPZ

2,055

2,055

7.42

490.4

490.4

Total

3,122

3,122

7.11

713.6

713.5

M&I

Ogee

553

552

7.40

131.6

131.4

Otto

1,056

1,056

6.39

216.8

216.8

SPZ

2,120

2,120

7.50

511.1

511.1

Total

3,729

3,728

7.17

859.5

859.3

Inferred

Ogee

39

38

6.70

8.3

8.3

Otto

425

425

7.63

104.4

104.4

SPZ

429

418

6.53

90.1

87.4

Total

893

882

7.06

202.8

200.1

Source: SRK, 2026

Notes:

1. Mineral Resources,

which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, market or other relevant issues. The deposit has been classified as

Measured, Indicated and Inferred based on confidence in the geological model and drill spacing. The quantity and grade of reported Inferred resources are uncertain in nature, and there has not been sufficient work to define these Inferred Mineral

Resources as Indicated or Measured resources. There is no certainty that any part of a mineral resource will ever be converted into reserves.

2. Mineral Resources are reported on an inclusive basis with units reported in in Metric Tonnes, grade is reported in grams/tonnes, and metal

in 000’s troy ounces.

3. Mineral Resources are reported within Mineable Stopes which have been defined based on metal price

assumptions,* variable metallurgical recovery assumptions, mining costs, processing costs, shipping, G&A costs. Metal Pricing is based on Gold (US$3,000 /oz). Other key assumptions include:

* Mining Costs (US$/t 190.18);

*

Autoclave Processing Cost (US$/t 131.13), Low-grade Oxide (US$/t 59.94)

* Autoclave Recoveries

ranging 77.6% - 93.4% (average 87.6%) and Oxide recoveries ranging 60.0% - 87.4% (average 68.9%);

* Shipping Costs (US$/t 16.04) and

G&A Costs (US$/t 22.29);

* Totaling US$/t 281.45 – 359.64 for underground mining and Processing;

* Based on the variable recoveries an average cut-off grade of 3.75—9.81 g/t (0.286 oz/st) has

been used reporting.

4. Mineral Resources have been depleted for previous underground mining activity, and reported on an attributable

basis

5. Mineral Resources are based on validated data, which have been subjected to QA/QC analysis, using capped, composited samples at 2

m (5ft) intervals. Estimation has been completed using a combination of OK and IDW estimation methodologies and classified based on confidence in the underlying data and drill spacing. Mineral resource tonnages have been rounded to reflect the

precision of the estimate.

6. The mineral resources were estimated by SRK Consulting (U.S.), Inc.

7. Rounding of some figures may lead to minor discrepancies in totals.

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Table 11-38: Exclusive Mineral Resource Statement

for Granite Creek Underground—Effective Date March 31, 2026

Classification

Domain

Mass

(000’s

tonnes)

Mass

Attributable

(000’s tonnes)

Au

Grade

(g/t)

Material

Content Au

(000’s t. oz)

Attributable Material

Content Au

(000’s t. oz)

Measured

Ogee

175

175

6.82

38.4

38.3

Otto

166

166

5.58

29.7

29.7

SPZ

31

31

6.44

6.5

6.5

Total

372

372

6.23

74.6

74.5

Indicated

Ogee

241

241

6.52

50.6

50.6

Otto

425

425

5.23

71.5

71.5

SPZ

771

771

4.70

116.6

116.6

Total

1,438

1,438

5.16

238.6

238.6

M&I

Ogee

417

416

6.64

89.0

88.9

Otto

591

591

5.32

101.2

101.2

SPZ

803

803

4.77

123.1

123.1

Total

1,810

1,810

5.38

313.3

313.1

Inferred

Ogee

38

38

6.71

8.3

8.3

Otto

417

417

7.62

102.0

102.0

SPZ

415

405

6.49

86.6

83.9

Total

871

860

7.04

197.0

194.2

Source: SRK, 2026

Notes:

1. Mineral

Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, market or other relevant issues. The deposit has been

classified as Measured, Indicated and Inferred based on confidence in the geological model and drill spacing. The quantity and grade of reported Inferred resources are uncertain in nature, and there has not been sufficient work to define these

Inferred Mineral Resources as Indicated or Measured resources. There is no certainty that any part of a mineral resource will ever be converted into reserves.

2. Mineral Resources are reported on an Exclusive basis with units reported in in Metric Tonnes, grade is reported in grams/tonnes, and metal

in 000’s troy ounces.

3. Mineral Resources are reported fully diluted within Mineable Stopes which have been defined based on metal

price assumptions,* variable metallurgical recovery assumptions, mining costs, processing costs, shipping, G&A costs. Metal Pricing is based on Gold (US$3,000 /oz). Other key assumptions include:

* Mining Costs (US$/t 190.18);

*

Autoclave Processing Cost (US$/t 131.13), Low-grade Oxide (US$/t 59.94)

* Autoclave Recoveries

ranging 77.6% - 93.4% (average 87.6%) and Oxide recoveries ranging 60.0% - 87.4% (average 68.9%);

* Shipping Costs (US$/t 16.04) and

G&A Costs (US$/t 22.29);

* Totaling US$/t 281.45 – 359.64 for underground mining and Processing;

* Based on the variable recoveries an average cut-off grade of 3.75—9.81 g/t (0.286 oz/st) has

been used reporting.

4. Mineral Resources have been depleted for previous underground mining activity, and reported on an attributable

basis

5. Mineral Resources are based on validated data, which have been subjected to QA/QC analysis, using capped, composited samples at 2

m (5ft) intervals. Estimation has been completed using a combination of OK and IDW estimation methodologies and classified based on confidence in the underlying data and drill spacing. Mineral resource tonnages have been rounded to reflect the

precision of the estimate.

6. The mineral resources were estimated by SRK Consulting (U.S.), Inc.

7. Rounding of some figures may lead to minor discrepancies in totals.

11.2.15

Mineral Resource Sensitivity

Table 11-39 Table 11-38 and Table 11-40 show the continuity of the grade estimates at various cut-off increments and the sensitivity of the mineral resource to changes in CoG. The reader should note that the

results presented herein have been limited to the current MSO shape and show limited growth on cut-off below 4.78 g/t (0.139 oz/st) Au, while comparison of in situ tonnage and grades may show greater growth.

The reader is cautioned that the values in Table 14 38 and Table 14 39 should not be misconstrued with the mineral resource statement. The values are only presented to show the sensitivity of the block model estimates to the selection of CoG. All

values are rounded to reflect the relative accuracy of the estimates.

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Table 11-39: Grade Sensitivity (Combined Measured

and Indicated), inside MSO on a 100% Project basis

Cut-Off

Classification

Mass

(000’s tonnes)

Au Grade

(g/t)

Material Content

Au (000’s t. oz)

Au (g/t) ≥ 0.00 g/t

M&I

3,729

7.17

859.5

Au (g/t) ≥ 0.50 g/t

M&I

3,324

8.04

859.0

Au (g/t) ≥ 1.00 g/t

M&I

3,274

8.15

857.8

Au (g/t) ≥ 1.50 g/t

M&I

3,210

8.29

855.2

Au (g/t) ≥ 2.00 g/t

M&I

3,145

8.42

851.5

Au (g/t) ≥ 2.50 g/t

M&I

3,075

8.56

846.5

Au (g/t) ≥ 3.00 g/t

M&I

3,007

8.69

840.5

Au (g/t) ≥ 3.50 g/t

M&I

2,923

8.85

831.7

Au (g/t) ≥ 4.00 g/t

M&I

2,813

9.05

818.4

Au (g/t) ≥ 4.25 g/t

M&I

2,742

9.18

808.9

Au (g/t) ≥ 4.50 g/t

M&I

2,742

9.18

808.9

Au (g/t) ≥ 4.75 g/t

M&I

2,562

9.50

782.7

Au (g/t) ≥ 5.00 g/t

M&I

2,457

9.70

766.4

Au (g/t) ≥ 5.25 g/t

M&I

2,356

9.90

749.7

Au (g/t) ≥ 5.50 g/t

M&I

2,254

10.10

732.1

Au (g/t) ≥ 5.75 g/t

M&I

2,148

10.32

712.8

Au (g/t) ≥ 6.00 g/t

M&I

2,043

10.55

693.0

Au (g/t) ≥ 6.25 g/t

M&I

1,944

10.78

673.5

Au (g/t) ≥ 6.50 g/t

M&I

1,851

11.00

654.5

Au (g/t) ≥ 6.75 g/t

M&I

1,754

11.24

633.7

Au (g/t) ≥ 7.00 g/t

M&I

1,662

11.48

613.5

Au (g/t) ≥ 7.25 g/t

M&I

1,576

11.72

593.7

Au (g/t) ≥ 7.50 g/t

M&I

1,493

11.96

574.0

Au (g/t) ≥ 8.00 g/t

M&I

1,327

12.49

532.7

Au (g/t) ≥ 8.50 g/t

M&I

1,185

13.00

495.1

Source: SRK, 2026

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Table 11-40: Grade Sensitivity Table (Inferred),

inside MSO on a 100% Project Basis

Cut-Off

Classification

Mass

(000’s tonnes)

Au Grade (g/t)

Material Content

Au (000’s t. oz)

Au (g/t) ≥ 0.00 g/t

Inferred

893

7.06

202.8

Au (g/t) ≥ 0.50 g/t

Inferred

727

8.66

202.6

Au (g/t) ≥ 1.00 g/t

Inferred

717

8.78

202.4

Au (g/t) ≥ 1.50 g/t

Inferred

702

8.94

201.8

Au (g/t) ≥ 2.00 g/t

Inferred

684

9.13

200.7

Au (g/t) ≥ 2.50 g/t

Inferred

660

9.38

199.0

Au (g/t) ≥ 3.00 g/t

Inferred

635

9.64

196.8

Au (g/t) ≥ 3.50 g/t

Inferred

612

9.88

194.3

Au (g/t) ≥ 4.00 g/t

Inferred

590

10.11

191.8

Au (g/t) ≥ 4.25 g/t

Inferred

581

10.20

190.5

Au (g/t) ≥ 4.50 g/t

Inferred

581

10.20

190.5

Au (g/t) ≥ 4.75 g/t

Inferred

551

10.51

186.2

Au (g/t) ≥ 5.00 g/t

Inferred

530

10.73

183.0

Au (g/t) ≥ 5.25 g/t

Inferred

512

10.93

180.0

Au (g/t) ≥ 5.50 g/t

Inferred

500

11.07

177.8

Au (g/t) ≥ 5.75 g/t

Inferred

486

11.22

175.3

Au (g/t) ≥ 6.00 g/t

Inferred

471

11.39

172.4

Au (g/t) ≥ 6.25 g/t

Inferred

451

11.63

168.4

Au (g/t) ≥ 6.50 g/t

Inferred

432

11.86

164.6

Au (g/t) ≥ 6.75 g/t

Inferred

420

12.01

162.0

Au (g/t) ≥ 7.00 g/t

Inferred

391

12.39

155.7

Au (g/t) ≥ 7.25 g/t

Inferred

370

12.69

150.8

Au (g/t) ≥ 7.50 g/t

Inferred

360

12.83

148.6

Au (g/t) ≥ 8.00 g/t

Inferred

301

13.82

133.8

Au (g/t) ≥ 8.50 g/t

Inferred

267

14.53

124.7

Source: SRK, 2026

11.2.16

Sources of Uncertainty

Areas of uncertainty that may materially impact the MREs include impacts of the underlying supporting data, geological interpretation,

estimation parameters and economic assumptions:

SRK has considered the various sources of the data and accounted for these in the

QP’s opinion through the estimations methodology and classification systems applied. Areas with lower drilling density or lower continuity either geologically or grade based have been considered when assigning the level of confidence. Some

areas within the existing mining operations have relatively limited drilling cover from the diamond and RC drilling, but in the QP’s opinion the visual reconciliation to the muck and sludge samplings is acceptable to maintain the assigned

confidence.

In the estimation process for this style of deposit there is a degree of sensitivity to the grades related to the selected

capping levels, which the QP has applied based on their years of experience in the style of mineralization. Changes in capping (up or down), would result in higher or lower overall grades, but the QP has selected values based on analysis in two

separate software packages to test the impact on the contained metal at various caps to select the most appropriate values. As in any production environment the grades (and hence capping values), should be monitored through the reconciliation

process to confirm performance of the model.

In terms of the level of study and the available information it is the QP’s opinion

that the level of density sampling remains relatively low and there is some degree of uncertainty in the level of variability in the density within the different units. Reduction in density could result in lower tonnage and contained metal. However,

as this mine is an operating entity with active reconciliation to production it is not considered by the QP at this time that any variability would be material.

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SRK notes that future economic assessment could result in a change in the CoG which would

potentially result in a change in the tonnage of available mineable material. Mineralization represented by the resource block model was evaluated for reasonable prospects of economic extraction for underground mining methods by applying an

appropriate CoG, these can be impacted by:

•

Changes to long term metal price assumptions

•

Changes to the input values for mining, processing, and G&A costs to constrain the estimate

•

Changes to local interpretations of mineralization geometry and continuity of mineralized domains, including the

stratigraphic and structural models

•

Changes in the capping values can impact the sensitivity of the grade estimates and reconciliation data should be

reviewed to ensure consistency with grade distribution

•

Material types (oxide vs sulphide) is based on visual inspection and changes in material type could impact

metallurgical recovery

•

Changes to metallurgical recovery and geo-metallurgical assumptions

•

Changes to assumptions with an existing agreement or new agreements

•

Changes to environmental, permitting, and social license assumptions

It is the QP’s opinion that the risks highlighted above can be reduced through additional exploration and technical studies, but the

Mineral Resources as presented have accounted for the confidence where applicable.

11.2.17

Reconciliation

Production Reconciliation Process

i-80 performs routine reconciliation between the mineral resource (previous model only), underground

production and a short-term block model (STBM). Surveyed mining extents for individual development rounds were used to define the mined volumes, which were then intersected with the block models to extract modeled grades and contained gold within

the exact mined shapes. Recorded production tonnages and associated grade-control assays (assigned to each mining round) were compiled, and tonnage-weighted grades and contained ounces were calculated and compared to the model predictions on a

volume-matched basis.

The QP considers this approach is generally consistent with industry best practice for underground reconciliation as

it is based on surveyed mined volumes, uses tonnage-weighted comparisons, and evaluates performance by domain and through time. However, some implementation details (e.g., exact production grade basis, classification cut-offs, and linkage of the

reconciled model to current Mineral Resource / Reserve statement will need to be to the updated resource and reserves model needs to be completed. The reconciliation to the previous model showed a bias between the earlier model estimates and actual

production. In aggregate, mined grades and contained gold were higher than predicted by both the longer-term model and the STBM, indicating an overall positive reconciliation (i.e., the earlier models underestimated production). This positive bias

was concentrated in high-grade oxide and sulfide material, while some low-grade oxide material reconciled negatively, suggesting that grades in that category were over-estimated.

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i-80 interpreted these results as evidence that the

earlier model was conservative overall, with domain-specific biases that needed to be addressed. In particular, the strong positive reconciliation in high-grade material and negative reconciliation in parts of the low-grade oxide inventory highlight

the importance of domain-specific parameterization, grade capping / clamping strategies, and classification thresholds. These findings were considered in the development and refinement of the current geological and estimation models and in the

assessment of uncertainty and risk in the pre-feasibility-level Mineral Resource and Reserve estimates. The reconciliation is therefore used as a diagnostic tool to inform model improvement, rather than as a basis for current public disclosure of

Mineral Resources and Reserves

Comparison to the Previous Mineral Resources

No Exclusive Resource have previously been reported for the project as this is the first declared Mineral Resources, with the previous report

limited to an initial assessment. Comparison of the inclusive Mineral Resources noted an increase in the Measured and Indicated portion of the deposit from 775 kt at grade of 10.50 g/t Au for 261 koz of contained gold, to 3,729 kt at 7.17 g/t for

859 koz of contained gold. There was a reduction in the Inferred Mineral Resources from 782 kt of material at an average grade of 13.0 g/t for 326 koz, to 893 kt at 7.06 g/t Au for 203 koz of gold.

The differences in the models are attributed to the following key changes:

•

Depletion of mined material to March 31, 2026

•

Infill drilling completed during 2025 targeting high-grade extensions and infill of the PEA model material

•

Increase in the gold price used for the RPEE assessment from $2,175/oz to $3,000/oz

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11.2.18

Relevant Factors

The QP is not aware of any additional Environmental, permitting, legal, title, taxation marketing or other factors that could affect resources.

The final domains were exported to .dxf format and provided to the i-80 Geology team for review;

several small areas needed manual adjustment to improve the geological interpretation (namely in areas of lower sampling volume), which was completed by SRK, with final review and agreement

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Table 11-41: Open Pit Mineral Resource Sensitivity

by Domain

Deposit

Cutoff Grade (g/t)

Total Process Material

(‘000s tonnes)

Total Process Material

(‘000s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(million tr

oz)

Measured

Pit B

0.1

5,131

5,656

0.84

0.025

0.139

0.15

4,399

4,849

0.96

0.028

0.136

0.2

3,900

4,299

1.06

0.031

0.133

0.25

3,501

3,859

1.16

0.034

0.131

0.3

3,153

3,476

1.26

0.037

0.128

0.35

2,856

3,148

1.36

0.040

0.124

0.4

2,629

2,897

1.44

0.042

0.122

0.45

2,391

2,635

1.54

0.045

0.118

0.5

2,187

2,410

1.64

0.048

0.115

Pit A

0.1

1,489

1,641

0.56

0.016

0.027

0.15

1,270

1,400

0.64

0.019

0.026

0.2

1,098

1,211

0.71

0.021

0.025

0.25

915

1,009

0.81

0.024

0.024

0.3

773

852

0.91

0.026

0.023

0.35

638

703

1.03

0.030

0.021

0.4

526

580

1.17

0.034

0.020

0.45

440

485

1.31

0.038

0.019

0.5

370

408

1.47

0.043

0.018

CX

0.1

18,417

20,301

0.85

0.025

0.501

0.15

15,362

16,933

0.99

0.029

0.489

0.2

13,493

14,873

1.10

0.032

0.478

0.25

12,117

13,356

1.20

0.035

0.469

0.3

11,130

12,268

1.29

0.037

0.460

0.35

10,282

11,334

1.36

0.040

0.451

0.4

9,577

10,557

1.44

0.042

0.443

0.45

8,944

9,859

1.51

0.044

0.434

0.5

8,344

9,198

1.58

0.046

0.425

Mag

0.1

18,446

20,333

0.87

0.025

0.515

0.15

16,524

18,214

0.96

0.028

0.508

0.2

15,055

16,596

1.03

0.030

0.500

0.25

13,683

15,083

1.11

0.032

0.490

0.3

12,603

13,893

1.18

0.035

0.480

0.35

11,672

12,866

1.25

0.037

0.470

0.4

10,721

11,818

1.33

0.039

0.459

0.45

9,852

10,860

1.41

0.041

0.447

0.5

9,114

10,046

1.49

0.043

0.436

Indicated

Pit B

0.1

1,491

1,643

0.42

0.012

0.020

0.15

1,098

1,210

0.52

0.015

0.018

0.2

823

907

0.64

0.019

0.017

0.25

587

647

0.80

0.023

0.015

0.3

433

477

0.99

0.029

0.014

0.35

347

382

1.16

0.034

0.013

0.4

279

307

1.35

0.039

0.012

0.45

243

268

1.49

0.043

0.012

0.5

207

229

1.66

0.049

0.011

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Deposit

Cutoff Grade (g/t)

Total Process Material

(‘000s tonnes)

Total Process Material

(‘000s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(million tr

oz)

Pit A

0.1

1,290

1,422

0.54

0.016

0.023

0.15

1,114

1,228

0.61

0.018

0.022

0.2

969

1,069

0.68

0.020

0.021

0.25

863

952

0.73

0.021

0.020

0.3

782

862

0.78

0.023

0.020

0.35

694

765

0.84

0.024

0.019

0.4

605

667

0.90

0.026

0.018

0.45

524

578

0.98

0.029

0.016

0.5

463

510

1.04

0.030

0.016

CX

0.1

6,949

7,660

0.69

0.020

0.154

0.15

5,406

5,959

0.85

0.025

0.147

0.2

4,423

4,876

1.00

0.029

0.142

0.25

3,755

4,139

1.14

0.033

0.137

0.3

3,364

3,708

1.24

0.036

0.134

0.35

3,102

3,419

1.31

0.038

0.131

0.4

2,875

3,169

1.39

0.040

0.128

0.45

2,652

2,924

1.47

0.043

0.125

0.5

2,482

2,736

1.54

0.045

0.123

Mag

0.1

15,095

16,640

0.60

0.018

0.292

0.15

13,067

14,403

0.68

0.020

0.284

0.2

11,509

12,687

0.74

0.022

0.275

0.25

10,179

11,221

0.81

0.024

0.265

0.3

9,029

9,953

0.88

0.026

0.255

0.35

8,018

8,838

0.95

0.028

0.245

0.4

7,160

7,893

1.02

0.030

0.234

0.45

6,245

6,884

1.11

0.032

0.222

0.5

5,416

5,970

1.20

0.035

0.209

Inferred

Pit B

0.1

72

79

0.40

0.012

0.001

0.15

52

57

0.51

0.015

0.001

0.2

50

55

0.52

0.015

0.001

0.25

45

50

0.55

0.016

0.001

0.3

39

43

0.60

0.017

0.001

0.35

30

33

0.67

0.020

0.001

0.4

24

27

0.75

0.022

0.001

0.45

23

26

0.76

0.022

0.001

0.5

21

23

0.80

0.023

0.001

Pit A

0.1

615

678

0.35

0.010

0.007

0.15

506

558

0.40

0.012

0.007

0.2

440

485

0.44

0.013

0.006

0.25

352

388

0.49

0.014

0.006

0.3

275

304

0.55

0.016

0.005

0.35

229

253

0.60

0.018

0.004

0.4

188

207

0.65

0.019

0.004

0.45

154

170

0.70

0.020

0.003

0.5

124

137

0.76

0.022

0.003

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Deposit

Cutoff Grade (g/t)

Total Process Material

(‘000s tonnes)

Total Process Material

(‘000s short tons)

Au Grade

(g/t)

Au Grade

(opt)

Au Contained

(million tr

oz)

CX

0.1

3,155

3,478

0.65

0.019

0.066

0.15

2,464

2,716

0.80

0.023

0.063

0.2

1,997

2,201

0.94

0.028

0.061

0.25

1,734

1,912

1.05

0.031

0.059

0.3

1,574

1,735

1.13

0.033

0.057

0.35

1,426

1,572

1.22

0.036

0.056

0.4

1,316

1,451

1.29

0.038

0.055

0.45

1,232

1,358

1.35

0.039

0.053

0.5

1,146

1,263

1.41

0.041

0.052

Mag

0.1

1,901

2,095

0.57

0.017

0.035

0.15

1,594

1,757

0.65

0.019

0.034

0.2

1,339

1,476

0.75

0.022

0.032

0.25

1,036

1,142

0.90

0.026

0.030

0.3

783

863

1.10

0.032

0.028

0.35

699

770

1.20

0.035

0.027

0.4

635

700

1.28

0.037

0.026

0.45

612

675

1.31

0.038

0.026

0.5

602

664

1.32

0.039

0.026

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Table 11-42: Underground Mineral Resource

Sensitivity (1 of 2)

Cut-Off

Classification

Mass

(Mt)

Average Value

Au (g/t)

Material Content

Au (g/t)

thousand t. oz

Au (g/t) ≥ 0.00 g/t

M&I

2.26

9.15

666.41

Au (g/t) ≥ 0.50 g/t

M&I

2.26

9.19

666.32

Au (g/t) ≥ 1.00 g/t

M&I

2.25

9.22

666.10

Au (g/t) ≥ 1.50 g/t

M&I

2.23

9.27

665.59

Au (g/t) ≥ 2.00 g/t

M&I

2.22

9.31

664.87

Au (g/t) ≥ 2.50 g/t

M&I

2.20

9.36

663.69

Au (g/t) ≥ 3.00 g/t

M&I

2.19

9.41

662.29

Au (g/t) ≥ 3.50 g/t

M&I

2.17

9.48

659.89

Au (g/t) ≥ 4.00 g/t

M&I

2.13

9.56

655.87

Au (g/t) ≥ 4.25 g/t

M&I

2.11

9.63

652.51

Au (g/t) ≥ 4.50 g/t

M&I

2.07

9.72

647.73

Au (g/t) ≥ 4.75 g/t

M&I

2.02

9.85

639.66

Au (g/t) ≥ 5.00 g/t

M&I

1.96

10.01

629.79

Au (g/t) ≥ 5.25 g/t

M&I

1.89

10.18

618.94

Au (g/t) ≥ 5.50 g/t

M&I

1.82

10.36

607.06

Au (g/t) ≥ 5.75 g/t

M&I

1.75

10.56

594.00

Au (g/t) ≥ 6.00 g/t

M&I

1.68

10.77

579.93

Au (g/t) ≥ 6.25 g/t

M&I

1.60

10.97

565.71

Au (g/t) ≥ 6.50 g/t

M&I

1.53

11.19

551.19

Au (g/t) ≥ 6.75 g/t

M&I

1.46

11.42

535.16

Au (g/t) ≥ 7.00 g/t

M&I

1.39

11.65

519.99

Au (g/t) ≥ 7.25 g/t

M&I

1.32

11.87

505.04

Au (g/t) ≥ 7.50 g/t

M&I

1.26

12.11

489.03

Au (g/t) ≥ 8.00 g/t

M&I

1.12

12.63

455.89

Au (g/t) ≥ 8.50 g/t

M&I

1.01

13.12

425.89

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Table 11-43: Underground Mineral Resource

Sensitivity (2 of 2)

Cut-Off

Classification

Mass

(Mt)

Average Value

Au (g/t)

Material Content

Au (g/t)

thousand t. oz

Au (g/t) ≥ 0.00 g/t

Inferred

0.44

9.88

138.37

Au (g/t) ≥ 0.50 g/t

Inferred

0.43

9.98

138.35

Au (g/t) ≥ 1.00 g/t

Inferred

0.43

10.03

138.30

Au (g/t) ≥ 1.50 g/t

Inferred

0.43

10.08

138.20

Au (g/t) ≥ 2.00 g/t

Inferred

0.42

10.16

137.96

Au (g/t) ≥ 2.50 g/t

Inferred

0.42

10.29

137.47

Au (g/t) ≥ 3.00 g/t

Inferred

0.41

10.41

136.92

Au (g/t) ≥ 3.50 g/t

Inferred

0.40

10.53

136.21

Au (g/t) ≥ 4.00 g/t

Inferred

0.40

10.64

135.41

Au (g/t) ≥ 4.25 g/t

Inferred

0.39

10.70

134.94

Au (g/t) ≥ 4.50 g/t

Inferred

0.39

10.78

134.26

Au (g/t) ≥ 4.75 g/t

Inferred

0.38

10.95

132.68

Au (g/t) ≥ 5.00 g/t

Inferred

0.37

11.13

131.01

Au (g/t) ≥ 5.25 g/t

Inferred

0.36

11.30

129.31

Au (g/t) ≥ 5.50 g/t

Inferred

0.35

11.44

127.91

Au (g/t) ≥ 5.75 g/t

Inferred

0.34

11.58

126.39

Au (g/t) ≥ 6.00 g/t

Inferred

0.33

11.76

124.49

Au (g/t) ≥ 6.25 g/t

Inferred

0.31

12.01

121.63

Au (g/t) ≥ 6.50 g/t

Inferred

0.30

12.27

118.82

Au (g/t) ≥ 6.75 g/t

Inferred

0.29

12.41

117.36

Au (g/t) ≥ 7.00 g/t

Inferred

0.28

12.66

114.47

Au (g/t) ≥ 7.25 g/t

Inferred

0.27

12.94

111.41

Au (g/t) ≥ 7.50 g/t

Inferred

0.26

13.07

110.02

Au (g/t) ≥ 8.00 g/t

Inferred

0.22

14.14

98.98

Au (g/t) ≥ 8.50 g/t

Inferred

0.20

14.70

93.94

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12

Mineral Reserve Estimates

Underground Mineral Reserves for the Granite Creek Project were estimated by SRK Consulting, for the Ogee, Otto, and SP zones located beneath

the historically mined open pit. The open pit Mineral Resource (Section 11) has no corresponding Mineral Reserve, no mineral reserves have been estimated for the open pit as part of this study as further technical work is needed to increase the

confidence in the modifying factors. Granite Creek is currently in production and is mined using an underhand drift-and-fill (DAF) method, with stopes backfilled

primarily using cemented waste rock backfill (CRF). The general location of the deposit is shown in Figure 12-1.

Source: SRK, 2026

Figure 12-1: General Locations of the Granite Creek Deposit

12.1

Key Assumptions, Parameters, and Methods Used

Potential mining blocks are generated from the resource block model based on the deposit geometry, metallurgical process, and block values.

Measured Mineral Resources were converted to Proven Mineral Reserves and Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the modifying factors described below; Inferred Mineral Resources were not converted to

mineral reserves, where Inferred material is contained within the mineral reserve designs, it was assigned zero grade and treated as internal dilution.

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Mining Dilution

Dilution in the mineral reserves consists of internal and external dilution. Internal (planned) dilution depends on the geometry of the deposit

and the minimum mining width: where the width of above-cut-off-grade mineralization is less than the 4.5 m (15 ft) minimum mining

width, a limited amount of below-cut-off material is mined to achieve that minimum width, and this is treated as internal dilution. External dilution consists of below-cut-off material that overbreaks into the designed stope due to the geotechnical properties of the adjacent rock (from the host rock or from backfill in adjacent

drifts); an external dilution of 10% was applied to the drift-and-fill production shapes at zero grade.

Mining Recovery

A

stope recovery factor of 100% was applied to the reserves, determined based on what current mining has achieved (the drifts tend to overbreak, and the planned material is recovered). A development recovery factor of 100% was used for all lateral

development.

Mineral Reserves Estimates

The gold price selected for Mineral Reserves has been based on review of market study as presented in section 16 of this report which has been

compared to market consensus forecast data through a database to which SRK subscribes, review by the QP of industry peers, plus discussion with i-80 technical teams. Based on the review SRK is considering the

price reasonable over a time period of 10 years. It is the QP’s opinion that this is a reasonable forecast for the time considered and the current time of the mine being in operation (ramping up production), to cover short-term pricing and

future long-term pricing. A 3D mine design has been created representing the planned reserve mining areas for the Granite Creek underground operation. The point of reference for the reporting of mineral reserves is the

in-situ diluted tonnage and grades contained within the Mine Design, which accounts for the modifying factors.

The design process resulted in 2.20 Mt (2.42 Mst) of Proven and Probable Mineral Reserves at an average grade of 7.87 g/t (0.23 oz/st) Au, for

all commodities gold only (no other commodity is estimated — see Section 12.5). The cut-off grades and metallurgical recoveries underlying the estimate are discussed in Section 12.3. Table 12-1 presents the Mineral Reserve statement as of March 31, 2026.

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Table 12-1: Mineral Reserve Estimate for Granite

Creek Underground – Effective Date March 31, 2026

Classification

Domain

Metric Units

Imperial Units

Tonnage

(000’ tonnes)

Au Grade (g/t)

Tonnage

(000’s s. tons)

Au Grade (oz/st)

Au Metal Content

(000’s t. oz)

Proven

Ogee

71

9.34

78

0.27

21.2

Otto

171

7.04

189

0.21

38.7

SP

40

11.21

44

0.33

14.5

Stockpile

26

7.33

29

0.21

6.2

Total

308

8.13

340

0.24

80.6

Probable

Ogee

88

7.95

97

0.23

22.4

Otto

366

6.70

403

0.20

78.8

SP

1,437

8.11

1,584

0.24

374.7

Total

1,890

7.83

2,084

0.23

475.9

Proven + Probable

Ogee

158

8.57

174

0.25

43.6

Otto

537

6.81

592

0.20

117.6

SP

1,477

8.19

1,628

0.24

389.2

Stockpile

26

7.33

29

0.21

6.2

Total

2,198

7.87

2,423

0.23

556.5

Source: SRK, 2026

Notes:

•

All figures are rounded to reflect the relative accuracy of the estimates. Totals may not sum due to rounding.

Mineral Reserves have been stated on the basis of a mine design, mine plan, and economic model.

•

Reserves are reported using a stope CoG range of 4.42 g/t (0.13 oz/st) Au – 6.41 g/t (0.19 oz/st) Au and an

incremental processing CoG range of 1.51 g/t (0.04 oz/st) Au – 2.91 g/t (0.08 oz/st) Au depending on the processing method. The CoG calculations assume a US$2,500/oz Au price, and 60% to 92% metallurgical recovery depending on the processing

method applied.

•

Operating costs include mining (US$190.18/tonne), processing (US$59.94/tonne to US$131.13/tonne), G&A

(US$22.29/tonne), Shipping Costs (US$/t 16.04). Additionally, a 6% royalty and 0.75% Excise Tax are applied, along with TC/RC charges of US$1.85/oz Au.

•

The reserves are mined using the underhand drift and fill method with cemented waste rock backfill. Mining

dilution of 10% is included in the reserve at zero grade.

•

Stockpile reserves reflect ore mined and placed on surface stockpiles prior to the effective date and not yet

processed.

•

The Mineral Reserves were estimated by SRK Consulting, U.S. (Inc.).

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All figures are rounded to reflect the relative accuracy of the estimates; totals may not sum

due to rounding. Mineral Reserves are stated on the basis of a mine design, mine plan, and economic model, using the underhand drift-and-fill method with cemented waste

rock backfill and a mining dilution of 10% included in the reserve at zero grade. Stockpile reserves reflect ore mined and placed on surface stockpiles prior to the effective date and not yet processed. The Mineral Reserves were estimated by SRK.

The qualified person preparing the mineral resource estimates must round off, to appropriate significant figures chosen to reflect order

of accuracy, any estimates of quantity and grade or quality.

12.2

Cut-Off Grades Estimates

Current estimated project costs and the calculated economic cut-off grade (CoG) for the underground

Mineral Reserve are shown in Table 12-2. A dynamic cut-off grade was used based on whether material is oxide or sulfide: a stope

cut-off range of 4.42 to 6.41 g/t (0.13 to 0.19 oz/st) Au was used for oxide material, and 5.21 to 6.17 g/t (0.15 to 0.18 oz/st) Au for sulfide material. Development headings generate additional material

grading between 1.51 and 2.19 g/t (0.04 to 0.06 oz/st) Au for oxide material and 2.47 to 2.91 g/t (0.07 to 0.08 oz/st) Au for sulfide material; this material is economic to process on an incremental basis and is therefore included in the mineral

reserve estimate.

Table 12-2: Underground Cut-off

Grade Parameters

Process

Method1

Parameter

Imperial

Value

Imperial

Unit

Metric Value

Metric

Unit

Gold Price

2,500

US$

/oz

80.3769

US$

/g

TC/RC

1.85

US$

/oz

0.0595

US$

/g

NSR Royalty

6%

Nevada

Excise Tax

0.75%

Oxides

Metallurgical

Recovery

60.0% - 87.4%

Mining Cost

172.53

US$

/st

190.18

US$

/t

Process Cost

54.38

US$

/st

59.94

US$

/t

Shipping

Cost

14.55

US$

/st

16.04

US$

/t

G&A Cost

20.22

US$

/st

22.29

US$

/t

Total Cost

261.68

US$

/st

288.45

US$

/t

Stope

Cut-off

Grade

0.129 – 0.187

oz/st Au

4.423 – 6.411

g/t Au

Processing

Cut-off

Grade2

0.044 – 0.064

oz/st Au

1.509 – 2.194

g/t Au

Sulfide

Metallurgical

Recovery

77.6% - 92.2%

Mining Cost

172.53

US$

/st

190.18

US$

/t

Process Cost

118.96

US$

/st

131.13

US$

/t

Shipping

Cost

14.55

US$

/st

16.04

US$

/t

G&A Cost

20.22

US$

/st

22.29

US$

/t

Total Cost

326.26

US$

/st

359.64

US$

/t

Cut-off

Grade

0.15–0.18

oz/st Au

5.21–6.17

g/t Au

Processing

Cut-off

Grade2

0.07–0.08

oz/st Au

2.47–2.91

g/t Au

The cut-off grade calculations assume a US$2,500/troy oz Au

price, 60.0% to 92.2% metallurgical recovery depending on the processing method applied, mining costs of US$190.18/tonne, processing costs of US$59.94 to US$131.13/tonne, G&A costs of US$22.29/tonne, a 6% net smelter return royalty, a 0.75%

Nevada excise tax, and treatment/refining charges of US$1.85/oz Au.

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12.3

Reserves Classification and Criteria

Mineral Reserves for the Granite Creek Project are classified as Proven or Probable. Under the S-K 1300

definitions at 17 CFR 229.1300, a Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource and can only result from conversion of a Measured Mineral Resource; a Probable Mineral Reserve is the economically mineable

part of an Indicated, and in some cases a Measured, Mineral Resource. Consistent with this framework, Measured Mineral Resources at Granite Creek were converted to Proven Mineral Reserves, and Indicated Mineral Resources were converted to Probable

Mineral Reserves, by applying the modifying factors described in Section 12.1 to potential mining shapes created during the mine design process; Inferred Mineral Resources were not converted to mineral reserves.

12.4

Multiple Commodity Reserve

Not applicable. The Mineral Reserve for the Granite Creek Project is estimated for gold only; no other commodity grade or quantity is reported

in the source technical documentation for this deposit.

12.5

Impact of Changes of Modifying Factors to the Mineral Reserve Estimates

Mineral reserves are subject to risks typically associated with

drift-and-fill operations. In the QP’s opinion, these risks could materially affect the mineral reserves and include, but are not limited to, the following:

•

Differences between the mineral reserves metal price assumption and actual metal prices.

•

Changes in the interpretation of the deposit geometry and grade continuity.

•

Changes to the estimated gold grades, estimation strategy and input assumptions.

•

Changes to the density values applied to the mineralized domains.

•

Changes to mining, processing, and G&A costs used to determine the

cut-off grade.

•

Changes to metallurgical recovery rates.

•

Changes in the geotechnical assumptions leading to potential impacts to schedule, dilution and/or mining

recovery.

•

Changes to mining method in portions of the deposit.

•

Assumptions related to ongoing access to the site, retention of mineral tenure, obtaining necessary

environmental, mining, and other regulatory permits, and maintaining a social license to operate with relevant stakeholders.

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13

Mining Methods

The Granite Creek Mine currently produces 600 st of ore per day (216,000 st per year assuming 360 days per year) and is operated by a local

contractor. The contractor has operated a number of mines in northern Nevada over the past thirty years. The mine is accessed through the bottom of the open pit with a 4.57 m (15 ft) wide x 4.87 m (16 ft) high decline with a gradient of 13%. Mining

of the Ogee and Otto zones are between 4790 level and 4150 level and the decline to the SPZ is developed to 3950 level.

The mining method

used in Granite Creek is underhand drift and fill with cemented waste rockfill. The production drifts are 4.57 m (15 ft) x 4.57 m (15 ft). Mineralization is loaded on 30 t (33.1 st) mining trucks and hauled to the surface via the decline.

Fresh air comes in from a fresh air portal at the bottom of the pit and is connected to the various workings through raises. The haulage ramp

serves as an exhaust drift.

The proposed mining method for the life of Granite Creek mine will remain the same as the current operations.

The decline will continue down another 650 vertical ft to the bottom of the mine at a 14% gradient.

The ventilation plan will remain the

same with fresh air coming in from the fresh air portal, down raises and exhausted out the haulage ramp. A new 3.2 m (10.5 ft) diameter fresh air raise will be developed to supply enough air to the bottom of the mine.

13.1

Parameters Relevant to Mine Designs and Plans

13.1.1

Geotechnical

WSP, conducted a site visit to the Granite Creek Mine on November 5-6, 2025. The purpose of the visit

was to review geological, geotechnical, mining, and infrastructure conditions relevant to the SPZ and planned mine expansion. The site visit included inspection of drill core, underground workings, ground support systems, cemented rockfill

production facilities, and the proposed ventilation raise location to surface. Observations made during the site visit were used to validate geological interpretations, geotechnical domains, rock mass classifications, mining methods, ground support

recommendations, and ventilation raise development assumptions presented in this report.

The South Pacific Ore zone will be accessed via

the SPZ Decline and mined using the Underhand Cut and Fill method, replacing open stopes with jammed CRF. i-80 has been mining other ore bodies at Granite Creek with a similar mining method with success since

2022.

Benchmarking in the Region and at Granite Creek

Northern Nevada is a well-established underground gold mining region with significant experience available to draw upon for benchmark

comparison. The underhand cut and fill mine method has been demonstrated to be well suited for the challenging ground conditions in the region with other mining operations successfully applying the method including Turquoise Ridge, Cortez Hills,

Goldstrike, and others. Design parameters available to compare include:

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•

Ground support design and implementation

•

Specialized heading advance techniques for difficult ground conditions

•

Backfill design and implementation

•

Minimum sill pillar thickness between mine levels

Benchmarking studies and a review of Warren (2016) indicate the SPZ mine method and designs in this report are in line with experience

gained at most underground mines in the region. Additionally, the current mine workforce draws from experience in the region including technical staff and operations.

Granite Creek has been operating successfully using underhand cut and fill since 2022. Installed ground support and CRF design have generally

included:

•

Moderate – good ground (Rock Mass Rating (RMR) 35 +) 2 m (8 ft) Swellex, 3 ft x 3 ft bolt spacing and wire

mesh and shotcrete where necessary.

•

Poor ground (RMR<30) 2 m (8 ft) Swellex, 2 ft x 2 ft bolt spacing and wire mesh and shotcrete where necessary.

This is typically in fault zones and ore body mining areas.

•

CRF backfill design of 700 psi with emplaced QA/QC program per Golder (2012) design.

The above ground support and backfill design has performed adequately thus far at the Granite Creek mine.

Cemented Rockfill Designs

CRF design strength determination was based on Stone et al., (2019) modified from Stone (1993) and Pakalnis (2005). The SPZ will implement

a CRF design strength of 700 psi (7% cement) with 2-inch minus crushed aggregate sourced from the CX West pit on surface. This is consistent with other ore body extraction designs at Granite Creek and with

recommendations by Golder (2012).

Backfill will be made using a typical pit mix configuration using a front end loader to mix aggregates

with a cement slurry into a consistent mixture. Mixed CRF will be loaded into haul trucks for dumping at the designated location underground. A jammer, (typically an LHD-or similar fitted for jamming) will be

used to “jam” CRF tight to the back underground.

i-80 will continue current practice

of their CRF QA/QC program consisting of:

•

Regular crushed aggregate sampling and testing targeting the ideal Talbot Curve for gradation distribution.

•

Regular CRF backfill cylinder sampling and testing to ensure minimum design strengths are met.

The CRF design strength, mixing operation, and QA/QC program for the SPZ is consistent with other operations in the area

operating in similar lithologies and ore extraction practices.

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Geotechnical Domains for Underground Mining and Development

Geotechnical domains in a mine are developed by dividing the rock mass into zones that exhibit similar engineering properties and will respond

similarly to mining activities. Table 13-1 indicates geotechnical domains developed for the SPZ based on evaluation of the geological model, geotechnical logging of core holes, underground observations, and

laboratory testing results.

Table 13-1: Geotechnical Domains

Geotechnical Domain

Lithology

Description

RMR/GSI

Upper Comus (Ocu)

Argillite

Strong rock (50 to 100 MPa) typically brittle in nature. Thinly bedded to massive

45 (25 to 50)

Lower Comus (Ocl)

Interbedded argillite and limestone, typically metamorphosed to hornfels (argillite) and marble (limestone) with calcsilicate mineral assemblages

Very strong rock >100 MPa moderately fractured to massive rock mass

60 (50+)

SPZ ore body and other faults

Breccia mixed with clay gouge. Clayey Gravel, roughly 10% to 15% clay

Angular breccia mixed with clay gouge. Clayey Gravel, roughly 10% to 15% clay.

20 (<25)

Laboratory Testing

Laboratory testing was performed to obtain strength and material parameters for the SPZ zone main rock types. Testing included 15 (5 with

modulus) Unconfined Compressive Strength (UCS), 27 Triaxial Compressive Strength (TCS), 10 Brazilian Tensile Strength (BTS), and 5 Unified Soil Classification (USCS) tests. Table 13-2 presents typical

properties for the SPZ mining area.

Table 13-2: Geotechnical Parameters for the Three

Geotechnical Domains

Domain

UCS

(MPa)

Tensile

Strength

(MPa)

Specific

Weight

(kg/

M3)

Youngs

Modulus

Ei (GPa)

Poisson’s

Ratio

USCS

Upper Comus

85

11

2788

99.9

0.25

NA

Lower Comus

134

14

2794

82.6

0.26

NA

SPZ Ore Body and Fault Zones

NA

NA

NA

NA

NA

Clayey

Gravel

(GC)

Note: Rock sample testing was performed at the WSP laboratory in Burnaby Canada and at the Earth

Mechanics Institute in Golden Colorado, USA. Soil Classification testing was performed at the WSP lab in Golden Colorado, USA.

Rock

Mass Classification

Geotechnical core logging provides the main source of information for classification of the rock masses in

each of the geotechnical domains within the SPZ mine area. Additional supplemental information includes underground observations from within the existing developments and observation of representative sections in the core shed. For this project, the

main classification system selected was the Rock Mass Rating (RMR) System (Bieniawski, 1976). Input parameters for the RMR system include:

•

Rock Quality Designation (RQD) from core logging based on Deer and Deer (1988)

•

Intact Rock Strength (ISRM R Strength) from core logging, underground mapping, and laboratory testing

•

Joint spacing from core logging and underground mapping

•

Joint condition rating from core logging and underground mapping

•

Groundwater rating is assumed to be damp-wet for empirical design, and

dry for numerical mode input parameters with groundwater conditions considered directly in the modeling conditions

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Table 13-3 indicates the range of RMR estimates by

source for each geotechnical domain along with the representative values selected for estimating rock mass strength.

Table 13-3: RMR Summary for Each Geotechnical Domain

Domain

Range

i-80 Mode

WSP Logging Mode

Observations Underground

Design RMR

Upper Comus

0 to 70

20 to 30

40 to 50

40 to 50

45

Lower Comus

0 to 90

30 to 40

60 to 70

60 to 70

60

SPZ/Faults

0 to 40

NA

20 to 30

0 to 20

20

Geological Structural Assessment

Geologic structures can generally be defined as large-scale (mine wide) and small scale (rock fabric). Structure at the Granite Creek Project

is highly complex and indicative of multiple deformation events. Thickness and orientation of major structures are constrained by core logging, underground mapping, and 3D interpretive modeling (i.e., Leapfrog). Most faults outside the main ore

bodies and Range Front Fault are typically steeply dipping and average 2.4 to 3 m (8 to 10 ft) thick, with a maximum thickness of 6 m (20 ft).

Rock fabric data are available from 11 core holes with televiewer optical and acoustic structural pics. Stereographic analysis of the compiled

data was used to define the major joint sets. Table 13-4 summarizes the major sets defined through the analysis and used as basis for kinematic excavation stability analysis discussed subsequently.

Table 13-4: Major Joint Sets Defined by Televiewer Data Analysis in Dips Software

Design Joint Set Name

Bedding

Dip°

Bedding

Dip Dir°

J2 Dip°

J2 Dip

Dir°

J3

Dip°

J3 Dip

Dir°

Excavation

Orientation

Typical Bedding

50

62

34

243

80

150

NE and NW

Steep Bedding

84

100

15

23

81

160

NE and NW

Stress Assessment

Local geologic and large-scale structural information indicate an extensional in situ stress regime. Vertical stress estimates are assumed to

increase linearly with depth and proportional to the unit weight of the rock. Laboratory results indicated an average rock density of 174.9 pounds per cubic foot (pcf) (2802 kg/m3) with a maximum depth below ground surface of 457 m (1500 ft) = 1,822

psi (12.56 megapascal (MPa)). Basin and range extensional tectonics and back analysis of modeling studies (Sandback et al., 2020) suggest generally isotropic horizontal stress have a K value (horizontal / stress) of 0.75. Therefore, maximum

horizontal stress is assumed to be 1,366.5 (9.42 MPa). Table 13-5 lists assumed stress values assumed for this study.

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Table 13-5: Stress Analysis from Assumed Depth and

k Value of 0.75

Stress

Depth (ft)

Depth (m)

Density (pcf)

K

Stress (psi)

Stress (Mpa)

Vertical

1,500

457.2

174.9

—

1,822.0

12.56

Horizontal

1,500

457.2

174.9

0.75

1,366.0

9.4

Ground Support

Ground support design methodology consisted of the following steps:

•

Nevada-specific empirical support design based on Warren (2016), Warren et al., (2018), Pakalnis et al., (2008),

Pakalnis (2014), and Warren et al., (2020) as a starting point for ground support design

•

Kinematic Factor of Safety (FOS) stability checks based on design joint sets and empirical ground support design

•

2D numerical modeling checks on stability of development drifts and mining drifts based on in situ stress

analysis, mining induced stress, and input parameters from core logging and laboratory testing

Ground Support

Design

Table 13-6 presents the ground support recommendations for Granite Creek SPZ based

on empirical design, kinematic analysis, 2D numerical modeling, and a review of benchmarking studies for other mines in the region.

Table 13-6: Ground Support Recommendations for Granite Creek SPZ Development and Mining

Ground Type

Typical

Geotechnical

Domain

Drift Dimension

(WxH ft)

RMR/ GSI

(range)

Primary bolt 8 ft 120

kN

Swellex-type

bolt (ft x ft) (1)

Secondary Bolt

Spacing 12ft 240

kN Swellex-type

(ft x

ft)

Shotcrete

Type I

Lower Comus

15x16

60 (50+)

4x4

As needed

As needed

Type II

Upper Comus

15x16

45 (25-50)

3x3

As needed

As needed

Type III

Fault Zones Development

15x16

20 (<25)

2x2

6 x 6

3” minimum

Type III (temporary)

South Pacific Zone ore cuts

15x15

20 (<25)

2x2

As needed

As needed

All Intersections

All

Up to 20 ft circular diameter

All

As required above

6x6

As needed

(1)

Welded wire mesh is required and should be installed simultaneously with the primary bolt pattern.

Operational Considerations

Stand-up time in weak ground will be a primary risk to development in fault zones and in the top cuts

of the South Pacific ore body. Experience in these conditions at Granite Creek has shown that spiling with grouted self-drilling hollow-bar MAI bolts has proven effective in stabilizing the back to allow

installation of primary ground support. General operating practices for ground types are presented in Table 13-7 below.

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Table 13-7: Recommended Operating Practices for

Ground Support Installation

Ground

Type

Typical

Geotechnical

Domain

RMR (range)

Span (ft)

Span (m)

Unsupported

Standup Time

Operational

Considerations

I

Lower Comus

60 (55+)

16

4.9

1 month to 1 year

10 ft blast rounds, longer standup time, shotcrete not necessary

II

Upper Comus Development

45 (35-55)

16

4.9

1 to 4 days

8 to 10 ft rounds, likely will stand until next shift can install support, shotcrete where necessary

III

Fault Zone Development and SPZ ore body top cuts

20(<35)

15 to 16

Immediate collapse - hours

4 to 6 ft light blast rounds or muck advance, spiling or flash shotcrete prior mucking likely necessary

Overstressed mine openings, particularly on mining levels, is a common occurrence in Nevada underground

mines due to the generally poor ground conditions and typically results in a time-dependent (squeezing) ground response or convergence of the openings. Successful management of these conditions typically involves mining and backfilling as soon as

possible (within weeks) before squeezing becomes a problem. The selected ground support elements of Swellex-type bolts have a relatively high strain capacity and are well suited for these conditions.

Granite Creek has not experienced evidence of overstressed ground in ore zones or development drifts. However, experience in the region and 2D

Finite Element Analysis indicate that as mining progresses deeper, time-dependent deformation may be encountered, particularly in the top cuts of ore drives.

The minimum sill pillar thickness at Granite Creek is 9 m (30 ft) per mine policy. Modeling using 2D Finite Element Analysis of the minimum

sill pillar thickness indicates potential, but manageable overstressing of the pillar if mining levels are in direct vertical alignment. Staggering level advance so that these openings are not in direct vertical alignment will help to reduce or

prevent this condition from occurring.

Production schedules and ground support maintenance planning need to provide sufficient buffer or

contingency to account for ground support rehabilitation efforts if these conditions are experienced. Mining production schedules should be sequenced to reduce open stope time and backfill as soon as possible, particularly in top cuts.

Ventilation Raises

Two ventilation raises are planned for development (VR1 and VR2) connected by 213 m (700 ft) of underground drift, with an additional 152 m

(500 ft) of development access drift. VR1 is planned from surface (elevation 1,694 m (5,557 ft)) to a depth of approximately 600 ft (elevation 4,957). VR2 is a 91 m (300 ft) internal raise from elevation 4,500 to 4200. Stability analysis based on

cover core holes along VR1 and near the VR2 alignment indicates that raise bore methods are feasible with ground improvements over sections of both alignments. Specific ground improvements include:

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•

The upper 18 m (60 ft) of alluvium in VR1 will require ground improvements such as grout injection and secant

piling.

•

Pre-excavation ground improvements will be required in VR1 from depths 18

to 61 m (60 to 200 ft) and the lower 23 m (75 ft).

•

Pre-excavation ground improvements will be required in limited sections

of the lower 30 m (100 ft) of VR2 within faulted intervals of greater than 2 m (6 ft).

Final lining is required in both

vent raises for long term stability (finished concrete or grouted in place steel liner) installed immediately after raise boring.

13.1.2

Hydrological

Groundwater Assumption

This study assumes the dewatering program will be effective and potential for adverse groundwater-related stability conditions will be

minimized. Accordingly, geotechnical analyses were completed assuming drained groundwater conditions.

Dewatering

There are seven existing sumps in the mine with a 60 HP pump in each sump. Dual 6-inch pipes are used

to transport the water along the decline to surface.

Water from the production faces is pumped by either

air-operated diaphragm or electrical pumps into either 2-inch or 4-inch HDPE discharge lines. The discharge lines are then routed

to the nearest main sump. Pipe size, pump-type, and pump size are adjusted as required to maintain safe access in the production headings. Most ore headings use a 2-inch

air-operated diaphragm pump while development headings most often use a 30 hp or 60 hp electric pump.

The Triple Sump consists of three 50 to 18 m (60 ft) long drifts spurring off of single access at the 4170 Level. Water from the mine is

discharged to Sump 8 and decants through drain holes to Sump 9 and then decants through a second set of drain to Sump 10 in an effort to clarify the water. A concrete weir and catwalk is installed in Sump 10 to allow the clarified water to flow to

the pumps while solids are left in the sump. The total installed capacity of the triple sump system is approximately 60,000 gallons.

Granite Creek has not seen significant additional increases in the amount of contact water encountered during mining since advance in the

exploration drift was completed. The exploration drift and the watercourse encountered in Sump 12 continue to serve as the primary point-sources into the mine. Total flows are estimated at 800 to 1000 gpm between the two sources. The balance is made

from water encountered in the ore headings; flow encountered in lower levels are generally balanced by a decrease in flow at higher levels on the same geologic structure. Figure 13-1 depicts the rolling

tracker for the contact water management system.

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Source: i-80, 2026

Figure 13-1: Contact Water Measurements

The current dewatering model estimates residual passive inflow into Granite Creek will exceed 2500 to 2700 gpm until the combined influence of

the second water treatment plant and additional dewatering wells advance the dewatering cone-of-depression below the current workings.

A Phase 2 upgrade to the dewatering system is in progress to achieve the following:

•

Discharge Line Upgrade:

•

The uphill piping system should be a single line capable of handling all the projected inflow.

•

HDPE is preferred to carbon-steel for material for installation in the decline.

•

Lift Stations:

•

The number of lift stations between the pit crest and the lowest point of the underground workings needs to be

minimized to control capital expenditures and ensure system reliability by decreasing the number of components (i.e., points-of-failure).

•

Lift Stations should be spaced as evenly as possible to balance head pressure to each pump and further increase

system reliability.

•

Station dimensions should fit in a standard muck bay (e.g., 15 ft W x 15 ft H x 50 ft L) where possible.

•

Pumps:

•

Pumps must operate within the pressure spec of the pipe.

•

Installed motor power should be delivered at 480V.

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•

Pump Style should allow for minimal maintenance in the presence of up-to

1% solids (or more).

•

Power:

•

Use of installed load-centers to transform voltage from feed to 480V is preferred to transformation at the

installation location.

•

Control Systems:

•

Variable Frequency Drives (VFDs): Granite Creek currently uses VFDs manufactured by Baker-Hughes and would prefer

to continue to use their product line to ensure parts commonality.

13.1.3

Ventilation

The following site visit observations and temperature/humidity data were collected to calibrate the mine ventilation model discussed under

Ventilation, below.

A site visit was completed on November 5, 2025, by WSP. A guided tour of the mine was completed with i-80 staff that included a visit and review of conditions at the main fan, booster, and auxiliary fans, as well as active working areas, the main haulage route, escapeways, and future ventilation raise locations.

Staff indicated the mine was very humid, with the likely source being groundwater seepage. However, neither SMD nor i-80 had humidity measurements available, so WSP collected this pertinent data while on Site

for use in calibrating the ventilation model. Overall ventilation practices, implementation, and management were discussed during the site visit with available i-80 and SMD personnel.

Temperature and humidity data collected during the site visit is shown in Table 13-8. Locations of

primary airflow in the mine were used to calibrate the existing mine Ventsim™ model since these data are believed to be the most reliable for calibration purposes. Non-working areas with limited airflow were excluded.

Table

13-8: Site Visit Temperature and Humidity Data

Test

Dry Bulb

(F)

Wet Bulb

(F)

Relative

Humidity

(%)

Dew Point

(F)

Location

Notes

1

59.2

50.0

52.2

41.7

Main fan (intake side)

2

58.7

49.7

52.8

41.5

Main fan (intake side)

3

80.0

77.2

88.5

76.3

4480 spiling roof area

(dead heading)

Discarded

4

78.3

71.5

72.3

68.6

4129

5

76.4

69.4

70.9

66.3

Lower decline three

(ventilation tube discharge point)

6

77.1

70.1

71.1

67.0

Exploration decline

(ventilation tube discharge point)

7

78.2

67.7

58.8

62.6

4110 future ventilation raise location

Discarded

8

79.0

65.6

49.6

58.5

Ventilation raise (4175 level)

Discarded

9

87.7

80.2

72.6

77.8

Exhaust near portal

Ventilation at the Granite Creek Mine is circulated through the two portals, one as intake and one as

exhaust. The haulage is on the only decline, which acts as the exhaust airway. Intake airflow is delivered through a series of raises that were developed through the center of the spirals on the decline. This Pre-Feasibility Study (PFS) expands upon

previous lower level of detail studies to describe

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improvements with enhanced engineering and sensitivity analyses required. Figure 13-2 shows the airflow scheme for the planned mine development.

Source: WSP, 2026

Figure 13-2: Future Mine Airflow Configuration

Weather

The

psychrometric properties of the surface air were calculated using the weather data collected by i-80 over the last two years. From the data recorded, WSP calculated average values for summer and winter to be

used in modeling.

Diesel Equipment

Diesel equipment operating underground produces pollutants that must be diluted to safe working levels according to the United States Mine

Safety and Health Administration (MSHA). The comprehensive list of every diesel machine on site that has the potential to go underground must be examined for calculating adequate airflow. WSP applied airflow rates based upon the engines fitted to

each machine, assuming a similar horsepower (hp) rating where an exact engine is not listed.

When considering the total hp and cubic feet

per minute (cfm) ratings for the equipment, on average 44 cfm is applied for each hp operating in the mine. This considers that all equipment is on and operating for a whole shift, which should be seen as the maximum airflow requirements, not the

typical. WSP assumes that 80% of the equipment will be working in the mine during any shift, meaning that only about 336 thousand cubic feet per minute (kcfm) will be required under normal circumstances. The other 20% of equipment is assumed to

be in the maintenance shop or on surface and is not considered a part of the airflow demand calculation.

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Relevant MSHA Standards

To remain in compliance, the mine must comply with the following MSHA standards. The ventilation models developed by WSP respect these

standards to the fullest extent possible. Changes in infrastructure are recommended based on non-compliance with any of these limits. The relevant standards are shown in Table

13-9.

Table 13-9: MSHA Standards

MSHA Standard

Value

Units

Minimum Airflow per Person

200

Cfm

Carbon Monoxide Maximum Concentration

50

Ppm TWA

Nitrogen Dioxide Maximum Concentration

5

Ppm TWA

Diesel Particulate Matter (DPM) Maximum Concentration Limit

160

µg/m3 TWA

Respirable Silica Maximum Concentration Limit

50

µg/m3 TWA

General Respirable Dust Maximum Concentration Limit

2.0

mg/m3 TWA

Maximum Temperature Dry-Bulb

120

°F

Maximum Working Temperature Dry-Bulb1

90

°F

Maximum Working Temperature Wet-Bulb1

79

°F

Source: WSP, 2026

Notes:

(1)

These temperatures are the MSHA action limit. Work can be performed above these temperatures if required,

however, regimented breaks and personnel monitoring are required which slows production significantly.

Heating and

Cooling Requirements

WSP modeled the winter and summer stages, with allowances for localize heat generation from backfill curing,

to see the more extreme scenarios of temperatures and found that no cooling is required in the summer, although air temperatures will feel quite warm around diesel equipment and operators should remain in enclosed cabs whenever possible. Open cab

drills and bolters are expected to have sufficient cool airflow to keep operating conditions for personnel below the MSHA action limits. Dry-bulb temperatures in the decline will be locally elevated, wherever

haul trucks are present, and all haul trucks should be outfitted with functional air conditioning units. The most critical method to avoid excess heat in at the working faces is to duct air from the intake raise over an air door through the decline

to each level. The temperatures become too hot when reusing air from the decline, because of the truck traffic.

In the winter scheme,

intake air low temperatures are at, or slightly below, freezing. This can cause icing in the intake raise if wet-bulb temperatures remain below freezing for extended periods of time because humidity drops out

of the air and forms ice around the raise circumference. If enough ice builds up, the resistance of a small, open cross-section in the raise can stall the fan and cut off airflow. To mitigate this risk of reduced airflow and fan damage, WSP

recommends installing a 1.8 MBTU/hr propane heater near the fan and ducting it into the plenum just upstream of the vanes. This amount of heat is enough to raise the temperature of the air intake by 12°F dry bulb (6°F wet bulb) and should be

sufficient to prevent icing for 95% of the operating days during each year according to the site recorded temperature data. For the remaining 5% of the days, i-80 may either turn the fan off temporarily and

operate at a lower production rate or rent another 1.8 MBTU/hr unit to manage the icing risk on a case-by-case basis. Each heater is estimated to cost about US$40/hour

to operate.

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DPM and Dust Management

WSP simulated DPM using Ventsim™, which derives the concentrations from particulate

mixing and dispersion using the source diesel hp or user set emission rate as an input. Since the decline is exhausting, the highest concentration of DPM is experienced there, about 500

µg/m3 steady state. Therefore, WSP recommends equipment movement other than truck haulage is minimized on the decline. This initiative is expected to be adopted by i-80 already considering traffic on the decline slows production.

The silica dust simulation offered

very similar results to DPM, except dust is expected to drop out of the air at some point as it moves up the decline so leakage of dust from the decline to the intake raises is modeled as 0

µg/m3. There was no data provided to WSP to calculate or simulate fugitive dust loading in the mine, so silica dust was the focus. i-80 is in

compliance with regulations on silica dust exposure but had a few readings that were above the action level of 25 µg/m3 in the testing data provided. WSP does not recommend any

infrastructure or administrative changes for silica dust exposure beyond those recommended for airflow, heat, and DPM. However, closed cabin equipment with HEPA® filters would reduce silica

dust exposure in the same way the filters will reduce DPM exposure.

New Ventilation Raise Infrastructure Design

Improving on previous studies, the ventilation modeling completed by the engineering team at the Granite Creek property identified that

sufficient airflow could be achieved with a single raise from surface to the 4600 level, with drifts to bypass adverse ground conditions, and the development of an additional raise connecting the 4500 level to the 4200 level. The raises are designed

to be 4 m (12 ft) excavated diameter and 3.2 m (10.5 ft) finished with a sealed collar. Environmentally permitted space is limited so WSP has designed the new installation to fit within a small footprint. WSP engaged a vendor to quote a fan of

sufficient size and power to provide the required airflow to the operations. One permanent and one temporary (rental) 1.8 MBTU heater have been included at the raise collar along with a 10,000-gallon (80%

filled) propane tank for a week supply of fuel. The general arrangement of the surface fan infrastructure is shown in Figure 13-3. Variations of this design have been proposed to reduce costs but are not

included in this report.

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Source: WSP, 2026

Figure 13-3: Surface Fan Installation General Arrangement

Ventilation Capital and Operating Costs

After developing the ventilation schemes for the future mining operation, WSP estimated the number of air doors, regulators, fans, and duct

work stock to purchase. Some of these items are available in inventory at Granite Creek or sister properties, while other items will need to be purchased new. To reduce the capital cost to i-80, WSP has

assumed that available inventory will be used where applicable. Table 13-10 shows a list of the main capital expenditures that will be required for the future mining operations and were included in the

estimated cost.

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Table 13-10: Capital Cost Estimate

Item

Cost

(US$)

Raisebore Excavation

8,820,000

Surface 400 HP Fan Installation1

3,870,000

Underground Air Door Installations

1,720,000

Underground Fan And Ductwork

Installations4

1,500,000

Total

$

15,900,000

Source: WSP, 2026

Notes:

1.

Approximately 25% of the costs sourced from CostMine™

(Costmine Intelligence, 2024), the remainder from budgetary quotations.

2.

A total estimated cost of US$1.25 million for injection grouting was added to the raisebore excavation

based upon previous WSP projects of similar scope. Injection grouting for ground improvement costs can vary widely depending on the actual ground conditions observed.

3.

Shipping and installation costs are estimated to be 3-5 times the

amount of the raw materials cost. This is based on a basket of recent contractor bids received for various projects. Shipping and installation costs are not applied to the raisebore excavation cost.

4.

No drifting included in cost estimates.

5.

Only shipping cost of 15% applied to underground fan and ductwork installations because installation is assumed

to be completed by the full-time i-80 ventilation crew labor, with cost captured as an OPEX.

6.

Values may not add exactly due to rounding.

The operating costs associated with ventilation are almost entirely captured in power consumption, heating and labor for moving fans and

installing / removing ductwork. Fans require limited maintenance and are designed to work for 10 to 20 years without planned maintenance service if used properly. Likewise, air doors and regulators are designed with robust and simplistic properties

for use in harsh mining environments. The doors and regulators are operated by a hydraulic over electric system with few moving parts. Servicing is limited to basic functions like greasing rotating parts and exchanging hydraulic fluid at periodic

intervals. Table 13-11 shows the annual operating expenditure calculation.

Table 13-11: Annual Operating Cost Estimate

Item

Count

Hourly OPEX1

(US$)

Annual Operating

Hours2,3

Annual Cost

(US$)

Fans

400 HP

2

44.15

8322

735,000

250 HP

1

29.80

8322

248,000

75 HP

4

17.69

8322

147,000

40 HP

7

5.78

8322

289,000

Air Doors

18

0.09

8322

14,000

Heater

1

38.80

2891

112,000

Ventilation Crew Labor

2

63.00

8760

1,100,000

Getman A64 (Scissor deck)

1

46.58

7884

367,000

Total

$

3,000,000

Source: WSP, 2026

Notes:

1.

Fan, air door and mobile equipment maintenance hourly OPEX sourced from CostMine™ (Costmine Intelligence, 2024). Hourly labor cost sourced from i-80 actual data.

2.

Fans and air doors are considered to have 95% availability and operate year-round. Mobile equipment

availability is not considered since multiple machines are available.

3.

Power cost given by i-80 Gold was US$0.08/kWh.

4.

Values may not add exactly due to rounding.

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Conclusions and Recommendations

After conducting a series of Ventsim™ simulations, WSP concluded that ventilation at

the Granite Creek property can be improved through the following measures: the installation of a 10.5-ft finished diameter raise from surface to the 4600 level, followed by a drift to bypass adverse ground conditions, and the development of an

additional 10.5-ft finished diameter raise connecting the 4250 level to the 4200 level. This ventilation configuration is expected to be sufficient for the remainder of the mine life as presented in the i-80

PFS mine plans and the updated schedule provided by i-80 that includes additional reserves. The main capital additions that accompany the raise-bored excavation include another

400-hp fan, the upgrade of the existing fan motor from 250 hp to 400 hp, one 1.8 MBTU propane heater, one 10,000-gallon propane tank, and the necessary electrical

components for safe operation of the fan. These items are estimated to have a capital cost of US$15.9 million. Additional investment could be made in enclosed cabs for the equipment items which currently have open cabs, but WSP does not see

this as a necessity to comply with MSHA regulations based on the modeling completed. However, i-80 should monitor conditions and implement controls as necessary based on actual results. Additionally, to better

control DPM and silica dust exposure, WSP recommends all headings are ventilated directly from the intake raises with the ductwork passing over air doors placed in the crosscuts which connect the decline to the intake raises. This method provides

fresh air directly to working faces for an improved working environment.

13.2

Production Rates, Expected Mine Life, Mining Unit Dimensions, and Mining Dilution and Recovery Factors

Mine Stope optimization (MSO) within the Deswik software was used to determine the potentially economically minable

material. The variable cut-off was calculated and written into the block model. 15 ft x 15 ft drifts were generated over the deposit at different orientations to generate minable envelopes. Level intervals

were set to match the current mine plan. A mine design was generated using the envelopes as guidance. Figure 13-4 shows the minable envelopes and the mine design.

Source: SRK, 2026

Figure 13-4: Production Design with MSO Guidance

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Level accesses are developed from the decline to access a maximum of five cuts. An attack

ramp provides access to the top cut (Cut 1), once the top cut is mined cemented backfill is placed and cured before the next attack ramp is developed to the bottom cut (Cut 2). A curing period of 21 days will be required before mining can occur

below. This continues until all the cuts are mined and backfilled. Figure 13-5 shows a schematic of the cut sequence.

Source: SRK, 2026

Figure 13-5: Schematic of Cut Mining Sequence

The mining sequence on a level generally follows a retreat sequence. The access drift is developed to the furthest point of the level and

retreats to the access. Mining adjacent to backfill requires 24 hours to allow the backfill to cure and prevent blast damage. Figure 13-6 shows the production drift sequencing of a typical level.

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Source: SRK, 2026

Figure 13-6: Production Mining Sequence

Underground infrastructure includes sumps for dewatering at roughly every 100 vertical ft. Load centers are located on the ramp to provide

power to the mining fronts. These cut outs start off as remucks as the decline is being developed and are repurposed after the ramp is complete.

The main maintenance shop is located on surface between the decline and the fresh air portal. A smaller maintenance bay is located underground

that stores hoses and consumables. The powder and cap magazine are located underground near the top of the mine. Figure 13-7 shows the location of the shop and powder and cap magazine.

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Source: SRK, 2026

Figure 13-7: Underground Infrastructure Location

A muck bay is placed at the entrance to each production pod and serves all the cuts within the pod. Trucks are loaded on the decline.

The cemented waste rock backfill plant is located on surface. The backfill material is loaded onto trucks at the plant and hauled to the drift

requiring backfill. Backfill is tight filled and allowed to cure.

The Granite Creek design resulted in 2.39 Mst at an average Au grade of

0.23 oz / st. Figure 13-8 shows the mine design colored by activity type.

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Source: SRK, 2026

Figure 13-8: Granite Creek Mine Design Looking Northwest

The production schedule is based on the mine design and Mineral Reserves discussed in Section 12. Productivity rates were generated based

on actual footage achieved by contractors at the mine as well as geotechnical inputs on ground quality.

The productivity rates used for

mine scheduling are based on the ground type defined in Section 13.1 for the top cuts and a constant rate for the undercuts, as they are mined under backfill. They are shown in Table 13-12.

Table 13-12: Productivity Rates

Activity Type

Ground Type

Rate1

Units

Capital Development

Type I

9.0

ft/d

Type II

8.0

ft/d

Type III

4.0

ft/d

Production

Type I

7.0

ft/d

Type II

6.0

ft/d

Type III

4.0

ft/d

Undercut

9.0

ft/d

Vertical Development

—

5.0

ft/d

Backfill

—

600

st/d

Source: WSP, 2026

(1)

All rates are per face. Multiple areas/faces are mined together to generate the production schedule.

General schedule parameters applicable to all underground mining activities are presented in Table 13-13.

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Table 13-13: Schedule Parameters for Underground

Mining

Schedule Parameters

Units

Value

Annual mining days(1)

days/year

365

Mining days per week

days/week

7

Shifts per day

shifts/day

2

Scheduled shift length

hrs/shift

12

Scheduled

Deductions

Line-Up and Safety Talk

hrs/shift

0.25

Travel - In

hrs/shift

0.25

Pre-Op Checks

hrs/shift

0.25

Lunch/Breaks (travel to lunchroom)

hrs/shift

1.00

Wash and grease at end of shift

hrs/shift

1.00

Operator unavailable and other interference

hrs/shift

0.25

Vehicle loading/Pick-up Area

hrs/shift

0.08

Travel - Out

hrs/shift

0.25

Total Non-Operating Time

hrs/shift

3.33

Total Operating Time

hrs/shift

8.67

Utilization of shift time

%

72

%

Source: WSP, 2026

Actual operational mining days are 360. For simplicity the schedule has been completed assuming 365 days, with

pro-rated productivity rates.

The production and development schedules were completed using Deswik

software. The mining operation schedule is based on 365 days per year, seven days per week, with two twelve-hour shifts each day. The production schedule continues with the current production target of 544.3 t (600 st)/d of ore and ramps up to 907.2

t (1,000 st)/d in 2028 as the SPZ is developed and headings become available.

Granite Creek has an

8.3-year projected mine life, from 2026 to 2034. Table 13-14 and Figure 13-9 summarize the production schedule.

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Table 13-14: Granite Creek Annual Production

Schedule

Period

Ore

Tonnes

(kt)

Ore Au

Grade

(g/t)

Au Metal

Content

(oz)

Waste

Tonne

(kt)

Backfill

Tonne

(kt)

Development

Length

(m)

20261

169.56

7.72

42,078

142.22

230.30

5,694

2027

203.65

7.48

48,969

132.82

244.31

6,050

2028

334.96

7.73

83,289

113.71

360.06

7,831

2029

333.23

7.73

82,769

82.44

366.25

7,215

2030

291.52

8.08

75,718

62.50

367.69

6,115

2031

339.27

7.89

86,084

66.07

378.97

6,993

2032

273.97

8.04

70,800

32.35

355.58

5,218

2033

175.89

8.09

45,768

11.43

225.79

3,160

2034

50.16

9.20

14,835

3.36

64.31

902

Total

2,172.20

7.88

550,311

646.90

2,593.26

49,178

Source: WSP, 2026

Last 8 months of 2026.

Source: WSP, 2026

Last 8 months of 2026.

Figure 13-9: Granite Creek Annual Production Schedule

Mine development drift and raise dimensions are described in Section 13.3. Mining dilution and mining recovery factors, and the

methodology used to derive them, are described in Section 12.1; a 10% external dilution (at zero grade) and 100% stope and development recovery were applied.

Dilution, Ore Loss & Grade Control

The ore control geologist targets viewing each active heading once per shift. The geologist takes a photo of the face and makes notes.

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Muck samples are collected by haul truck drivers at the windrow. Development headings receive

one sample per round and are automatically shipped as waste unless otherwise directed by the geologist. For ore headings, a sample is collected at the rate of roughly one sample per two trucks. The driver uses a hand-held sample scoop to fill a

sample bag half full (five to ten pounds), walking along the windrow and taking a scoop every 2 m (5 ft), making sure to collect both coarse and fine material. Sample bags have a tag with duplicate bar codes separated by a perforation. The

perforated portion of the tag includes space to handwrite sample source information including mine level, heading ID and distance, date and shift. The haul truck driver places completed sample bags in a designated location near the mine office

trailer. The ore control geologist collects the samples accumulated from the previous day and night shifts and uses the sample tag information to generate a sample submittal for the laboratory and fill the information to the ore control database.

The ore control geologist inserts QA/QC samples into the sample stream. A contract driver transports muck samples, Over the Road (OTR) truck samples as well as any drill samples to the Lone Tree laboratory once per day.

The Lone Tree lab analyzes samples for Au grade by fire assay and cyanide absorption, sulfide %, TOC, CO3, and preg-rob potential. The results

are used to characterize each round as oxide or autoclave refractory, high grade or low grade, or waste. Assays must be approved by the database administrator before the ore control geologist can enter assay results in the ore control database and

flag windrows for routing. The geologist ties color coded flagging associated with the assessed ore type to a lath at the end of the windrow. The windrow can then be moved to the stockpile corresponding to its ore type. The process typically takes

three days from mine face mucking to ore type determination and flagging.

High grade oxide and sulfide ores are screened to 3 inches at

the stockpile. The minus 3-inch portion is loaded for shipment to the appropriate ore processing location (autoclave or oxide mill). Screened oversize oxide material is placed in the low-grade oxide stockpile, which is shipped to the heap leach facility on a low priority basis. Oversized sulfide material is transferred to long term on-site low-grade sulfide stockpiles and is not shipped.

Ore is shipped to processing facilities using contract

OTR trucks. The truck driver receives a ticket number at the security gate and gives the ticket number to the loader operator at the stockpile. The loader operator loads the first bucket, then spills a small portion of every other bucket into a

small sample pile on the ground during the loading process. Once the loaded truck departs, the loader operator collects a sample from the sample pile and labels the bag with an ID associated with the truck ticket number. Sample bags are waterproof

to preserve moisture content. The loader operator places the samples at the designated sample location at the end of the shift, where they are collected by the ore control geologist who prepares a laboratory sample submittal and enters the sample

information into acQuire. Trucks are weighed near the security gate when departing the mine, and security personnel email a report of truck tons and ticket numbers at the end of the shift.

13.3

Requirements for Stripping, Underground Development, and Backfilling

Stripping

Not

applicable. Granite Creek is an underground operation; underground access is via the existing Granite Creek open pit and a decline (see Section 13.1), and no additional pit stripping is required to support underground production.

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Underground Development

Capital development is sized 15 ft x 15 ft with an arched back and includes the decline, sumps, ventilation drifts and mine load center. The

decline gradient averages 14%. Development avoids the main structures where possible. Where it intersects the faults, adequate ground support is used.

Operating development includes attack ramps to access the level, muck bays, production drifts and internal accesses to connect different pods.

The production drifts are 15 ft x 15 ft with a flat back, while all other development is 15 ft x 15 ft with an arched back. A 15 m (50 ft) offset between the production drift and a turn is maintained to allow for proper mucking and backfilling by

the LHD.

Fresh air raises are 3.2 m (10.5 ft) in diameter and are developed using a raise bore machine. A drill and blast drop raise

method is used for raises less than 30 m (100 ft).

All capital and operating development are performed by contractors.

Backfilling

The

mine production sequence includes the use of CRF to fill the voids left by the production drift. As the mining method is underhand drift and fill, the CRF needs to have adequate strength to provide a stable roof to work under. A higher strength CRF

with 7% cement is used to fill the drifts. For the life of mine, 2.86 Mst of backfill is required.

CRF is produced at the surface plant

and hauled underground to the drifts using trucks. A modified LHD with an extended boom and push plate attached to the end is used to tight fill the drift. Backfill aggregate is sourced from waste rock mined in the CX West open pit.

13.4

Required Mining Equipment Fleet and Machinery, and Personnel

Material Handling System

The mine plan assumes 10 t (11.0 st) LHDs will muck the ore from the production face to the muck bay located near the level access. From the

level access, 30 t (33.1 st) trucks are loaded and then hauled up the decline. Waste rock will be handled similarly and hauled to surface. The material haulage distances and cycle times were calculated using the haulage profile in Deswik. It is

assumed that the trucks will return with backfill.

The one-way haul distance for the waste

material increases as the mine gets deeper and averages 2,667 m (8,750 ft) over the life of the mine. The ore haulage distance averages 2,713 m (8,900 ft) over the life of the mine. Average roundtrip cycle times for waste and ore material is one

hour.

Equipment & Fleet – Contract Mining

The mobile equipment fleet will consist of diesel vehicles provided by the contractor. The estimate assumes equipment availabilities of 80% to

85%, a job efficiency factor of 83% (50 minute hour), and an activity efficiency factor that varies by activity.

The owner equipment will

include pickup trucks for staff and a core drill for definition drilling.

Table 13-15 summarizes

the mobile equipment required throughout the mine life.

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Table 13-15: Mobile Equipment Fleet

Equipment Type

Number

Required

2 Boom Jumbo

3

Bolter with Screen Handler

6

10 t Development LHD

4

Emulsion Loader

3

30 t Truck w/ Push Box

8

Scissor Lift

1

Multi-Lift

1

Boom Truck

1

Fuel Lube Truck

1

Grader

1

Shotcrete Truck

1

Transmixer Truck

1

Personnel Carrier

2

UG Forklift

1

Diamond Drill

1

4 x 4 Utility Vehicle

5

4 x 4 ATV

3

Mine Rescue Vehicle

1

Total Equipment

44

Source: WSP, 2026

Manpower & Operations – Contractor Model

Labor levels are estimated based on the production schedule and equipment needs. Development and production are performed by contractors. Staff

labor includes both the mine technical staff from owner and the contractor. Mine technical staff will work 4 @ 10-hour shifts per week. Operations will work twelve hours per shift, two shifts per day and seven

days per week.

The mine manpower required is shown in Table 13-16. This assumes a two to four crew

rotation for operations and maintenance.

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Table 13-16: Granite Creek Manpower

Labor Category

Position

Staffing

Technical Staff

Staff

UG Mine Manager

1

Staff

UG Mine General Foreman

1

Staff

Senior Mining Engineer

1

Staff

Short Term Planning Engineer

1

Staff

Geotechnical Engineer

1

Staff

Senior Mine Geologist

1

Staff

Mine Geologist

3

Staff

Surveyors

2

Staff

Administrative Assistant

1

Staff

Safety

1

Technical Staff Subtotal

13

Operation

Hourly

Lead Development Miner (jumbo)

5

Hourly

Development Miner (bolter)

9

Hourly

Blaster

5

Hourly

Development Helper (Services)

1

Hourly

LHD Operator

6

Hourly

Truck Operator

13

Hourly

Shotcrete/Transmixer Operator

7

Hourly

Lead Construction Miner

2

Hourly

Construction Miner

2

Hourly

General Laborer

4

Operation Subtotal

54

Maintenance

Staff

Maintenance Superintendent

1

Staff

Maintenance Planner

1

Hourly

Shop Supervisor

1

Hourly

Mechanic

5

Hourly

Electrician

2

Hourly

Instrumentation

2

Hourly

Welder

2

Hourly

Mechanic Helper

2

Hourly

Electrician Helper

2

Maintenance Subtotal

18

Mine Total Manpower

85

Source: WSP, 2026

13.5

Final Mine Outline Map

Figure 13-10 shows the final mine outline map.

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Source: WSP, 2026

Figure 13-10: Final Mine Outline Map

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14

Processing and Recovery Methods

This section describes the methods by which ore from the Granite Creek Project is processed and gold is recovered. These methods include whole

ore pressure oxidation via the Lone Tree facility or a third-party facility.

14.1

Lone Tree Facility Historical Processing

The Lone Tree Mine is located immediately adjacent to US Interstate 80 (i-80), approximately 19 km (12

miles) west of Battle Mountain, NV, 80 km (50 miles) east of Winnemucca, NV, and 193 km (120 miles) west of Elko, NV. Mining commenced at Lone Tree in April 1991 with the first gold pour in August 1991. In 1993, a POX circuit was added to the

facility, which included a SAG / ball mill circuit, followed by a thickening circuit, the POX process for refractory gold ores, and finally CIL, carbon stripping, and refining.

In 1997, a 4,500-tpd flotation plant was constructed to make concentrate to supplement the whole ore

feed to the POX circuit. Excess concentrate was shipped to Newmont’s (now NGM) Twin Creeks POX plant or Carlin roaster. The Lone Tree processing facilities were shut down at the end of 2007. Since that time, the mills have been rotated on a

regular basis to lubricate the bearings. In general, the facility is still in place with most of the equipment sitting idle.

The objective

of i-80 Gold Corp is to refurbish and restart the POX circuit and associated unit operations, including the existing oxygen plant, to similar conditions as it was operating before the shutdown, while meeting

all new regulatory requirements. The flotation circuit is not being considered for restart. The POX circuit will have capability to operate under either acidic or basic conditions but is initially anticipated to be operated in an acidic environment.

In order to restart the process plant, new environmental regulations must be met limiting allowable mercury emissions. In February 2011,

the NDEP and the EPA brought about new standards to limit mercury emissions to 127 lb of mercury for every million tons of ore processed. In order to meet this requirement, the Lone Tree Facility will require several modifications. In 2022 and 2023

a number of multidiscipline site inspections were completed to assess the current facility condition and develop the requirements for the restart plan.

14.2

Lone Tree Facility Process Description

An overall process flow diagram for the Lone Tree Facility is shown in Figure 14-1. The process flow

sheet includes the following major processing areas:

•

Ore reclaim, grinding, and thickening and acidulation

•

Pressure oxidation

•

POX off-gas Treatment and quench water loop

•

Neutralization, carbon-in-leach,

and cyanide destruction

•

Tailings thickening and filtration

•

Acid wash, carbon stripping, and carbon regeneration

•

Electrowinning and refinery

•

Plant and instrument air

•

Oxygen plant

•

Reagent preparation and storage

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•

Process and plant services cooling towers

•

Water distribution

•

Steam generating plant and propane storage

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Source: i-80 Gold, 2025

Figure 14-1: Overall Process Flow Diagram for the Lone Tree Facility

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14.2.1

Key Process Design Criteria

The Lone Tree POX Facility restart will have minimal changes made from the 1993 process design criteria (PDC). A new PDC was developed based on

the expected production sources as defined byi-80.

Key process design criteria are summarized in

Table 14-1.

Table 14-1: Summary of Key Process

Statistics

Criteria

Units

Value

Annual Mill Throughput

tonnes (tons)

827,806 (912,500)

Daily Throughput (per calendar day)

tonnes (tons)

2,268 (2,500)

Operating Throughput of Ore to Autoclave Circuit

mtph (stph)

111.2 (122.5)

Grinding Circuit Product Size,

P80

µm (US mesh)

74 (200)

Operating Time / Availability

%

85

Design Sulfur Treatment Rate

mtph (stph)

2.4 (2.7)

Gold Recovery

%

Varies

Silver Recovery

%

Varies

Source: i-80, 2026

The feed to the Lone Tree autoclave will consist of material from the various mines. The design scenarios showing the feed chemistry and grades

are shown in Table 14-2.

Table 14-2: Design Scenarios

for Lone Tree Autoclave Feed

Design Scenario

Unit

Case 2

Case 4

Case 6

Case 8

Case 10

Blend

100% Granite Creek

100% Ruby Hill

100% Cove

20% GC

40% RH

40% Cove

60% RH

40% Cove

Sulfide Sulfur

%

1.12

1.88

1.13

1.43

1.58

Carbonate

%

11.09

15.92

9.21

12.27

13.24

Organic Carbon

%

0.60

0.15

0.47

0.36

0.28

Gold Grade

g/t

11.3

5.83

11.3

9.26

7.89

Source: i-80, 2026

Specific to the Granite Creek Project, dependent on the characteristics of the ore, gold recovery is expected to vary over the life of the

mine. LOM average of refractory ore via pressure oxidation is expected to be 87.5%, with a range of 78.8% to 94.0%. The dependency of this recovery is discussed in Section 13.

14.2.2

Ore Reclaim

The purpose of the ore reclaim area is to store and reclaim ore for processing, which has been shipped to the Lone Tree processing facility via

highway trucks.

ROM ore will be delivered to the ore stockpile area. Production from various mining locations, specifically Granite Creek,

Cove, and Ruby Hill, will be dumped at designated locations within the storage area and blended into facility feed stockpiles.

The

stockpile will be sited to the north of existing plant buildings in an area of existing disturbance. The stockpile will include a lined base graded to drain to a geomembrane lined collection pond in order to convey and contain precipitation that

falls on the stockpile area. The collection pond is sized to contain run-off produced from the stockpile area during a 100-year /

24-hour storm event.

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Ore will be hauled to the site and placed on the stockpile using conventional over-the-road tractor trailers. A front-end loader will be used to consolidate the trailer dumped piles of ore, and to transfer ore

from the stockpile to an existing feeder hopper located at the south end of the stockpile.

The stockpile area has been sized to provide

capacity for approximately 154,000 tonnes(170,000 tons) of ore to accommodate the shipment of ore to site and provide allowance for necessary blending.

A layout diagram of the area is shown in Figure 14-2.

Source: i-80 Gold, 2026

Figure 14-2: Ore Stockpile Area Layout Diagram

Additionally, the reclaim area is utilized for feed blending for the Lone Tree POX circuit. This blending will be used to manage sulfide sulfur

concentrations, gold grades, and carbonate grades feeding the autoclave to ensure stable circuit operation within the design criteria for the plant.

14.2.3

Grinding

The purpose of the grinding area is to reduce the ROM ore particle size to support target sulfide oxidation kinetics within the autoclave

resulting in increased gold recovery. The ROM feed particle size distribution will be sufficient to direct feed the ore to the grinding circuit without the need for pre-crushing. The grinding area contains a

SAG and ball mill circuit to produce the required grind size. The

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SAG mill is fed via a conveyor from the dump hopper. The SAG mill discharges over a vibrating screen with oversize returned to the SAG mill feed conveyor. Vibrating screen undersize discharges

into the cyclone feed pump box together with ball mill discharge. The slurry is then pumped to the cyclones with the cyclone underflow gravitating to the ball mill and the cyclone overflow is directed to the POX feed thickening circuit.

14.2.4

Thickening and Acidulation

The purpose of the thickening area is to prepare the slurry for autoclave processing by densifying the product of the grinding circuit to

improve storage capacity of the downstream slurry storage tanks, improve the autoclave heat balance by reducing the water transferred to the autoclave, and improve the solids flow through the autoclave feed pumps. The thickened slurry is acidified

in two acidulation tanks with sulfuric acid to reduce the carbonate content in the autoclave feed to target a ratio of 1.1:1 carbonate-to-sulfide sulfur ratio. The

acidulation tanks double as feed storage tanks to provide a combined storage capacity of 12 hours.

The storage tanks ensure continuous

feed to the autoclave plant, unaffected by upstream throughput variations.

14.2.5

Pressure Oxidation

The POX autoclave circuit includes the slurry pre-heaters, autoclave feed, autoclave, pressure let down

vessels (flash system) and the POX ancillary services: autoclave agitator seal system, oxygen supply, high pressure cooling water, and high-pressure steam. The Lone Tree Facility restart expects the autoclave to operate under acidic conditions.

Slurry Heaters

The purpose of the slurry heaters is to capture excess energy discharged from the autoclave and

pre-heat the feed slurry prior to the autoclave process thereby reducing the total energy input required to operate the autoclave. The heating is achieved in two stages consisting of a series of two refractory

lined counter-current splash slurry heater vessels.

The heat source is flashed steam released from the autoclave discharge slurry during

the pressure letdown process. The splash slurry heaters are direct contact heat exchangers that provide a means of heat recovery via steam condensation. This reduces the off-gas load on the downstream off-gas equipment and reduces the required input steam from the steam boiler.

Autoclave Feed

The purpose of the autoclave feed area is to increase the pressure of the pre-heated slurry

above the autoclave operating pressure to facilitate transfer into the autoclave at the required pressure using the autoclave feed pumps. These pumps are positive displacement pumps selected for the design temperatures and pressures.

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Autoclave

The purpose of the autoclave is to oxidize the refractory sulfide minerals under acidic conditions to liberate the gold trapped in the sulfide

sulfur minerals. The autoclave at Lone Tree is designed to operate at 198°C (389 °F) and 297 PSI(g) with a slurry residence time of 40 to 50 minutes and consists of four compartments. The design expects up to 97% cumulative sulfide sulfur

oxidation through the autoclave. High purity oxygen is introduced to all four compartments of the autoclave at controlled rates. Due to the low sulfur grades, steam is required to be continuously fed to the first compartment of the autoclave to

maintain the kinetically required oxidation temperature. The addition of oxygen and steam support the required oxidation rates to achieve the target sulfide sulfur oxidation. The autoclave slurry is discharged through a level control choke valve and

is fed to the high-pressure flash vessel in the pressure letdown system.

Flash System

The purpose of the flash system is to reduce the pressure and temperature of the autoclave discharge slurry, making it suitable for subsequent

unit operations downstream. The oxidized slurry undergoes a controlled pressure and temperature reduction process as it passes through two stages of flash vessels located downstream of the last autoclave compartment. The flash steam is recycled

upstream to preheat the autoclave feed slurry in Section 17.2.5.

Slurry Coolers

The purpose of the slurry coolers is to reduce the temperature of the incoming slurry from the

low-pressure flash vessel to prepare it for the downstream neutralization and CIL circuits through a series of six water-cooled shell and tube heat exchangers.

14.2.6

POX Off-Gas Treatment

The purpose of the POX off-gas treatment area is to effectively eliminate particulate matter present in

the POX vent stream, while simultaneously reducing the temperature and volume of the vent gas through direct contact condensation. The POX off-gas treatment circuit also includes a mercury removal step to

minimize autoclave mercury emissions to the environment.

Quench Vessel

The vent streams originating from the autoclave, low temperature and high temperature slurry heaters are directed to the quench vessel. The

purpose of the quench process is to reduce the gas load to the venturi scrubber by cooling and condensing the steam from the vent gas stream prior to feeding the scrubber. This process also indirectly aids in removing entrained particulate droplets

and contaminants. Non-condensed vapor and gases discharge from the top of the vessel and are fed to the Venturi Scrubber. Condensed solutions are directed to the quench water circuit.

Venturi Scrubber

The Venturi scrubber acts as the final stage of particulate gas cleaning prior to the mercury abatement carbon beds. Water containing removed

particulates is directed to the quench water circuit.

Emergency Relief Vessel

The purpose of the emergency relief vessel is to serve as an engineered discharge point for effectively managing relief events in the autoclave

and the off-gas circuit. Excess pressure beyond vessel Maximum Allowable Working Pressures (MAWP) within the system can occur due to loss of quench water, high pressure conditions in the autoclave, or elevated

pressure levels in the high-pressure flash vessel.

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The vessel mitigates the consequences of an emergency release by separating any potential

slurry carryover from the gaseous emissions and reducing the velocity of the vent gas before it discharges to atmosphere. The gaseous emissions are discharged outside at the top of the POX mercury abatement structure. The collected solution in the

relief vessel is directed to the neutralization area sump pump, which transfers it to the neutralization distributor tank.

POX

Mercury Abatement

The purpose of POX mercury abatement system is to remove mercury and volatile organic compounds (VOC’s)

from the POX vent gas via adsorption on to activated carbon before the final discharge to the atmosphere.

Quench Water

Heated quench water and condensed steam in the POX off-gas circuit is collected, cooled, and reused

within the POX off-gas circuit. The quench water tank collects solutions from the quench vessel and Venturi scrubber, and discharges to the quench water clarifier for entrained solids separation. The clarified

solution is cooled via three cooling towers prior to being recirculated to the quench vessel. Excess cooling solution created from condensing vapors in the off gas is bled from the circuit to the mill water tank. The separated solids from the

clarifier underflow are pumped to the neutralization circuit where they rejoin with the cooled autoclave discharge slurry.

Cooling

Towers

The purpose of the cooling area is to reject heat absorbed within the process to the atmosphere. The solution cooling area

includes the process service cooling circuit and the plant service cooling circuit. The process cooling circuit rejects the heat from the autoclave cooling circuit and the elution circuit heat exchangers. The plant service cooling circuit provides

trim heat rejection from various equipment support systems throughout the plant. Standard treatment of the cooling water is performed with biocide and corrosion inhibitor.

14.2.7

Neutralization

The purpose of the neutralization circuit is to neutralize all free acid in the slurry, precipitate the heavy metals as their hydroxides, and

raise the pH to approximately 10.5 to ensure cyanide stability in the carbon-in-leach circuit for personnel safety and process optimization. The neutralization circuit,

comprised of a distributor and three tanks, is dosed with lime slurry to raise the pH of the autoclave discharge slurry. Sparged oxygen converts ferrous iron to ferric iron, reducing downstream cyanide consumption. The neutralized slurry from this

circuit is then fed to the CIL circuit for gold recovery.

14.2.8

Carbon-in-Leach

The purpose of the CIL circuit is to leach and extract gold and silver from the oxidized and neutralized slurry

using cyanidation and carbon adsorption. The CIL circuit provides retention time of 22 to 24 hours. The CIL circuit consists of six mechanically agitated tanks arranged in series. The agitators prevent solids settlement and maximize contact time

with cyanide and activated carbon to improve gold and silver recovery. Oxygen (as opposed to air) is sparged into each tank to maximize the rate of gold dissolution. Each tank is fitted with a carbon retention screen to keep activated carbon with

each tank while allowing slurry to flow to the next tank. Slurry and carbon are pumped upstream through the

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circuit to provide counter current flow of the carbon relative to the slurry flow. Loaded carbon from the first CIL tank is pumped to an elution circuit for carbon stripping and regeneration.

Stripped, regenerated and sized barren carbon is fed to the last tank of the CIL circuit. The leached slurry is transferred from CIL to the cyanide destruction circuit.

14.2.9

Carbon Elution and Regeneration

The purpose of the elution circuit is to elute precious metals from the loaded carbon and transfer the resulting solution of high gold

concentration (pregnant eluate) to the refinery to produce doré. Barren carbon is thermally regenerated prior to addition to the CIL circuit.

Carbon Acid Wash

The purpose of acid wash is to rinse the loaded carbon from CIL with dilute nitric acid solution prior to the carbon stripping process.

Carbonate scale builds up on the activated carbon during the CIL process and fouls the carbon’s adsorption properties by depositing a layer of scale. If left intact, over time the scale will limit the adsorption capacity of the carbon and will

cause softening of the carbon in the regeneration kiln. The loaded carbon from CIL is first treated within the carbon acid wash vessel prior to treatment within the carbon stripping vessel.

Carbon Stripping

The purpose of the carbon strip circuit is to strip the acid washed loaded carbon of the adsorbed gold using a Pressure ZADRA Strip scheme at

149°C (300°F) and 414 kPag (60 psig). Pressure ZADRA strip uses several bed volumes of a recirculated solution to strip the precious metals off the loaded carbon. The cyanide solution is buffered by caustic to assist with gold elution. The

stripped carbon is then sent to carbon regeneration circuits. The resulting pregnant eluate is processed in the electrowinning circuit.

Elution Mercury Abatement

The purpose of elution mercury abatement system is to condition the off-gas leaving the pregnant and

barren solution tanks to remove fine particulate, solution aerosols, and condensed and gas phase mercury using absorption on to a new carbon bed.

Carbon Regeneration

The purpose of the carbon regeneration circuit is to restore the activated carbon’s ability to recover gold from the cyanidation circuit

solutions. The circuit also permits the introduction of new carbon to the process and removes carbon fines from the process.

As carbon is

used in the CIL and elution circuits, the surface and internal pore structure becomes contaminated with organic species. The organics foul the carbon, slow the gold adsorption rate, and decrease the gold loading capacity of the carbon. The carbon

reactivation electric kiln is a horizontal rotary kiln that heats the carbon in an inert atmosphere at 750°C (1400°F) and is specifically designed to remove the organic fouling and regenerate the carbon to an active state where it can be

used again to adsorb leached gold in the CIL circuit. The rotary kiln operates continuously until a batch of carbon is processed.

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Carbon Fines Handling

Carbon fines that are too small to be mechanically separated from the slurry in the CIL circuit, and would lead to precious metal losses, are

collected and transferred by gravity from the reactivated carbon vibrating screen, carbon reactivation feed vibrating screen, kiln feed hopper, and carbon reactivation electric kiln. The carbon fines are dewatered in a filter press and discharged

into supersacks for external sale.

Refinery

The purpose of the refinery circuit is to recover concentrated gold from cyanide solutions via electrowinning (EW) and produce doré

bars.

Electrowinning

The EW circuit recovers gold from the pregnant eluate by applying a voltage across cathodes and anodes in cells. Pregnant eluate is pumped

through three EW cells to electrowin the gold and silver, along with other contaminants, into a sludge that plates onto the cathode mesh material. The plated sludge is manually washed from the cathodes and then melted and refined. Cathodes are made

from stainless steel mesh, and anodes are stainless steel punch plate.

Refining

The refining process is to produce doré bars by smelting retorted sludge mixed with fluxes. The fluxes form a slag to which contaminates

report to.

The sludge from the EW cells is first processed in a mercury retort oven to dry the sludge and remove the co-captured mercury from the precious metal recovery steps. The retorted gold sludge is then processed in a melt furnace to remove contaminants through slag and produce the final doré bars that are processed

by third party refineries for sale.

14.2.10

Cyanide Destruction

The purpose of the cyanide destruction (CND) circuit is to effectively reduce the concentration of cyanide in the final tail discharge and the

recycled process water to ensure compliance with predefined environmental standards and regulations. CND also improves the safety of the operation by reducing cyanide concentrations outside the CIL and elution circuits. The CND circuit targets a

specific concentration limit of 2.5 mg/L of residual weak acid-dissociable (CNWAD) cyanide. This reduction is accomplished through the application of the SO2/air cyanide destruction process

using oxygen in a series of two tanks, which oxidizes the cyanide (to cyanate) to meet the required concentration level. The cyanide destruction circuit is fed directly from the slurry discharge of the CIL circuit.

14.2.11

Tailings Management

The purpose of the tailings circuit is to increase the density of the detoxified tailings to aid with dry stacking of tailings residue.

Additionally, this circuit produces process water for internal use within the facility. The tailings circuit consists of a thickener as a first stage of solids densification. The thickener underflow is fed to a tailings filtration circuit which

dewaters the tailings sufficiently to support tailings dry stacking. The de-watered tailings from the filter presses are then dry stacked at the tailings storage facility.

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The water removed from the tailings slurry is used as process water within the facility to

offset water requirements. Excess process water is processed via a reverse osmosis circuit to provide supplemental permeate, water to offset freshwater requirements.

Tailings Thickening and Overflow Clarification

The purpose of the tailings thickening area is to provide the initial stage of densification and to recycle process water for the facility via

thickener overflow using a conventional thickener. Dilute flocculant solution is added to the thickener feed to promote settling and clarify the thickener supernatant. The thickener underflow of 55 to 65 wt.% solids is pumped to a filtration feed

tank.

Tailings Filtration

Thickener underflow at 55% to 65% solids by weight is further dewatered via the four horizontal plate and frame tailings filter presses. They

are operated in coordinated filtration cycles, two operating and two standbys, to produce a filter cake with a target 15% moisture (17.65% geotechnical moisture). There are two different sizes of filter presses, with the legacy filter presses from

the prior operation being repurposed to provide additional filtration capacity, if needed to supplement the newly installed filter presses. The filter cake is collected by a mobile loader and transferred to the filtered tailings facility, the filter

press filtrate reports back to the tailings thickener.

Process Water

The tailings thickener overflow tank is the primary process water tank for the facility. The process water is used directly in grinding, CIL,

tailings filtration, and reagent make-up areas. Water in excess of that which is needed is treated in the reverse osmosis circuit. In case of process upsets surge water can be directed from the tailings

thickener overflow tank to the water surge pond. The pond provides surge capacity and has the ability to discharge to the tailings thickener deaerator tank to be fed back through the thickener if needed for solids removal and the overflow tank for

distribution to the other circuits if it does not contain solids.

Reverse Osmosis

The purpose of reverse osmosis (RO) is to treat excess process water to obtain a clean permeate water that can be used as fresh water

throughout the facility. The RO circuit comprises pressure filtration, an ultrafiltration system, and an RO system.

14.2.12

Water Distributions

There are eight types of defined water services at Lone Tree:

•

Fresh water – Is generally used for reagent make-up, and water

washing streams.

•

Gland water (fresh water) – Is used to supply gland water to slurry pumps.

•

Mill water – (grinding thickener overflow and process water) Is used to provide dilution water within the

milling circuit.

•

Potable water (fresh water) – Is used for safety showers and sanitary uses, generated from a new potable

water skid.

•

Demin water fresh water) – Is primarily used to supply the steam generating plant provided from a

demineralization water plant.

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•

Process water (tailings thickener overflow) – Is used for mill water makeup. Additionally, it feeds the

reverse osmosis circuit.

•

Quench water (condensed steam from autoclave off-gas) – Is used

within the POX off-gas circuit as the source of direct cooling water.

•

Excess water (bleed stream from RO, Demin and Cooling towers) – Is discharged from the main processing

facility to the excess water treatment facility (battery limit for restart project).

14.2.13

Reagents

Each set of compatible reagent preparation and storage systems is located within dedicated containment areas to prevent erroneous mixing of

reagents. Storage tanks are equipped with level indicators, instrumentation, and alarms to reduce the risk of spills during normal operation. Appropriate ventilation, fire and safety protection, safety shower stations and Safety Data Sheet stations

are located throughout the facility.

The required reagents with their estimated consumption rates are as follows:

•

Ammonium bisulfite (65%) – 2.3 L/t (0.56 gal/ton)

•

Activated carbon – 0.036 kg/t (0.073 lb/ton)

•

Cooling tower antiscalant and corrosion inhibitor – 0.017 L/t (0.004 gal/ton)

•

Biocide – 0.004 L/t (0.001 gal/ton)

•

Caustic (50%) – 0.17 L/t (0.041 gal/ton)

•

Copper sulfate – 0.11 kg/t (0.215 lb/ton)

•

Flocculant – 0.13 kg/t (0.259 lb/ton)

•

Fuel – 0.23 L/t (0.054 gal/ton)

•

Glycol – 0.021 L/t (0.005 gal/ton)

•

Lime – 0.017 kg/t (0.033 lb/ton)

•

Nitric acid (35%) – 0.13 L/t (0.030 gal/ton)

•

Phosphate (10%) – 0.004 L/t (0.001 gal/ton)

•

Sodium cyanide (30%) – 0.32 kg/t (0.63 lb/ton)

•

Sodium sulfite (10%) – 0.004 L/t (0.001 gal/ton)

•

Sulfuric acid (93%) – 175 kg/t (350 lb/ton)

•

Propane – 10.9 L/t (2.61 gal/ton)

14.2.14

Oxygen Plant

High purity oxygen is primarily used for the oxidation of sulfides during the POX process, for iron conversion from ferrous to ferric in the

neutralization circuit, and for cyanide to cyanate conversion in cyanide destruction. Furthermore, during cyanidation, the addition of oxygen maximizes the rate of gold dissolution. At Lone Tree, oxygen will be supplied by a vendor owned and

operated oxygen plant through a Sale-of-Gas (SOG) contract, of which the vendor is responsible for supply, operation and maintenance of the facility. The oxygen plant is

mandated to supply up to 200 stpd of gaseous oxygen at 375 psi(g), delivered to the autoclave, and at reduced pressure to the neutralization tanks, CIL tanks, and cyanide destruction tanks for plant uses. Pressure reduction will be by i-80 Gold for the individual circuits.

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14.2.15

Compressed Air

The Lone Tree facility includes separate instrument and plant air systems to support the facility’s air requirements with excess capacity

and redundancy built into the air system design.

14.2.16

Utilities Consumption

The plant consumptions for water and power are provided for the average processing case below and consider the design blend of material to be

processed within the Lone Tree Facility for the design life of operation.

14.2.17

Water Consumption

Table 14-3 provides a summary of the water consumption by type for the Lone Tree processing facility.

Table 14-3: Lone Tree Facility Water Consumption by Type

Type

Consumption

(m3/h)

Consumption

(gpm)

Mill Water

352

1 550

Fresh Water

130

570

Low Pressure Gland Water

24

105

High Pressure Gland Water

39

170

Demineralized Water

25

110

Potable Water

3

15

Source: i-80, 2026

Electrical Power Requirements

The estimated annual electrical energy requirements for the Lone Tree processing facility are summarized by area in Table 14-4. Annual power requirements are 142,090 MWh or 156 kWh/ton based on 85% availability.

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Table 14-4: Lone Tree Facility Energy Usage by

Area

Area

Annual Energy Consumption

(MWh/y)

General Plant Wide

2,250

Water System

930

Potable Water

240

Process Water (RO and Process Water Tank)

4,900

Ore Reclaim

770

Refinery

2,310

POX Grinding

26,920

POX Grinding Thickening and Acidulation

1,890

Neutralization and CIL and Acid Storage

6,540

Carbon Stripping

4,090

Cyanide Destruction

690

Reagents

2,640

Plant Air and Propane

3,310

POX and POX Utilities

15,540

POX Demineralized Water System

2,660

Tailings Filtration

13,690

Plant Wide Electrical and Instrumentation

4,000

ABS and CN Storage

160

POX Mercury Abatement

900

Quench Water Treatment

4,020

Oxygen Plant

40,090

Existing Plant Areas

3,570

Total

142,090

Source: i-80, 2026

14.3

Third-Party Refractory Processing

Prior to the recommissioning of Lone Tree, production designated as refractory will be processed by a third party as part of a toll milling

agreement, details of which are provided in Section 19. Material will be transported by over-the-road trucks to the toll processor. The basic processing steps are

described in the following sections. A block flow diagram of this process is shown in Figure 14-3.

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Source: i-80 Gold, 2026

Figure 14-3: Third-Party Pressure Oxidation Facility Flow Diagram

The pressure oxidation facility at the toll miller consists of SAG milling followed by two-stage ball

milling. The final grinding product is 80% passing 200 mesh. The cyclone overflow reports to a thickener. Thickener underflow reports to an acidification circuit where sulfuric acid is added as necessary to ensure adequate autoclave free acid

solution levels. The free acid concentration for Turquoise Ridge Complex ore needs to be maintained at >18 g/L. Thickener overflow solution is returned to the milling circuit. After acidulation, ore slurry is added to two identical autoclaves

that are operated in parallel. Two stages of flash heat recovery are used. Autoclave discharge is cooled before reporting to the lime neutralization circuit. Autoclave waste gas is cooled and scrubbed before discharging to the atmosphere. Oxide ore

and acidic oxidized sulfide ore slurry are combined in the neutralization circuit. After neutralization with the carbonate oxide ore and supplemental lime, the ore slurry reports to a CIL circuit where the ore is leached in cyanide solution to

extract the gold. Final tailings slurry is pumped to the tailings area. Tailings settle and decant solution is reclaimed and reused in the grinding circuit. Loaded carbon from the CIL circuit is transferred to the recovery plant. After acid washing

to remove inorganic contaminants, the carbon is transferred to the pressure Zadra stripping circuit. Gold is stripped from the carbon using caustic and cyanide solution at elevated temperature and pressure. Pregnant solution from the stripping

circuit is pumped to an electrowinning circuit where precious metal is removed from the solution as sludge. The sludge is filtered, dried in a mercury retort, mixed with fluxes, and refined into doré bars. After carbon stripping, the barren

carbon reports to the kiln regeneration circuit and returns to the CIL circuit.

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Refractory material from Granite Creek is fed to the autoclave as a proportion of the overall

feed blend and processed in monthly batches. The quantity of gold in each batch is determined through net scale truck scale weights, moisture samples and gold assays by both parties. The gold recovery is determined using a bench scale autoclave test

and then scaled to the refinery production for that particular month.

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15

Infrastructure

15.1

Summary

Previous operators of the Granite Creek Project constructed a large portion of the infrastructure required to operate the Granite Creek

underground mine. This includes:

•

The connection to the NV Energy grid

•

5 MW substation

•

13.8 kV distribution network

•

Administration building

•

Dry facilities

•

Three of the active dewatering wells

•

Two of the four permitted rapid infiltration basins

Since acquisition of the project, i-80 has added to the existing infrastructure with completion of:

•

Dewatering well GCW5

•

Deepening of well BPW5 to 677 m (2,220 ft)

•

Expansion of the dry facilities to accommodate 176 personnel

•

Installation of a rental Water Treatment Plant to treat 800 gpm for discharge to the RIB’s

Infrastructure construction in progress includes:

•

Drilling and completion of dewatering wells GCW14 and GCW15

•

Construction of a 3,300 gpm water treatment plant

•

Construction of the two remaining permitted RIBs

•

Expansion of the contact water handling system and tie it into the water treatment circuit

After completion of the infrastructure construction currently underway, the electrical power distribution, dewatering, water treatment,

administrative and related support infrastructure will be adequate to support the planned underground mine operations. The cost to complete infrastructure construction as of June 1, 2026, is US$5.7 million.

15.2

Access and Local Resources

The Granite Creek Mine is located 32 km (20 miles) northeast of Golconda. From Golconda follow NV789 (paved) northeast for approximately 26 km

(16 miles) followed by the Getchell Mine Road (unpaved) four miles to the Granite Creek access road. The towns of Winnemucca and Battle Mountain are 27 km (17 miles) west and 61 km (38 miles) east of Golconda respectively via Interstate 80.

Osgood Mining Co. (OMC) provides transportation only for senior level employees. The mine and drilling contractors provide light vehicles for

employee transportation. Parking at Granite Creek is available for personal and company vehicles adjacent to the administration building.

Winnemucca and Battle Mountain are both established communities serving the mines and ranches in the surrounding areas. As of the 2020 census,

the population of the cities of Winnemucca and Battle Mountain were 8,431 and 3,705, respectively.

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Northern Nevada hosts several operating gold mines that produced a combined 3.5M ounces in

2024. The area is home to a skilled workforce, contractors and suppliers that can fulfill all the needs of Granite Creek.

15.3

Administrative Facilities

OMC offices, dry facilities, warehouse, truck scale and security office are located just off the Getchell Mine Road at the entrance to the

property (Figure 15-1). The dry facility is housed in the main administration building and a double wide trailer (Figure 15-2). It can serve a total of 176 OMC and

contractor employees. Office space is available to accommodate 25 employees and warehouse area totals 5,000 square feet (ft2) enclosed plus additional yard area. Site security and visitor check-in are provided at the main gate. Security also operates the truck scale where all ore trucks entering and leaving the property are weighed.

Source: OMC, 2026

Figure 15-1: Granite Creek Overview

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Source: Google Earth, 2026

Figure 15-2: Granite Creek Administration Area

15.4

Dewatering

15.4.1

Dewatering Wells

There are four dewatering wells currently operating at the Project (APW-1, BPW-3, BPW-5 and GCW-6). In total, five wells existed prior to i-80’s purchase of the

property in 2021. BPW-2 and BPW-4 were both drilled and tested in 2008 but were never outfitted with permanent pump assemblies due to lower than anticipated production

rates and limited effectiveness for underground operations. GCW-6 was drilled in 2022, tested in Q1 2023 and put online Q1 2024. In Q3 2024, BPW-5 was deepened to a

total depth of approximately 671 m (2,200 ft), targeting the AP Offset structure for increased hydraulic head depressurization of the planned underground progression. As of Q1 2026, GCW-14 has been drilled and

tested and is awaiting completion of GCW-15 in Q2 2026 to allow for accurate sizing of the pump assemblies at both locations (Figure 15-3).

The currently operating wells (APW-1, BPW-3, BPW-5 and GCW-6) are pumping from the CX block hydrogeologic unit at a combined average rate of approximately 1,300 gpm with additional pumping of approximately 1,350 gpm

collected from sumps in the underground mine workings. Water is discharged to the Rapid Infiltration Basins (RIBs) on the southeast portion of the mine boundary for infiltration into the downgradient alluvial basin aquifer. A pipeline has been

constructed to connect the dewatering circuit to a water treatment circuit, so the water can be treated if required. Currently,

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BPW- 3 (600 gpm) does not require treatment and discharges directly to the RIBs. Approximately200 gpm of

BPW-3’s production is routed to the site freshwater supply tank for mining operations, dust suppression and as support for fire suppression. The pump parameters for the dewatering wells are listed in

Table 15-1.

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Table 15-1: Granite Creek Active Dewatering Wells

Well Identifier

Mine Coordinates

Elevation

(ft amsl)

Casing

Diameter

(inches)

Well Depth

(ft bgs)

Static

Water

Level

(ft bgs)

Screened

Interval

(ft bgs)

Average

Pumping

Rate

(gpm)

Pump

Power

(hp)

Pump

Set-Depth

(ft bgs)

Easting

(ft)

Northing

(ft)

APW-1

9890.2

10154.3

4722.3

18

620

380

120 -140

160 -180

200 -600

210

100

574

BPW-3

10188.9

9474.8

5057.1

18

1391

780

500 -540

580 -620

660 -700

740 -1380

600

400

1307

BPW-5

10387.1

11126.9

5093.9

18

12

2222

708

679 -1380

1400-2222

380

200

1290

GCW-6

10310.9

11742.9

5153.2

14

2093

742

803 - 2083

110

150

1950

GCW-14

11272.8

12953.1

5091.1

14

2178

758

966 - 2168

TBD

TBD

TBD

GCW-15

11176.6

12298.1

5086.2

14

TBD

TBD

TBD

TBD

TBD

TBD

Source: LRE, 2026

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15.4.2

Monitoring Wells and VWPs

Monitoring wells and Vibrating Wire Piezometers (VWP) are used to collect hydrogeological data in support of mining operations. Currently,

there are 41 active monitoring wells and 16 active vibrating wire piezometers across 6 locations (Figure 15-3). Construction and recent water level data are provided in Table

15-2.

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Source: LRE, 2026

Figure 15-3: Map of Dewatering Wells, Monitoring Wells, Piezometers at Granite Creek

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Table 15-2: Summary of Locations, Construction

Information and Water Levels for Dewatering Wells, Monitoring Wells, and Piezometers

Identifier

Local Mine

Coordinates

Elevation of

Land Surface

(ft amsl)1

Year

Completed

Inclination

(degrees)2

Open Interval of Well

or VWP Setting

Geologic

Unit(s)

Depth

(ft bls)4

Static Water Level

Comments

Easting

(ft)

Northing

(ft)

Depth

(ft bls)3

Elevation

(ft amsl)

Elevation

(ft amsl)5

Date

(dd/mm/yyyy)

Dewatering Wells

APW-1

9889.60

10152.60

4722.28

2005

-90

120 to 600

4602 to 4122

Ocl

308.1

4414.18

27/08/2024

CX Pit (active)

BPW-2

9804.81

10554.13

4762.98

2008

-90

200 to 920

4563 to 3843

Ocl

385.4

4377.58

23/11/2024

CX Pit (inactive)

BPW-3

10188.94

9474.81

5057.14

2008

-90

500 to 1380

4557 to 3677

Ocl

662.53

4394.61

10/07/2024

South of CX Pit (active)

BPW-4

10806.54

9132.50

5011.94

2008

-90

540 to 1380

4472 to 3632

Ocl

627.5

4384.44

07/10/2024

South of CX Pit (inactive)

BPW-5

10387.06

11126.94

5093.90

2008 / 2024

-90

679 to 2222

4415 to 2872

Ocu; Ocl

703.06

4390.84

17/12/2024

Between Mag and CX West Pits (active)

GCW-06

10310.90

11742.90

5153.20

2022

-90

803 to 2083

4350 to 3070

Ocu; Ocl

1852.3

3300.90

27/12/2024

Between Mag and CX Pits (active); deepened 2025

GCW-14

11272.8

12953.1

5091.1

2025

-90

966 to 2166

4195 to 2295

Ocu, Ocl

747.57

4343.5

1/30/2027

North of Mag Pit (Inactive) 14 inch diameter casing

GCW-15

11176.6

12298.1

5086.2

TBD

TBD

TBD

TBD

Ocu, Ocl

TBD

TBD

TBD

North of Mag Pit (Inactive) 14-inch diameter casing. Currently drilling

Water Supply Wells

WW-7

15630.88

10407.68

4787.39

1984

-90

230 to 420

4557 to 4367

Qal

232.82

4554.57

6/30/2003

E of county road (inactive); 12-inch diameter casing; no sounding port for

water level measurement; alternate identifier PW7; owned by NGM

WW-8

15141.92

8899.17

4756.11

1987

-90

210 to 560

4546 to 4196

Qal

190

4566.11

31/08/2000

East side of county road (inactive); no sounding port for manual DTW

VWPsa

iGS22-17

11081.00

11264.70

4830.70

2022

—

—

—

Ocu

Mag Pit (active)

iGS22-17D_4188

10801.60

11478.90

—

2022

-62

736

4188

Ocu

—

—

—

iGS22-17C_3735

10588.50

11598.00

—

2022

-62

1250

3735

Ocl

—

—

—

iGS22-17B_3391

10428.30

11688.90

—

2022

-62

1640

3391

Ocl

—

—

—

iGS22-17A_3233

10354.60

11732.70

—

2022

-62

1820

3233

Ocl

—

—

—

iGS22-25

10411.80

11448.50

5104.00

2022

—

—

—

Ocu

Between Mag and CX Pit (active)

iGS22-25D_4281

10081.30

11747.90

—

2022

-63

937

4281

Ocl

—

—

—

iGS22-25C_4190

10049.50

11779.40

—

2022

-64

1038

4190

Ocl

—

—

—

iGS22-25B_4077

10010.20

11819.40

—

2022

-63

1164

4077

Ocl

—

—

—

iGS22-25A_3978

9975.60

11855.50

—

2022

-63

1275

3978

Ocl

—

—

—

iGS22-26

11247.10

12508.00

5092.00

2022

—

—

—

Qal

North of Mag Pit (active)

iGS22-26D_4199

11147.30

12590.70

—

2022

-83

902

4199

Ocu

—

—

—

iGS22-26C_3983

11124.00

12609.50

—

2022

-82

1120

3983

Ocu

—

—

—

iGS22-26B_3635

11087.10

12638.80

—

2022

-82

1472

3635

Ocu

—

—

—

iGS22-26A_3384

11061.10

12659.10

—

2022

-80

1725

3384

Ocl

—

—

—

iGS23-10A

11005.50

12407.30

5111.00

2023

—

—

—

Ocu

North of Mag Pit (active)

iGS23-10A_4412

10935.20

12481.40

—

2023

-81

707

4412

Ocu

—

—

—

iGS23-10A_3861

10857.40

12517.30

—

2023

-81

1264

3861

Ocu

—

—

—

iGS23-10A_3700

10830.50

12526.90

—

2023

-81

1428

3700

Ocu

—

—

—

iGS23-10A_3579

10811.70

12533.50

—

2023

-81

1552

3579

Ocl

—

—

—

iGS23-02A

11091.40

11430.30

4826.60

2023

—

—

—

Ocu

Mag Pit (active)

iGS23-02A_4128

10792.40

11906.90

—

2023

-52

898

4128

Ocu

—

—

—

iGS23-02A_3747

10634.20

12132.80

—

2023

-52

1371

3747

Ocu

—

—

—

iGS23-02A_3641

10592.10

12195.10

—

2023

-52

1500

3641

Ocl

—

—

—

Monitor Wells

GMWCX-1

9844.60

12588.10

5258.10

1997

-90

505 to 545

4753 to 4713

Ocl

382.8

4875.30

11/12/2024

GMWCX-2

8409.50

11701.20

5580.20

1997

-90

465 to 505

5115.2 to 5075.2

Kgd

278.92

5301.28

05/11/2024

GMWCX-3

8235.70

10619.10

5321.30

1997

-90

278 to 318

5043.3 to 5003.3

Cpy

274.45

5046.85

05/11/2024

GMWCX-4

7785.30

7991.50

5335.50

1997

-90

610 to 670

4725.5 to 4665.5

Cpy

506

4829.50

05/11/2024

GMWCX-5

10341.30

8970.70

5073.00

1997

-90

490 to 523

4583 to 4550

Ocl

DRY

DRY

11/05/2024

GMWCX-5D

10008.58

8670.70

5109.77

2008

-90

1000 to 1080

4109.77 to 1029.77

Ocl

711.68

4398.09

23/12/2024

Replacement for GMWCX-5

AMW-1

9857.09

10652.34

4762.31

2005

-90

900 to 960

3862.31 to 3802.31

Ocl

383.55

4378.76

23/12/2024

BPZ0802

11070.54

9224.37

4995.80

2008

-90

—

—

Qal

DRY

DRY

—

BPZ0803

9329.24

11014.22

5194.53

2008

-90

—

—

Ocl

678.8

4515.73

16/05/2024

GMW-HLMW-1

14221.21

4917.50

4692.27

1989

-90

254

—

Qal

264.06

4428.21

27/09/2023

GMW-RCH-588

15614.11

10449.75

4788.60

1995

-90

505

—

Qal

193.94

4594.66

28/09/2023

GMW-W7A

20480.58

10155.92

4657.70

1995

-90

110

—

Qal

125.94

4531.76

28/09/2023

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Identifier

Local Mine

Coordinates

Elevation of

Land Surface

(ft amsl)1

Year

Completed

Inclination

(degrees)2

Open Interval of Well

or VWP Setting

Geologic

Unit(s)

Depth

(ft bls)4

Static Water Level

Comments

Easting

(ft)

Northing

(ft)

Depth

(ft bls)3

Elevation

(ft amsl)

Elevation

(ft amsl)5

Date

(dd/mm/yyyy)

MW 8

15140.35

8999.35

4757.01

1998

-90

212

—

Qal

154.87

4602.14

02/10/2023

RCH-1101

11440.58

12178.09

5060.51

1992

-90

264

465

Ocu

270.71

4789.80

05/06/2024

RCH-1280

13419.41

9893.00

4850.00

1991

-90

—

—

Qal

211.25

4638.75

23/12/2025

RCH-1305

10859.00

10333.45

5012.79

1992

-90

300

—

Ocu

DRY

DRY

—

Records indicate well has collapsed

RCH-1308

10981.17

9783.77

4996.49

1992

-90

285

—

Ocu

DRY

DRY

—

Records indicate well has collapsed

RCH-1309

10642.73

10744.50

5048.88

1991

-90

337

—

Ocu

DRY

DRY

—

Records indicate well has collapsed

RCH-1515

12602.88

9750.12

4895.68

1993

-90

219 to 465

4676.68 to 4430.68

Qal

258.1

4637.58

05/06/2024

RCH-1516

12395.45

10333.65

4904.70

1993

-90

229

—

Qal

DRY

DRY

—

Records indicate well has collapsed

RCH-1517

12458.51

10705.76

4915.72

1993

-90

229

—

Qal

DRY

DRY

—

Records indicate well has collapsed

RMW2NE

15915.24

7792.24

4718.68

2005

-90

178 to 218

4540.68 to 4500.68

Qal

104.16

4614.52

15/02/2024

RMW2SE

15902.63

7292.08

4711.48

2005

-90

158 to 198

4553.48 to 4513.48

Qal

91.35

4620.13

15/02/2024

RMW2W/GMWMW2A

13432.00

9194.40

4838.20

1992

-90

140

500

Qal

203.91

4634.29

20/03/2024

RMW3NE

15621.68

6584.88

4710.37

2005

-90

158 to 198

4552.37 to 4512.37

Qal

90.52

4619.85

15/02/2024

RMW3SE

15488.44

6250.25

4706.13

2005

-90

158 to 198

4548.13 to 4508.13

Qal

87.59

4618.54

15/02/2024

RMW3W

13681.94

7080.52

4804.22

2005

-90

278 to 318

4526.22 to 4486.22

Qal

179.19

4625.03

15/02/2024

WELL 10

10454.99

11200.32

5084.57

1992

-90

242

542

Ocu

DRY

DRY

16/05/2024

Records indicate well has collapsed

WELL 2A

13452.52

9148.66

4839.80

1992

-90

144 to 450

4695.8 to 4389.8

Ocu

204.15

4635.65

05/06/2024

WELL 6

8381.24

9858.41

5168.70

1998

-90

274

—

Cpy

248.75

4919.95

16/05/2024

WSW-W#11

10907.57

12300.59

5117.49

1998

-90

105 to 505

5012.49 to 4612.49

Ocu

DRY

DRY

—

WSW-W#2

13024.00

9420.00

4861.90

1992

-90

251 to 555

4610.9 to 4306.9

Ocu

228.83

4633.07

05/06/2024

WSW-W#9B

10323.65

11966.86

5173.91

1998

-90

264 to 617

4909.91 to 4556.91

Ocl

DRY

DRY

—

Rib Piezometers

RPZ2E

14833.50

8020.25

4768.41

2005

-90

58 to 138

4710.41 to 4630.41

Qal

90.03

4678.38

15/02/2024

RPZ2N

14732.84

8348.93

4771.00

2005

-90

58 to 138

4713 to 4633

Qal

128.5

4642.50

15/02/2024

RPZ2S

14595.45

7762.05

4773.27

2005

-90

58 to 138

4715.27 to 4635.27

Qal

81.05

4692.22

15/02/2024

RPZ2W

14496.42

8090.13

4785.56

2005

-90

58 to 138

4727.56 to 4647.56

Qal

88.58

4696.98

15/02/2024

RPZ3E

14388.32

6615.05

4763.06

2005

-90

58 to 138

4705.06 to 4625.06

Qal

DRY

DRY

15/02/2024

RPZ3N

14377.55

6939.79

4772.25

2005

-90

58 to 138

4714.25 to 4634.25

Qal

DRY

DRY

15/02/2024

RPZ3S

14140.12

6414.30

4766.55

2005

-90

58 to 138

4708.55 to 4628.55

Qal

DRY

DRY

15/02/2024

RPZ3W

14124.26

6741.97

4778.56

2005

-90

58 to 138

4720.56 to 4640.56

Qal

DRY

DRY

15/02/2024

Source: LRE, 2026

(a)

Feet above mean sea level; for wells, elevation of land surface at surface casing; for VWPs elevation of

surface casing at land surface is provided;

(b)

Degrees from horizontal at bottom of well or depth of VWP along inclined borehole using IDS survey

(c)

Feet below land surface for wells; feet along inclined borehole for VWPs based on IDS inclination survey and

Leapfrog Geologic Model positioning

(d)

Feet below land surface for wells

(e)

Feet below land surface for wells; depth to water subtracted from collar elevation

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15.4.3 Underground Contact Water Management

Contact water discharge volume reached a maximum of 1,400 gpm when mining the South Pacific exploration drift started in late 2024 (Figure 15-4). The current groundwater model predicts inflow to the underground workings will decrease from a maximum of approximately 1,500 gpm to a minimum of approximately 500 gpm as the cone of depression created

by the surface dewatering wells reduces the hydraulic head on the underground workings (Section 10.4.6). (Spheros Environmental 2026).

Source: OMC, 2026

Figure 15-4: Granite Creek Weekly Average Passive Infiltration

Water encountered in underground excavations is collected in a series of sumps excavated along the main decline. Water from the sumps is pumped

from the mine via two six-inch diameter SDR 11 HDPE pipelines routed through the primary development to the surface (Figure 15-5). Sumps eight through ten serve as the

main collection point for contact water.

Expansion of the contact water managements system to handle the anticipated inflow increase will

add a 250-HP submersible pump in Sump 10, pumping into a 12-inch diameter discharge line. A series of three 450-HP booster pumps

will be located at appropriate intervals along the discharge line to overcome static head and friction losses. All pumps will be powered by variable frequency drives. The underground contact water system expansion is estimated to cost

US$2.79 million (Jones 2026).

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Source: OMC, 2026

Figure 15-5: Underground Contact Water Collection and Discharge System

15.4.4

Water Treatment and Rapid Infiltration Basins

Water from the dewatering wells that is not utilized for operations is currently discharged to Rapid Infiltration Basins (RIBs) on the east

side of the Getchell Mine Road through HDPE pipelines. Two of the four permitted RIBs (NEV2005102) have been constructed to date, with discharge to one of the two cells at any given time (Figure 15-6). When

RIB maintenance is required, discharge is routed to the dormant cell. Current dewatering efforts are well under the permitted 6,900 gpm threshold of the RIBs and the RIB infiltration is sufficiently limiting surface ponding in the active cell. As

dewatering efforts have increased in recent years, i-80 has constructed the remaining two permitted cells (1 and 4) in Q2 2026. This will increase the surface area of the infiltration basins, allowing for

proper future operations and cycling between cells for maintenance as the produced volumes increase with planned dewatering infrastructure upgrades.

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A portable rental water treatment plant capable of 800 gpm is in use currently to treat water

from three wells (APW-1, BPW-5 and GCW-6) which does not meet the minimum required water quality standards for direct discharge

to the RIBs (Figure 15-7). The Mag Pit serves as a diversion route for off-spec plant water but is limited by continuous monitoring and adjustments within the closed

loop plant system.

Due to the depth and planned production rates of wells GCW-14 and GCW-15, a second plant has been designed to treat approximately 3,300 gpm and is planned to begin operating in Q3 2026(Figure 15-8). The original plant will also be available

for treatment, allowing for a combined total treatment capacity of approximately 4,100 gpm.

Source: Forsgren, 2026

Figure 15-6: Map of Current and Planned RIBs at Granite Creek

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Source: Practical Mining, 2026

Figure 15-7: Existing Water Treatment Plant (800 gpm capacity)

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Source: Forsgen, 2025

Figure 15-8: WTP-2 Process Flow

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15.4.5

Fresh Water Supply

WW-8, east of Getchell Road, has historically been utilized as a potable water well for the Project

(Figure 15-3). The well is completed in basin alluvial deposits to a depth of 5,177 m (80 ft) and can produce approximately 60 gpm. The well is currently not in use but may be utilized as a potable water

source if needed for future operations.

15.5

Electrical

The Granite Creek project is connected to NV Energy’s 120kv transmission line that parallels the Getchell Mine Road. The Main substation

is located adjacent to the mine access road and has a rated capacity of 5,000 kva. The current substation capacity is sufficient to support the project through to the completion of underground mining (Table

15-3). Power is distributed throughout the site at 13.8 kv (Figure 15-9). Pad mounted or pole mounted transformers are located adjacent to each dewatering well, water

treatment plant, backfill aggregate crushing plant and the mine portal. The underground mine is supplied by a single 4.16 kv feeder entering at the main portal. A second 13.8 or 4.16 kv underground feeder is planned northeast of the portal that will

enter the mine through a borehole or the planned ventilation raise.

Substation repairs and main transformer replacement were initiated in

Q1 2026 following a main transformer fire. As of the effective date of this report the substation has returned to operation.

Table 15-3: Projected Electrical Demand by Area

Area

kva

Dewatering Wells

700

WTP1

70

WTP2

490

Existing Underground Feed

1,100

Future Underground Feed

1,100

Portal Facilities

680

Admin, Dry and Offices

30

Miscellaneous

70

Total

4,240

Source: Woolever, 2026

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Source: OMC, 2026

Figure 15-9: Granite Creek Electrical Distribution and Dewatering System

15.6

Portal Facilities

The mine is accessed through either of two portals, one providing fresh air intake and the other exhaust. Existing portal facilities also

include an equipment maintenance shop, fueling area with diesel and oil storage, contractor offices, and electrical switchgear supplying power to surface facilities and the 4.16kv underground feeder (Figure

15-10).

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Source: OMC, 2026

Figure 15-10: Granite Creek Portal Facilities

The lined fuel and oil storage area contains a double-lined 10,000-gallon diesel tank, a 500-gallon gasoline tank, minor amounts of lubricants and oils, and a waste oil holding tank.

The mine

contractor supplies a backfill plant consisting of a cement silo, a colloidal grout mixing tank, pumps, and a mixing sump where aggregate is mixed with the cement grout with a front-end loader. The mine

contractor also supplies a shotcrete mixing plant with a dedicated cement silo and auger mixer.

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15.7

Backfill Aggregate Crusher

Backfill aggregate is mined from the CX West pit and crushed to a nominal 4-inch maximum size. A

contractor is mobilized as needed to mine and crush backfill aggregate. A typical campaign produces 300 to 400 ktons of aggregate.

15.8

Screen Plant

Before transporting mineralized material to the process plant, OMC removes material greater than nominal four inches using a diesel-powered

screen plant located on a lined stockpile area near the CX pit ramp. The grade of the course material removed is below the process cutoff grade and is sent to the waste rock disposal area. Figure 15-11 shows

the general arrangement of the lined pad and screen plant.

Fuel and oil storage for surface support equipment is also staged on the lined

screen and stockpile pad. It consists of a double-lined 10,000-gallon diesel tank, a 500-gallon gasoline tank, minor amounts of lubricants and oils, and a waste oil

holding tank.

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Source: OMC, 2026

Figure 15-11: Screen Plant Area General Arrangement (OMC 2026)

15.9

Waste Rock Disposal (WRD)

The infiltration basin located at the bottom of the CX Pit (Figure 15-10) must remain in service until

commissioning of WTP-2 is complete. During that period mined waste rock will be disposed of in the C pit.

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Mined waste over the Life of Mine (LOM) will be placed in the bottom of the CX Pit and the C

Pit as shown in Figure 15-12. The Waste Rock Disposal (WRD) plan has sufficient volume for all waste rock mined over the current LOM plan (Table 15-4).

Table 15-4: WRD Capacity

WRD Phase

Color

Capacity

(tons)

Capacity

(tonnes)

CX Pit

Green

475,000

430,912

C Pit

Blue

700,000

635,029

Total WRD Capacity

1,175,000

1,065,941

LOM Plan Waste

660,000

598,742

Excess WRD Capacity

515,000

467,200

Source: OMC, 2026

Waste rock from the underground mine does not contain significant sulfide minerals and is dispersed among an abundance of oxidizing material.

The high acid neutralization potential of the waste stream facilitates disposal in unlined waste rock facilities.

WRD in the CX pit will

be placed in 20 to 40-foot lifts across the CX Pit. C Pit WRD will be dumped from the 5200 elevation in a single lift approximately 76 m (250 ft) high. The WRD slope is the natural angle of repose and is

estimated to be 33 degrees. The in-place density for both dumps is planned at 0.058 tons/ft3.

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Source: OMC, 2026

Figure 15-12: C (Blue) and CX (Green) Pit WRD Plan

15.10

Waste Disposal

Granite Creek does not have an active permitted Class III landfill. A contractor transports household and

non-hazardous waste to an authorized landfill located offsite. Hazardous wastes are collected in the appropriate manor and temporarily stored onsite until removed and transported to an approved disposal site

for the type of waste being disposed of.

The administration offices and dry are serviced by a septic system installed in the early

1990’s. The system consists of a 4000 gal septic tank and leach field. The septic tank is pumped annually. Portable toilets are located in suitable locations across the site. These are serviced regularly by a qualified contractor and the waste

disposed of offsite at an approved facility.

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15.11

Communications

Granite Creek site communications infrastructure is built on a resilient fiber-optic backbone that spans surface and underground areas,

providing high-bandwidth, low-latency connectivity for operational systems, control rooms, and data centers. The fiber network supports enterprise services (voice over IP, CCTV, telemetry, SCADA, and

high-speed data transfer), links surface and underground network nodes, and enables centralized monitoring, automated reporting, and remote access for maintenance and engineering teams. Redundant routing and diverse-path cabling are implemented

where possible to minimize single points of failure and to preserve critical communications during maintenance or incidents.

Wireless

coverage is provided by UHF radio across the surface areas and VHF radio throughout the underground workings. Surface UHF systems deliver clear, mobile voice communications for haulage, processing, and surface crews and integrate with dispatch and

incident response systems. Underground VHF channels are optimized for penetration in tunnels and confined spaces, ensuring reliable miner-to-control-room voice

communications, emergency signaling, and personnel tracking support. Both radio systems are tied in to the fiber network and dispatch consoles, so voice, data, and alarms are consolidated for situational awareness; routine testing, battery/spare

management, and periodic coverage surveys maintain reliability and safety compliance.

Additionally,

i-80 Gold Corp has a FEMCO, a simple, rugged mine-phone system, that provides a dedicated copper-wire voice circuit to metal handset phones installed at key surface and underground locations (shafts, drifts,

refuge stations, pump rooms, crushers, access points and control rooms). The system is intentionally low technology: line-powered drop lines feed magneto-style or dry-cell metal telephones that ring locally

and require no active switching equipment in the handset, ensuring survivable voice communications during power outages and after blast events. FEMCO trunks terminate at protected junction boxes and a small central exchange/gateway that provides

console access for dispatch and cross-connects into the primary fiber backbone and radio dispatch consoles. Typical design features include corrosion-resistant, armored copper cable, sealed metal enclosures, surge/protection modules at entry points,

labeled station circuits, and redundant routing of critical drops where practical. Operational practices include periodic loop and continuity testing, battery/dry-cell replacement schedules (or line-power

verification), maintenance logs for handset and cable condition, and weekly functional checks of gateway-to-dispatch connectivity and ringing. Limitations and

mitigations are: FEMCO delivers highly reliable voice-on-wire but has no native data, CCTV, or IP functionality, i.e., it cannot use the gateway to bridge alarms or

status signals to SCADA/CCTV or carry an isolated emergency voice conversation into the enterprise systems. The use of the FEMCO system is included in emergency response plans by ensuring clear signage of phone locations and FEMCO maintenance with

fiber and radio testing to preserve end to end communications resilience.

15.12

Equipment

Mobile and fixed equipment additions planned over the LOM include the following:

•

Front end loader

•

Cat 745C articulated haul truck (leased)

•

Air compressor

•

Portal truck weigh scale

•

Back-up generator

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16

Market Studies

Ore will be predominantly processed at the Lone Tree processing facility where the final product will be doré, which will be refined

through a certified refinery (North American or global) and sold to any one of several financial institutions at the spot gold price on public markets. There is also a small revenue stream from gold-bearing carbon fines and slags, which are

byproducts of the doré production process. These products are purchased by third parties at the spot price at contracted payable rates with associated fees for treatment.

Prior to refurbishment and restarting up of the Lone Tree facility, the material will be processed via a third party. A toll milling agreement

is currently in place, governing the terms of the processing of Granite Creek sulfide material when being processed by the third party. Oxide material from Granite Creek will be sold to a third party until Lone Tree is operational.

16.1

Market Information

Market Demand

A

report (World Gold Council, 2025) by the World Gold Council in November 2025 summarized the global demand for gold as follows:

“Economic, social and geopolitical trends have contributed to fundamental shifts in demand for gold. Across the globe, rising tensions

and financial market uncertainties are fueling interest among institutional investors and retails savers, supported by increased ease of access to physical gold and gold-backed funds. Among central banks, there has been sustained gold buying since

the Global Financial Crisis, accelerating still further in recent years. In China and India, rising wealth has fostered retail appetite for bars, coins and jewelry, reinforcing gold’s cultural and financial significance across Asia. In

industry, gold’s durability, conductivity and malleability make it a core component of AI infrastructure and other advanced technologies.”

The distribution of demand for gold is illustrated in Figure 16-1. As the demand for gold is high, and

the product is in a highly refined form, the gold produced at Lone Tree from the Granite Creek Project is not exposed to material market risk in terms of its ability to sell products or receive competitive terms. The toll milling contract is

effective until December 31, 2027, at which point it is expected that the Lone Tree processing facility will be operational and capable of processing the refractory ore from Granite Creek.

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Source: World Gold Council, 2025

Figure 16-1: Average Annual Demand for Gold – by Sector

Gold Price Assumptions

The selection of gold prices for the purpose of delineating Mineral Resources and Mineral Reserves is described below.

Figure 16-2 shows the daily spot gold price since January 2020, along with the three-year trailing

average. At the end of July 2026, the spot gold price was US$4,042/oz and the three-year trailing average price was US$3,099/oz.

Source: i-80, 2026

Figure 16-2: Daily Spot Gold Price

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In accordance with industry accepted practices, higher metal prices are used for the MREs

ensuring the mineral Reserves are a subset of, and not constrained by, the Mineral Resources, and which satisfies the test of reasonable prospects for economic extraction.

For the Project, the following gold prices were assumed:

•

US$3,000/oz for Mineral Resources

•

US$2,750/oz for the economic model

•

US$2,500/oz for Mineral Reserves

16.2

Contracts and Status

Dore and Byproducts

i-80 Gold has treatment and refining contracts in place with their respective service providers for the

processing of doré and associated byproducts. The terms of these contracts are typical of other gold mining companies in Nevada and globally.

Toll Milling Autoclave Contract

The agreement that i-80 Gold currently has in place allows for the Company to process ore at a

third-party processing facility at a rate of approximately 30,000 tonnes (33,069.3 st) per month provided the material meets the specifications for such material set out in the contract. The Company will be credited with gold recovered from this

material during processing and pay for processing based on a fee schedule of costs required to process the material.

Key specifics of the

contract are as follows:

•

Mass of material will be determined by net truck weights of material delivered and moisture content determined by

sampling, monthly

•

Gold grade of material will be determined by the mean of assays of a split sample of each truck delivery, by i-80 and the third-party processor, with the option to use a third umpire sample assay to settle difference outside the splitting limit of 5%, monthly

•

Gold recovery of material from Granite Creek, processed by the third-party processor will be determined from

bench top autoclave tests and CIL bottle roll tests replicating plant conditions, and reconciled to the actual recovery obtained during the time period of processing, monthly

•

Silver production credit will be allocated quarterly based on the silver produced and the delivered mass and

grade of material from Granite Creek during that quarter

•

Costs of processing under the Tolling Agreement will be determined monthly as follows:

•

Costs for processing, annual downtime and maintenance, selling, tailings storage and disposal, sustaining

capital and G&A costs are determined for each monthly processing lot

•

Additional costs are incurred if the material is determined to be

non-conforming to the specifications of the contract

•

Tolling charges are determined based on the prevailing gold price for the month of processing and range between

US$53.92 and US$238.88 per tonne of ore processed

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•

Specifications for conforming material include the following:

•

CO3 less than 15%

•

Organic carbon less than 0.5%

•

Sulfide sulfur above 1.0%

•

Moisture less than 8%

•

Preg-rob value less than 40%

•

Maximums on certain constituents including the following:

•

Mercury, 50 ppm

•

Arsenic, 5,000 ppm

•

Lead, 150 ppm

•

Zinc, 200 ppm

•

Total Copper, 0.25%

•

Selenium, 3 ppm

•

Barium, 500 ppm

•

Chromium, 100 ppm

•

Cadmium, 3 ppm

•

Cyanide Soluble Copper, 250 ppm

•

Antimony, 1,500 ppm

•

Free gold, none

•

Iron, 8%

•

Tramp metal and plastic, none

Ore Purchase Agreement

The ore purchase agreement that i-80 currently has with the third party enables i-80 to sell the oxide material produced from the Granite Creek mine. It is purchased by the buyer at a pre-determined payable rate dependent on the gold grade, the sulfide

content and the gold price. The range of payable rates is 40% to 75%, with a minimum accepted grade of 0.075 oz/t, a 3.4 point decrease in payable rate with every 0.1% increase in sulfide sulfur and a maximum accepted sulfide content of 1.0%. The

measurement of trucks and associated samples, along with mill scale readings and samples are used to determine the gold content sold by i-80.

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17

Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups

The site is a producing underground operation built on a historical mine site that has been impacted by operations

and exploration since the 1940’s. The majority of historical surface disturbances have been reclaimed in accordance with state and federal requirements. In the valley to the east of the site, there are several center-pivot irrigation systems

that are utilized for growing alfalfa. These adjacent water users may be beneficially impacted by the mine contributing groundwater to the rapid infiltration basins (RIBs). They are located at distance (and in the opposite direction of the

prevailing wind), thus making them unlikely to be impacted by noise or dust from current operations.

17.1

Environmental Study Results

The original Pinson Mine was seeking authorization from the U.S. Department of the Interior - Bureau of Land Management (BLM) in 1975, just six

years after the promulgation of the National Environmental Policy Act (NEPA). As such, the BLM was required to prepare an Environmental Assessment (EA) analyzing the potential impacts associated with the mine operations. The Environmental

Assessment, Pinson Project, Humboldt County, Nevada, for Lacanex Mining Company, Ltd. (Dames & Moore, 1975) examined the baseline conditions and potential environmental impacts for a number of resources, including: socioeconomics,

geology, soils, hydrology and hydrogeology, ecology, meteorology, and archeology.

A subsequent EA was prepared in 1992 (Environmental

Management Associates, 1992) for an expansion of the Pinson Mine which examined: physiography, geology, soils, hydrology and hydrogeology, air quality, vegetation and range resources, wildlife, cultural and paleontological resources, visuals, noise,

land use, and socioeconomics. A third EA was prepared (JBR, 2003) for an amendment to the closure and reclamation plans of the Pinson Mine with respect to the CX Pit and long-term management of discharge of Granite Creek, which was being diverted

through a series of pipes and culverts through the mining area. In addition to a reexamination of the previously analyzed resources, environmental justice was added to the assessment.

Based on the previous analyses conducted on the older Pinson Mine, the BLM found that the amendment to the existing Plan of Operations by

Osgood for the Granite Creek Underground Project (which only has a small surface footprint and nexus to federal public lands) had been sufficiently analyzed to allow for a Determination of NEPA Adequacy (DNA) and approval by the BLM.

The remaining activities associated with the Granite Creek Underground Project are being conducted on private lands under jurisdiction and

purview of the Nevada Division of Environmental Protection - Bureau of Mining Regulation and Reclamation (NDEP-BMRR). The following section discusses more recent environmental studies relevant to the current Granite Creek Project.

Geochemistry

The

site has had limited geochemical characterization throughout its history. Most of the geochemical characterization test work was performed by Water Management Consultants, Inc. (WMC) in 1998. This study involved acid-base accounting (ABA), metals

enrichment by acid-digestion and ICP-MS, and kinetic tests. Fifty-one rock samples were tested statically, and 15 of those samples were selected for kinetic humidity

cell testing (HCT). Samples were selected from a variety of lithologies and locations but were designed to primarily focus on the Mag and CX future pit wall material (WMC, 1998).

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Results from the ICP-MS analysis showed that major

element abundance was controlled by rock type, with calcium abundant in carbonate-bearing rocks, and silica and aluminum concentrations abundant in silici-clastic rocks. Arsenic was found to be elevated in three samples and associated with

hydrothermal deposits. ABA results indicated that the rock had low acid-generating potential (AP). Tested rocks had low sulfur, neutral paste pH, and abundant neutralization potential (NP) that resulted in 49 out of 51 samples being classified as non-acid generating based on the Canada Centre for Mineral and Energy Technology (CANMET) standards of NP / AP >3 (WMC, 1998).

Kinetic cells were run for a minimum of 20 weeks, with some running for a total of 28 weeks. Rates of acid rock drainage (ARD) generation were

tracked weekly, and the change in acidity and alkalinity was used to provide a quantitative estimate of whether the retained alkalinity would outlast the acid generation. Only one cell showed potential for acid generation: an argillite with 0.47%

sulfur. This sample had consistently acidic pHs (4 to 2.2 S.U.) with sulfate in the hundreds to thousands of mg/L range. All other kinetic cells had neutral to basic pH (7 to 9 S.U.), alkalinity between 20 to 40 mg/L, and no quantifiable acidity.

Monthly leachate analyses largely confirmed the weekly results. Metals in leachate were generally within the reference values, but some samples showed elevated levels of antimony and arsenic multiple times after the initial stabilization period.

Additionally, the two cells that were uncertain under the acid-generation calculation showed levels of aluminum, antimony, arsenic, copper, iron, lead, and thallium after the initial stabilization period that were above NDEP Profile I reference

values at the time. While the remaining cells were not over the reference value at the time, the arsenic and antimony reference values were lowered in 2006 to 0.006 and 0.010 mg/L, respectively, which makes all the cells retrospectively over the new

reference values. However, there was not a particular rock type that was consistently exhibiting acid generating or metal leaching potential (WMC, 1998).

Since 1998, the site has performed periodic sampling characterizing waste rock authorized for disposal by backfill to the bottom of the CX pit.

In the sample set taken from 2005 to 2022, theNP / AP ratio of this rock has varied from 2.5 to 568, confirming the presence of limestone layers within the Comus Sediments (see Section 7.0), that will readily neutralize any acid generated from

the dissolution of sulfide minerals . In addition to ABA tests, Meteoric Water Mobility Procedure (MWMP) tests have been performed on the waste rock deposited in the bottom of the CX Pit. The minimum and maximum MWMP concentrations of constituents

in exceedance of NDEP Profile I reference values is tabulated in Table 17-1 (Stantec Consulting Services, 2023).

Table 17-1: MWMP Results of Rock Placed in CX Pit 2005 to 2022

Analyte

NDEP Profile I Reference Value

Maximum

Minimum

Antimony (mg/L)

0.006

0.070

0.001

Arsenic (mg/L)

0.010

2.200

0.015

Nitrate (mg/L)

10

150.0

0.6

Sulfate (mg/L)

500

800

3

Total Dissolved Solids (TDS) (mg/L)

1000

1,900

27

Quarterly sampling of the waste rock for the sample set taken from 2023 to 2025 indicates theNP / AP ratio of

this rock has varied from 12.7 to 62.4, confirming that the materials remain consistent with previous characterization data. Additionally, MWMP tests have been conducted on the waste rock deposited in the bottom of the CX Pit. The MWMP results for

the rock placed in the CX pit are tabulated in Table 17-2.

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Table 17-2: MWMP Results of Rock Placed in CX Pit

2023 to 2025

Analyte

NDEP Profile I

Reference Value

Maximum

Minimum

Antimony (mg/L)

0.006

0.601

<0.003

Arsenic (mg/L)

0.010

1.3

<0.005

Nitrate (mg/L)

10

57.4

0.2

TDS (mg/L)

1,000

1,970

300

Note: Data was not available to calculate the weighted average. The geometric mean and mean were calculated for

comparative purposes.

The project, with the current quarterly geochemical data, is consistent with the historical assessment that the

backfilled rock does not appear to pose an ARD risk and only appears to pose minimal metal leaching (ML) risk in regard to antimony and arsenic release.

Onsite Water Quality

Water is sampled from many sources:

•

Underground dewatering wells (APW1, BPW3, BPW5, GCW-06)

•

Background groundwater wells

•

Underground mine sumps

•

Contact Water

•

Surface water

•

Mag Pit lake

•

RIBs

•

Influent and effluent from the WTP

Bedrock groundwater from the underground dewatering wells has variable chemistry, which is reflective of the variable groundwater flow between

different mineralized and unmineralized geologic units. In general, the groundwater from unmineralized blocks has lower arsenic and antimony than from mineralized blocks. The pH of bedrock groundwater ranges between 6.8 to 8.4 S.U., with TDS from

180 to 1,500 mg/L and alkalinity between 74 and 134 mg/L. Some bedrock groundwater also exceeds the NDEP Profile I reference values for some metals, particularly for arsenic, cadmium, iron, manganese, nickel, and zinc (Osgood Mining Company LLC,

2023).

The site has maintained continuous bedrock groundwater quality monitoring for the purposes of compliance with NDEP-BMRR Water

Pollution Control Permits (WPCP). A summary of the most recent water quality data for antimony and arsenic is compiled in Table 17-3.

Table 17-3: Water Quality April 2023 to January 2025

Location ID

Antimony Ranges (mg/L)

Arsenic Range (mg/L)

APW1

0.0025 – 0.0041

0.029 – 0.059

BPW3

0.0025

0.02 – 0.027

BPW5

0.0025

0.012 – 0.037

GCW-06

0.0035 – 0.0085

0.33 – 1.0

RIB Distribution Pipeline

0.0025

0.023 – 0.028

WTP Effluent

0.0025 – 0.003

0.005 – 0.015

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Most natural water meets NDEP Profile I reference values for antimony, though there are

occasional exceedances. The majority of bedrock groundwater onsite exceeds NDEP Profile I reference values for arsenic; however, the WTP effluent water quality demonstrates the WTP’s ability to meet arsenic reference values. An arsenic

attenuation study is ongoing to address the high arsenic concentrations found in most groundwater.

Underground mine water is most impacted

by metal leaching with analyzed samples collected from April to August 2024 exhibiting NDEP Profile I reference value exceedances for arsenic (ranging from 0.03 mg/L to 0.375 mg/L), antimony (ranging from 0.015 mg/L to 0.115 mg/L) and thallium

(ranging from 0.003 mg/L to 0.014 mg/L). Range-front background alluvial groundwater quality has historically been relatively consistent over the period of record, with few exceedances of NDEP Profile I reference values, and some indication that

natural chemical attenuation is occurring (Enviroscientists, Inc. Water Management Consultants, 2005). Alluvial groundwater on the project is monitored by numerous wells and is well understood. Onsite alluvial groundwater generally meets NDEP

Profile I water quality reference values, with most trace metals at or below analytical laboratory limits (Osgood Mining Company LLC, 2023).

Granite Creek is located adjacent to the site and flows ephemerally during the spring and summer in response to snow melt and precipitation

events. It is currently diverted through a series of pipes and culverts around the southern rim of the CX Pit to the original stream channel location downgradient of the pit. The water quality is consistently good, with all constituent

concentrations below the NDEP Profile I reference values for surface water (Osgood Mining Company LLC, 2020).

The Mag Pit lake water

quality has been monitored consistently since 2015. Samples have been taken from the top, middle, and bottom of the water column to establish any chemical differences in water quality with depth. Over 10 years of sampling, the average arsenic

surface concentration of the Mag Pit is 0.029 mg/L, the middle of the water column in the Mag Pit has 0.029 mg/L arsenic, and the bottom of the Mag Pit has 0.032 mg/L arsenic. All layers show consistently high TDS (around 1,000 mg/L) and sulfate

(490 to 580 mg/L). The bottom of the Mag Pit also appears to be elevated in manganese up to maximum of 0.53 mg/L in 2021 (LRE Water, 2024).

An Ecological Risk Assessment (SRK, 2025) conducted using the historical Mag Pit lake water quality results concluded that harmful effects to

potential receptors, as a result of exposure to existing water in the Mag Pit lake, are not likely.

Pit Lake Future Water

Quality

The current mine plan assumes that the CX Pit will be backfilled and will have a pit lake. LRE work has supported the

post-closure formation of the CX Pit lake. An investigation of future CX Pit wall material and backfill was conducted by WMC in 1998. This was followed up by a pit lake model to predict future pit lake water quality. WMC found that the majority of

future pit wall rock were acid-neutralizing, with paste pH >7 S.U. and average sulfur content of 0.053% weight. Most rock had significant neutralization potential, with an average of 115 tons of CaCO3 per 1,000 tons of rock.

The lake will behave as a hydrologic sink with no discharge

of impounded waters to the surface or groundwater. The waste rock backfilled to the bottom of the CX Pit will be inundated by rising post-mining pit lake waters to an estimated minimum depth of approximately 9 m (30 ft). Inundation of the backfill

will cut off the oxygen supply and reduce or eliminate the potential for the backfill to generate acidic conditions. Pit lake predictive modeling reports indicate that long-term post-mining lake water will be in compliance with WPCP NEV2005103, with

many metals concentrations less than the analytical laboratory reporting limit, significantly below Nevada reference values except arsenic (0.019 mg/L).

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Ongoing work from LRE (unpublished as of the effective date) addresses the post-closure

formation of a lake in the Mag pit which shows that it will also be a hydraulic sink. The results of post-closure water quality model simulations (Piteau, 2023) indicate that no parameters are likely to exceed NDEP Profile III reference values

(established to assess post-closure pit lakes in the State of Nevada) over the course of the 100-year simulation time frame. The CX Pit lake is predicted to be slightly alkaline with TDS on the order of 1,612

mg/L to 4,290 mg/L.

In the event that post-mining arsenic concentrations rise above acceptable levels, the modeling predicts that the

addition of ferric sulfate solution (Fe2(SO4)3) at a rate of 0.23 grams per

gallon would reduce the concentrations of arsenic to below the analytical laboratory reporting limit in the short and long term. The model result is well supported by the ferric sulfate dosing program that was tested at the CX Pit lake in 2001 to

help reduce arsenic concentrations reported in the pit lake at that time (Beale & Feehan, 2005).

Known Environmental

Issues

SRK is not currently aware of any known environmental issues that could materially impact the Osgood’s ability to

extract the Mineral Resources or Mineral Reserves of the Granite Creek Underground Project.

However, there does exist an environmental

issue that may impact project economics. Petroleum Contaminated Soils (PCS) were discovered in 2025 within the CX Pit waste rock backfill material. While Osgood estimates that only about 290 to 320 cubic yards of PCS material was disposed of in the

backfill, it is not known precisely where that material was placed, resulting in the possibility that most, if not all, of the 1.17 million tons of waste rock will need to be rehandled and sampled for contamination. Osgood is currently working

with the NDEP-BMRR on the preparation of a PCS Management Plan and remedial efforts to address this issue.

17.2

Requirements and Plans for Waste and Tailings Disposal, Site Monitoring, and Water Management During

Operations and After Mine Closure

Environmental management plans have been prepared as part of the state and federal

permitting processes authorizing the Granite Creek Project. Requisite state environmental management plans which are part of the WPCP and Reclamation Permit programs under Nevada Administrative Code (NAC) 445A.398 and NAC 519A.270, respectively,

include:

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•

Fluid management plan

•

Monitoring plan

•

Emergency response plan

•

PCS management plan (pending)

•

Temporary and seasonal closure plans

•

Tentative plan for permanent closure

•

Reclamation plan

Federal environmental management plans incorporate many of the same plans as are required by the State of Nevada; these are specified in Title

43 of the Code of Federal Regulations Part 3809.401(b) (43 CFR § 3809.401(b)) and include:

•

Water management plan

•

Rock characterization and handling plan

•

Spill contingency plan

•

Quality assurance plan

•

Reclamation plan

•

Monitoring plan

•

Interim management plan

The state environmental management plans were originally submitted to the NDEP-BMRR as part of the WPCP and Reclamation Permit applications.

Most of the management plans were updated and resubmitted as part of amendments to the Granite Creek Underground Project.

Ore

Handling

No onsite processing occurs at the Granite Creek Underground Project at this time. All mineral ore mined at the site is

transported offsite for processing under agreement at other permitted facilities. Ore-grade mineralized material encountered during exploration activities is transported to the surface, crushed or remain as

run of mine material, and stockpiled in a temporary stockpile on the mineralized material stockpile pad or clay-lined ore storage pad. The stockpiled material is then transported offsite to a permitted precious metal processing facility for recovery

of precious metals. Transport for offsite processing is contracted to a locally based trucking firm.

Waste Rock Management

Non-ore rock (waste rock) is transported from the underground by trucks to the surface. Under the

terms and conditions of the WPCP (NEV2005103), placement of non-ore rock is authorized for the placement into CX and C pits. The non-ore rock from the CX decline has

been, and will continue to be, end-dumped over the edge of the 4760 bench towards the bottom of the CX Pit.

There are no plans at present to commence construction of a decline originating in the southwest highwall on the 4700 bench of the Mag Pit.

However, if such construction is initiated at some point in the future, the Mag decline may produce non-ore rock which may also be placed in the approved but not yet constructed engineered waste rock dump

located southeast of the existing Mag Pit or within the Mag Pit on private lands. Non-ore rock will be dumped to construct ten-foot lifts built with an overall slope of

three horizontal to one vertical (3H:1V) to facilitate reclamation. The non-ore rock will be dumped at the angle of repose. To eliminate run-on and minimize potential

surface water degradation, the non-ore rock will not be constructed within ephemeral drainage channels.

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The WPCP requires monthly and/or quarterly monitoring of

non-ore rock, as well as non-ore rock specifically originating in the CX Underground Development and ore. The non-ore rock

sampling program at the Project includes collection of composite samples of non-ore rock and ore mined during the quarter and analysis by a State-certified analytical laboratory for NDEP Profile I constituents

via MWMP testing, and static ABA to assess the potential for the non-ore rock to generate acid. Results of the required periodic monitoring will be analyzed to ensure protection of the waters of the State, and

alert Osgood to the need, if necessary, to alter non-ore rock handling and storage procedures.

Tailings Management

See Item 15 (Infrastructure), Tailings Disposal, for the Project’s tailings disposal status.

Water Management

Mine Dewatering

Mine dewatering is discussed in Item 15 (Infrastructure), Power, Water and Pipelines.

Water Treatment Plant

Treatment of dewatering water is discussed in Item 15 (Infrastructure), Power, Water and Pipelines.

An additional WTP (a modular twin of the existing system) will be designed and built to accommodate the greater water disposal needs of the

project.

Stormwater

As required by NAC 445A.433(1)(c), stormwater from the watershed upgradient of the Project is diverted around the Project utilizing the Granite

Creek ephemeral drainageway, including twin 8-inch diameter HDPE pipes through which surface water flows for a time generally during the first two quarters of the year and in relation to snowmelt occurring in

the bowl near the top of the Osgood Mountains west of the Project. Osgood continues to maintain the diversion and monitor the flow as required by the Permit to ensure the integrity and adequacy of the diversion system. The Mining Stormwater General

Permit for Facility ID MSW-266 will remain active until the facility is permanently closed, reclamation is completed, and bonding is released, thus ensuring continued erosion and sedimentation control.

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Monitoring Requirements

Monitoring of the Granite Creek Mine / Project is accomplished on multiple levels and across various regulatory programs; these include:

•

Water Quality pursuant to NAC 445A.398(3)(a) and (c) and WPCPs, including baseline and operational, surface

and groundwater water quality, ponds, pit lake (as applicable), stormwater, etc.

•

Waste Rock generated during the quarter will be analyzed for meteoric water mobility procedures (MWMP), NDEP

Profile I parameters, and geochemical properties.

•

Air Quality and emissions monitoring through the Surface Disturbance Permit and Class II Air Quality

Operating Permit.

•

Surface disturbances, reclamation, and revegetation monitoring through the plan of operations and reclamation

permit.

•

Stormwater flows in Granite Creek and inspections of the facility, including Best Management Practices (BMPs),

structural controls, and high-risk, non-contact offsite discharge areas.

Currently, the site has been executing all environmental monitoring requirements required to maintain the permits associated with the

underground operation. All permit monitoring requirements are up to date.

Human Health and Safety

Public safety will be maintained throughout the life of the Project. All equipment and other facilities will be maintained in a safe and

orderly manner. The Project area will remain fenced during the life of the Project to ensure public safety. Site access will be restricted to employees and authorized visitors by not allowing unauthorized visitors into the Project Area. The site has

a plan for responding to emergencies. The intent of this plan is to outline and establish responsibilities and guidelines for actions to be taken by Project personnel in the event of a spill or pollutant release on the Project property.

17.3

Project Permitting

The Granite Creek Project includes both public and private lands within Humboldt County, Nevada. Therefore, the project falls under the

jurisdiction and permitting requirements of Humboldt County, the State of Nevada (principally the various bureaus within the NDEP), and federally through the BLM (though a limited nexus to federal jurisdiction exists at this time). Table 17-4 presents the list of permits and authorizations under which the Osgood Mining Company is currently permitted to carry out mining operations and reclamation activities at Granite Creek.

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Table 17-4: Granite Creek Mine Project Permits

Permit Name

Agency

Permit Number

Plan of Operations Granite Creek Mine Project

BLM

NVNV105956928 (Legacy Serial # NVN-064101)

Class II Air Quality Operating Permit

NDEP-BAPC

AP1041-3086.02

Mercury Operating Permit to Construct

NDEP-BAPC

MOPTC AP1041-3089 (de minimis)

Water Pollution Control Permit - Rapid Infiltration Basins

NDEP-BMRR

NEV2005102

Water Pollution Control Permit - Granite Creek Mine

NDEP-BMRR

NEV2005103

Granite Creek Mine Nevada Reclamation Permit

NDEP-BMRR

0047

Granite Creek UG Mine Nevada Reclamation Permit

NDEP-BMRR

0242

Mining Stormwater General Permit

NDEP-BWPC

NVR300000: MSW-42365

Onsite Sewage Disposal System

NDEP-BWPC

GNEVOSDS09S0177

Hazardous Materials Permit

Nevada State Fire Marshal

133539 (renews annually)

Waters of the United States Jurisdictional Determination

USACE

Approved Jurisdictional Determination (AJD) received from USACE November 19, 2025 (valid until November 19, 2030)

Water Rights

The Nevada Division of Water Resources (NDWR) is responsible for quantifying existing water rights, monitoring water use, and distributing

water in accordance with:

•

Court decrees

•

Reviewing water availability

•

Reviewing the construction and operation of dams (among other regulatory activities)

Water appropriations are managed through the NDWR and the State Engineer’s office; this is important to the Granite Creek Project because

the Kelly Creek Area hydrographic basin (Area No. 0) in which the operations are located, has been designated through NDWR Order No. O-536. Groundwater basins are typically designated as needing increased

regulation and administration by the State Engineer when the total quantity of committed groundwater resources (water rights permits) approach or exceed the estimated perennial yield (average annual groundwater recharge) from the basin. By

designating a basin, the State Engineer is granted additional authority in the administration of groundwater resources within the designated basin. Designation of a water basin by the State Engineer does not necessarily mean that the groundwater

resources are being depleted, only that the appropriated water rights exceed the estimated perennial yield.

Osgood Mining Company,

LLC’s water rights at the Granite Creek Mine have a total combined duty of 3,746 ha (9,256 acre)-feet annually (AFA), of which 1,149 AFA is for consumptive use (Osgood, 2024). Water usage for the Project is managed via three certificated water

rights (798 AFA) and 12 permits (8,458 AFA), three of which are block permits. All water rights are subject to State Engineer’s Order 1087 (Block Order). An additional 1,364 AFA are currently ‘Ready For Action’ by the NDWR under

permits 94443 and 94864. All use from the mine, including consumptive and non-consumptive use, is reported monthly on a site pumpage report and the specific meter readings are recorded and subsequently

uploaded monthly to NDWR’s online meter database.

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Current Permitting Activities

There are currently no pending permitting actions before the NDEP nor the BLM with respect to the Granite Creek Underground Project.

Performance or Reclamation Bonding

Pursuant to state and federal regulations, any operator who conducts mining operations under an approved plan of operations and/or reclamation

permit must furnish a bond in an amount sufficient for stabilizing and reclaiming all areas disturbed by the operations. Bonding amounts are based on the operator complying with all applicable operating and reclamation requirements as outlined in

the regulations at 43 CFR § 3809.420 and NAC 519A.350 et seq. Section 17.5 provides additional details on the currently calculated financial assurance estimate for the operations. Current reclamation cost estimates (RCEs) will

remain in effect until updated as part of a state and federal permitting action (pending) or next 3-year bond review (anticipated in 2028 and 2029).

17.4

Local Individuals and Groups

Osgood Mining currently has no plans, formal negotiations, or executed agreements with local communities or individuals as part of the

technical and economic factors necessary to demonstrate reasonable prospects for economic extraction of Granite Creek Project Mineral Resources.

Stakeholder Engagement

The following information on community relations and stakeholder consultation has been provided by Osgood Mining personnel (2026).

Mining activities at the Granite Creek property (formerly known as the Pinson Mine) date back to the 1940s, with intermittent periods of

operation continuing to the present day. Throughout its history, the operation has played an important role in the economic development of Humboldt County by creating employment opportunities and generating both direct and indirect economic benefits

for the region.

Osgood Mining Company periodically hosts Town Hall meetings in Winnemucca, Nevada, to provide local stakeholders with

operational updates and opportunities for open dialogue. The company also partners with the Nevada 95-80 Regional Development Authority during its annual economic development conference and collaborates with

the Humboldt County School District, Great Basin College and the Mining Industry Foundation, to support educational initiatives that strengthen opportunities for local students and the broader community.

Beyond these partnerships, Osgood Mining places a high priority on maintaining positive, long-term relationships with local government

officials, ranchers, neighboring landowners, and other community stakeholders, ensuring their perspectives are heard, respected, and considered throughout the development process. Through these ongoing efforts, Osgood Mining strives to be a

responsible and trusted community partner, supporting the long-term well-being of the region and its residents throughout the life of the Granite Creek Project.

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17.5

Mine Closure

Closure Planning

Mine closure and reclamation requirements in the State of Nevada are addressed on several levels and by several authorities:

•

Federal requirements are generally covered in the plan of operations under the BLM’s 43 CFR §

3809.401(b)(3) which state that, at the earliest feasible time, the operator shall reclaim the area disturbed, except to the extent necessary to preserve evidence of mineralization, by taking reasonable measures to prevent or control on-site and off-site damage to the federal lands.

•

State of Nevada requirements are stipulated in both the WPCP’s tentative plans for permanent closure (TPPC)

and final plans for permanent closure (FPPC) under NAC 445A.396 and 445A.446 / .447, respectively, and the reclamation permit requirements under NAC 519A.

•

The Humboldt County Regional Master Plan outlines broad goals encouraging responsible mining operations that

minimize environmental footprints and protect the county’s natural capital. It supports projects that remain compatible with long term land use planning but relies on external agencies to mandate the actual technical reclamation.

The state closure and stabilization requirements under the WPCP pertain to process and

non-process components (sources), such as mill components, heap leach pads, tailings impoundments, pits, pit lakes, waste rock dumps, ore stockpiles, fueling facilities, and any other associated mine

components that, if not properly managed during operation and closure, could potentially lead to the degradation of waters of the State. A mining facility operator / permittee must submit a TPPC as part of any application for a new WPCP or

modification of an existing permit. A FPPC must be submitted to the agency at least two years prior to the anticipated closure of the mine site, or any component (source) thereof. This plan must provide closure goals and a detailed methodology of

activities necessary to achieve chemical stabilization of all known and potential contaminants at the site or component, as applicable. The FPPC must include a detailed description of proposed monitoring that will be conducted to demonstrate how the

closure goals will be met.

Disturbance on both public and private land is subject to state mine reclamation regulations (NAC 519A).

Reclamation of disturbed areas resulting from activities associated with the facilities will be completed in accordance with NDEP and BLM regulations. The objectives of the reclamation work include the following:

•

Ensure public safety

•

Reduce or eliminate potential environmental impacts

•

Return the site to a condition that will support future use

•

Control infiltration, erosion, sedimentation, and related degradation of existing drainages in an effort to

minimize off-site impacts

•

Employ reclamation practices using proven methods that do not require ongoing maintenance

•

Considering these objectives, reclamation activities are designed to:

•

Stabilize the disturbed areas to a safe condition

•

Protect both disturbed and undisturbed areas from unnecessary and undue degradation

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To the extent practicable, reclamation and closure activities will be conducted concurrently

to reduce the overall reclamation and closure costs, minimize environmental liabilities, and limit financial assurance exposure. The revegetation release criteria for reclaimed areas are presented in the “Guidelines for Successful Revegetation

for the Nevada Division of Environmental Protection, the Bureau of Land Management, and the U.S.D.A. Forest Service” (NDEP, 2016). The revegetation goal is to achieve the plant cover similar to adjacent lands as soon as possible, which, on a

denuded salt playa, is relatively simple.

Mine Closure Design Criteria

The following design criteria and assumptions were applied by Mountain Valley Professionals (MVP, 2026) in the development of the mine closure

cost estimate:

•

All key regulatory reclamation and closure requirements will be met.

•

All regrading of mine waste structures will be performed during operations.

•

An approved rangeland grass mix will be used. Grassland and wildlife habitat with grazing is the anticipated

post-mining land use.

•

All buildings, power lines, and other infrastructure will be removed.

•

The RIBs will be filled in and reclaimed.

•

Granite Creek will be restored to a channel similar to the pre-mining

flow path.

•

The future water quality in the post-closure Mag Pit lake should meet NDEP Profile III reference values (SRK,

2025) under the current closure scenario (i.e., no open pit mining, independent pits at closure). The lake will be a terminal sink for water; therefore, a long-term risk to groundwater quality has not been identified.

•

Other than the issues discussed above, there are no other potential water quality or water quantity impacts at

the Granite Creek Mine upon closure.

Closure Cost Estimate

This current estimate for the reclamation and closure of the Granite Creek Project (US$8.64 million) (MVP, 2026) was prepared in the Nevada

Standardized Reclamation Cost Estimator (SRCE), Version 1.4.1 Build 17c. The SRCE model has been in use since 2006 in the State of Nevada after validation by both state and federal regulators and mining industry representatives. The estimate

utilized a cost data file (CDF) prepared by the NDEP-BMRR, which was released on August 1, 2025. The CDF utilizes the unit rates below:

•

Labor rates from federally mandated Davis-Bacon rates

•

Rental equipment rates quoted from Cashman Caterpillar in Reno, Nevada

•

Miscellaneous unit rates from Nevada mining vendor quotes (e.g., seeding, well abandonment, etc.)

•

Costs for some activities and supplies are from the 2024 RS Means Heavy Construction database (where activities

include labor, they are modified to use the Davis-Bacon wages)

The SRCE model utilizes first principles to calculate

various costs for activities related to mining operations. Inputs for these equations range from equipment efficiencies, labor efficiencies, fuel consumption rates, area calculations, unit rates for labor / equipment / consumables, etc. Some costs

estimated in the SRCE model (such as those for demolition) are estimated based on productivities and crews from the RS Means Heavy Construction database but use the standardized labor and equipment rates included in the CDF. Other, site-specific

costs may be calculated by the operator and included in one of the user sheets.

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The rates for the CDF are supplied by the NDEP-BMRR and vetted for usage in reclamation

estimates throughout the state of Nevada, as well as several surrounding jurisdictions. Davis-Bacon labor rates are based on government contracts with select labor unions and may be higher than those that would be incurred by an operator in a

self-performed closure scenario where in-house or non-union contract labor can be used. The costs within a reclamation estimate prepared for a regulatory agency often

have additional overhead costs related to government oversight of the closure project; the same is true of the values associated with equipment. The rates within the government-prepared CDF are leased rates (which include capital and operating

costs), as opposed to an owner / operator fleet already having a majority of the equipment on hand and partially or fully amortized or potentially easier access to equipment. The bond cost estimate includes 10% for contractor overhead and profit, 6%

for engineering and design, 6% for contingency, 6% for government project management, and 4% for bonding and insurance. The total indirect markup of the reclamation bond estimate is 37%. While this total markup is likely sufficient to cover the

project management and overhead (G&A) costs in a self-performed closure, they are not detailed enough to make a judgement whether they are adequate in this case. Normally, a self-performed LoM closure cost will include a project-specific list of

G&A costs for both management and overhead items like telephones, office supplies, electricity, etc.

Limitation on the Closure

Cost Estimate

The government required reclamation cost estimate was utilized to provide a basis for financial assurance. This type

of estimate reflects the cost that the government agency responsible for closing the site would be responsible for in the event that an operator fails to meet their obligation. If Osgood, rather than the government, closes the site in accordance

with their current mine plan and approved closure plan, the cost of closure is likely to be different from the financial assurance cost estimate approved by the government. There are a number of costs that are included in the financial assurance

estimate that would only be incurred by the government, such as government contract administration. Other costs (such as head office costs, a number of human resource costs, taxes, fees, and other operator-specific costs that are not included in the

financial assurance cost estimate) would likely be incurred by Osgood during closure of the site.

Furthermore, because closure of the site

is not expected until 2032, based on the forecast reserve production plan, the closure cost estimate represents future costs based on current expectations of site conditions at that date. In all probability, site conditions at closure will be

different from currently expected; therefore, the current estimate of closure costs is unlikely to reflect the actual closure cost that will be incurred in the future.

17.6

Adequacy of Plans

Given the robust state and federal regulatory requirements in Nevada and review of the available documentation, it is SRK’s opinion that

the current plans are adequate to address any issues related to environmental compliance, permitting, and local individuals or groups.

17.7

Commitments to Ensure Local Procurement and Hiring

i-80 Gold Corp. maintains specific corporate governance policies for all its operations, including

those at Granite Creek, including providing and supporting access to employment and procurement opportunities for communities and Indigenous groups (i-80 Gold Supply Chain Policy, March 31, 2025).

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18

Capital and Operating Costs

18.1

Underground

18.1.1

Expected Accuracy of Cost Estimates

Estimation of capital and operating costs is essential to the evaluation of the economic viability of a prospective project. These factors,

combined with revenue and other expense projections, form the basis for the financial analysis. Capital (CAPEX) and operating (OPEX) costs for Granite Creek were estimated based on actual data from i-80, the

life of mine plan described above, estimates of materials and labor based on that design, predicted consumption of power and supplies, budgetary quotes for major equipment, labor requirements, and estimates from contractors.

18.1.2

Capital Cost Estimates

The mining capital cost estimate is based on a first principal cost model buildup and actual costs from

i-80. The total capital estimate is US$60.74 million. A 5% contingency, or US$2.77 million, is included in the total CAPEX.

The CAPEX estimate includes lateral and vertical capital development, mine infrastructure, mobile equipment and infill drilling.

Development costs are derived from the mining schedule prepared by SRK. The prepared mining schedule includes footages of development combined

with unit costs and is based on site specific data to estimate the cost of these development activities. The breakdown of the estimated capital costs is shown in Table 18-1.

Table 18-1: Estimated Mining Capital Cost

Item

US$M

Capital Lateral Development Cost

9.34

Capital Vertical Development Cost

12.71

Mine Infrastructure and Equipment

4.75

Rebuilds

0.90

Infill Drilling

30.27

Contingency (5%)

2.77

Total

60.74

Source: SRK, 2024

US$M: US$ million

18.1.3

Operating Cost Estimates

The life of mine operating costs are US$408.13 million. SRK estimated the required mining equipment fleet, required production operating

hours, and manpower to arrive at an estimate of the mining costs that the mining operations would incur. The mining costs were developed from first principles with client inputs on actual cost data and contractor quotes. The mining operating costs

are presented in the following categories:

•

Drilling

•

Blasting

•

Mucking

•

Hauling

•

Backfill

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•

Support Equipment Operations

•

Miscellaneous Operations (various support operations, etc.)

•

Mine Engineering (mine technical personnel and technical consulting)

•

Mine Administration and Supervision (mine and maintenance supervision, etc.)

•

Contractor Equipment Rental

A summary of the LoM unit mine operating costs is presented in Table 18-2.

Table 18-2: Mining Operating Costs

Category

US$M

US$/st

US$/t

Operating Development

157.68

65.07

71.73

Backfill

48.74

20.11

22.17

Truck Haulage

36.16

14.92

16.45

Mine Services and Maintenance

70.53

29.10

32.08

Power

12.15

5.01

5.52

Technical Services and Management

13.07

5.39

5.94

Contractor Equipment Rental

69.81

28.81

31.76

Total

408.13

168.42

185.65

Source: SRK, 2026

US$M: US$ million

Figure 18-1 shows the OPEX cost profile for the life of mine.

Source: SRK, 2026

Figure 18-1: Yearly Mine Operating Cost Profile

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Key inputs and assumptions used to estimate the OPEX cost are as follows:

•

Development and production are performed by contractors. Contractors provide labor and equipment. The Owner

provides the consumables such as diesel, explosives, ground support, etc.

•

Backfill costs assume the haulage and placement of rockfill to the stopes. Costs for operating the backfill plant

and aggregate costs are estimated separately.

•

Material movement assumes haulage of material from the stope to the portal ore pad.

•

Power costs only assume the additional power needed to support the life of mine plan. The current power

requirement is estimated separately.

18.1.4

Closure and Reclamation

Total reclamation costs are estimated at US$8.6M. An additional approximately US$4M is estimated for environmental clean up and reclamation for

a total closure cost estimate of US$12.7M. Reclamation costs are only for the underground mine-related disturbance. Legacy reclamation costs are included in the open pit estimates.

18.1.5

Infrastructure

PMC constructed significant infrastructure prior to i-80’s acquisition of the Granite Creek

Project. Work is ongoing for the completion of RIBs one and four, water treatment plant #2, and wells GCW-14 and GCW-15. Estimated capital spending remaining for the Project as of June 1, 2026, are shown

inTable 18-3. Capital estimates are from existing construction contracts, current and prior purchase orders, and OMC engineering estimates.

Table 18-3: Infrastructure Capital Cost

Estimates1

Area

Capex Remaining1 ($M)

Source

WTP#2

$

6.7

Forsgren

Dewatering Wells GCW14 & GCW15

$

2.6

OMC

Monitoring Wells Rehab

$

1.7

OMC

RIBs 1&4 and Rib Monitoring Wells

$

0.5

OMC

Potable Water Treatment System

$

0.8

OMC

Electrical Distribution

$

0.3

OMC

Equipment

$

1.3

Mobile Equipment

$

1.9

OMC

Light Vehicles

$

1.0

OMC

Total

$

16.6

Source: Practical Mining, 2026

(1) As of June 1, 2026

US$M: US$ million

For purposes

of the economic analysis, Surface Mine Capital also includes approximately US$5.6 million of costs capitalized in connection with underground mine development. These costs are incremental to the US$16.6 million infrastructure capital cost

estimate presented in Table 18-3.

18.2

Surface Operating Costs

Operating costs for the surface infrastructure and support functions were derived from historical costs during the period from January 2024

through May 2026 and engineering estimates as appropriate (Table 18-4).

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Surface labor totals 23 hourly, supervisory and contract personnel tasked with WTP operation

and maintenance, well operation and maintenance, ore waste and backfill aggregate movement to and from the portals, screen plant operation, haul road maintenance, and maintenance of all site buildings and facilities.

Electrical power consumption was calculated from historic billed kwh plus new surface loads planned during the

life-of-mine (Table 18-5). New electric loads added in the underground mine are captured with the underground mining costs.

Existing power demand for the underground mine is included with the infrastructure baseline demand. Power costs are estimated at the historic average of US$0.077/kwh.

Table 18-4: Infrastructure Unit Operating Cost Estimates

Area

Operating

(US$/ore

ton)

Operating

(US$/ore

tonne)

Source

WTP#1

3.83

4.22

Veolia

WTP#2

17.06

18.80

Forsgren

Electrical

8.61

9.49

OMC

Surface Labor

9.56

10.53

OMC

Surface Support Equipment and Supplies

8.67

9.56

OMC

Backfill Aggregate

10.11

11.14

OMC

Total

57.83

63.75

Source: Practical Mining, 2026

Table 18-5: Annual Infrastructure Operating Costs (US$000’s)

Area

2026

2027

2028

2029

2030

2031

2032

2033

2034

Total

Surface Labor

1,356

2,776

2,782

2,777

2,777

2,777

2,784

2,776

2,077

22,883

Surface Support

1,655

2,046

2,822

2,990

2,883

3,094

2,682

1,846

739

20,758

WTP1

889

936

988

1,078

1,112

1,112

1,112

1,112

834

9,171

WTP2

1,965

4,500

4,621

5,042

5,202

5,202

5,202

5,202

3,902

40,839

Electrical

1,117

2,205

2,327

2,538

2,618

2,618

2,620

2,618

1,962

20,623

Backfill Aggregate

1,160

3,223

3,713

3,739

3,745

3,793

3,374

1,084

369

24,200

Total

8,143

15,686

17,252

18,164

18,338

18,596

17,775

14,638

9,882

138,474

Source: Practical Mining, 2026

Operating costs for over-the-road haulage of ore from Granite

Creek to either Lone Tree or third-party processors are part of a contracting agreement with a commercial hauler, with rates as follows:

•

US$5.44 per wet ton for ore transport

•

US$175 per hour for travel time

•

Additional fuel surcharge

Costs for transporting ore to the destinations for processing is shown in Table 18-6, based on historic

2026 invoices received.

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Table 18-6: Unit Costs for Over-the-Road Ore Haulage

Destination

Operating

(US$/ore ton)

Operating

(US$/ore tonne)

Lone Tree

16.21

17.87

Third-Party Ore Processing Plant

7.10

7.83

Source: i-80 Gold, 2026

18.2.1

Autoclave Capital Cost Allocation (Processing Fee)

i-80 plans to operate their mines in a hub and spoke arrangement.

i-80’s plan is to ship ore to the Lone Tree Mill acting as the hub. This is analogous to third-party ore or concentrate processing that is currently being used for ore being mined from Granite Creek

(i.e. Nevada Gold Mines). Given that Lone Tree is not a captive processing facility, i-80 has elected to use the mill as a toll treatment facility both internally and potentially externally. As the Lone Tree

Mill will be operating in a toll treatment capacity, i-80 intends to allocate the Lone Tree Mill’s refurbishment capital cost to each of the mine sites.

The capital cost estimate was developed by an experienced team of engineers, designers and cost estimators from Hatch.

The capital costs for the project have been prepared in accordance with standard industry practices and guidelines provided by the Association

for the Advancement of Cost Engineering (AACE) international recommended practice 47R-11. Using the guidelines provided in practice 47R-11, the cost estimate is

considered to be an AACE Class 3 estimate.

Basis of Estimate

The estimate is inclusive of the entire mill scope from feed ore stockpile to filtered tails material placement in the tailings storage

facility.

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The estimate includes the following key components:

•

Labor costs comprised of unit manhours, productivity adjustments and contractor-provided labor rates.

•

Equipment and bulk material costs for permanently installed equipment and facilities.

•

Costs associated with relocation, replacement and refurbishment items based on the site findings to bring the

process operations up to name plate production.

•

Construction indirect costs for the support and maintenance of the site during the construction period.

•

Other indirect costs such as EPCM, freight, first fills, and spares.

•

Contingency has been applied in accordance with AACE Class 3 Guidelines and a Qualitative Risk Assessment of

the estimate.

•

Owner’s Cost provided by i-80 for inclusion in the capital cost

estimate.

•

Forward escalation has been excluded from the capital cost estimate.

•

All costs are expressed in Q3 2025 United States Dollars.

Estimate Classification

The Mill estimate follows established guidelines for producing a CAPEX according to the Hatch procedures in alignment with AACE estimating

standard practice. The estimate was completed to a high-definition AACE Class 3 estimate with an intended accuracy of -15% to +15%.

Direct Costs

Direct costs

are generally quantity based and include all the permanent equipment, materials, labor, and subcontractor costs associated with the physical construction of the permanent facility / asset.

Discipline Estimates, Quantity Development, and MTOs

The following section outlines the basis for how the engineering discipline cost estimates and quantities were developed by the engineering

team. The cost estimates for all tagged equipment and some contracts were generally developed through the use of technical specifications and the use of firm or budgetary bid packages. The material quantities were developed from the 3D model, new

drawings, and engineering calculations and lists, supported by existing drawings information collected during site visits. The cost estimates for MTOs were obtained using budget pricing.

All engineering input to the estimate was controlled through the use of estimate control sheets and material

take-off forms.

The disciplines involved in direct input to the estimate are as follows: Civil,

Architectural, Structural, Mechanical, Piping, Electrical, and Instrumentation & Controls.

Indirect Costs

Indirect costs include items that are necessary for the completion of the project but are not directly related to the direct construction

costs. These costs are detailed in the section below and were generated by the project management and estimating team with input from i-80.

The indirect costs categories used within this estimate are as follows: Construction Indirects, Vendor Representatives, Spare Parts, First

Fills, Freight and EPCM.

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Capital Cost Summary

The estimated capital cost is presented in Table 18-7. The total Lone Tree Mill Restart project capital

cost estimate is US$430 million USD as of Q3 2025.

Table 18-7:Capital Cost Summary

Capital Cost Estimate Summary

Description

Total Amount (US$M)

Subtotal Direct Cost

252.6

Subtotal Indirect Cost

109.3

Total Direct + Indirect Cost

361.9

Contingency

43.4

Subtotal Cost (Excluding Owner’s Cost)

405.3

Owner’s Cost

25.0

Total Project Cost

430.3

Source: Hatch, 2026

US$M: US$ million

Capital

Cost Conversion to Sustaining CAPEX Cost

As i-80 intends to operate the Lone Tree Mill in a

toll treatment arrangement, the overall capital cost must be converted into a sustaining cost per ton of processed ore. The calculated yearly operating cost for the refurbishment can be attributed to each operating mine based on their throughput.

To convert from a fixed capital cost estimate to an sustaining cost estimate the following assumptions were used:

•

Nameplate capacity for each year of operation

•

Design mine life of 20 years

•

Discount rate set to 0%

•

Cost equally attributed to each ton of ore processed

These assumptions have the potential to have a significant effect on the capital attribution to each year. For instance, should the mill

operate for 25 years the effective cost per ton processed would decrease. In any year when the design tons to be processed are not processed this will inflate the effective operating cost per ton.

The selected assumptions align with the design and are to the best of the Authors knowledge accurate for the operating life of the Mill.

The Financial mine life for the CAPEX conversion was set by i-80 at 20 years. This is in contrast with

the project’s design life of 15 years. This financial mine life was set based on i-80’s operating experience and the benchmarking of the other Autoclave facilities in Northern Nevada. The

expectation is that the sustaining capital and maintenance costs included in the Lone Tree Mill operating cost estimate cover the expected expenses to keep the plant operating for 20 years (Table 18-8).

Table 18-8: Sustaining CAPEX Assumptions for Lone Tree

Description

Units

Value

Capital Cost

US$M Q3 2025

430

Life of Mine

Years

20

Solids Throughput

st / annum

912,500

Design Life Solids processed

Millions of st

18.25

Sustaining Processing Cost

US$ / st

23.56

Source: Hatch, 2026

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The sustaining cost for the refurbishment of the Lone Tree Mill is US$23.56/st. This cost is

applicable to every ton processed at Lone Tree.

Processing Costs – Lone Tree

The operating cost for the Lone Tree Processing Facility has been summarized by component as shown in Table

18-9.

Table 18-9: Lone Tree Processing Facility

Operating Cost Summary

Category

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Power

14.61

13.25

Labor

13.62

12.36

Maintenance

7.86

7.13

Consumables

61.87

56.13

Facility

7.21

6.54

Total

105.16

95.4

Source: Hatch, 2026

Power

The

electrical power requirements are estimated by plant area based on the equipment sizing for all equipment included on the electrical load list. A unit power cost of US$0.10/kWh has been assumed based on NV Energy rate sheet and preliminary contract

with i-80.

Electrical costs are summarized in Table 18-10.

Table 18-10: Lone Tree Pressure Oxidation-CIL Power

Costs

Area

Annual Consumption

(MWh)

Annual Cost

(US$)

Unit Cost

(US$/ton)

Unit Cost

(US$/tonne)

Comminution

26,628

2,662,900

2.92

3.22

POX and Utilities

19,412

1,941,200

2.13

2.34

CIL, Elution and Refinery

11,206

1,120,600

1.23

1.35

Tailings

19,898

1,989,900

2.18

2.40

Ancillary

22,775

2,277,600

2.50

2.75

Oxygen Plant

20,998

2,099,800

2.30

2.54

Total

120,917

12,092,000

13.25

14.61

Source: Hatch, 2026

Labor

Labor costs

include the salaries, taxes, and wages for 90 staff members required to directly support the restart of the Lone Tree processing facility, including administration, operations, maintenance, and technical services personnel. Additional roles required

to support the entire site and the management office at Lone Tree have been captured separately by i-80 and are excluded from the cost estimate.

The organizational structure for plant personnel is shown in Figure 18-2.

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Source: i-80 Gold, 2026

Figure 18-2: Organizational Chart for Lone Tree Pressure

Oxidation-CIL

Labor costs are summarized in Table

18-11.

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Table 18-11: Lone Tree Pressure Oxidation-CIL Labor Costs

Area

Total Headcount

Annual Cost

(US$)

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Shared Personnel

23

3,640,000

4.40

3.99

Comminution

14.3

1,648,000

1.99

1.81

POX and Utilities

10.3

1,208,000

1.46

1.32

CIL, Elution and Refinery

16.3

1,850,000

2.23

2.03

Tailings

26.3

2,936,000

3.55

3.22

Ancillary

0

—

—

—

Oxygen Plant

0

—

—

—

Total

90

11,282,000

13.63

12.36

Source: i-80 Gold, 2026

Maintenance

Maintenance material costs are calculated based on previous experience and similar sized plants. The maintenance costs are estimated as annual

costs based on the partial direct costs from the capital cost estimate. Partial direct costs exclude install labor, architectural, steel, concrete, and earthworks costs.

Maintenance costs are summarized in Table 18-12.

Table 18-12: Lone Tree Pressure Oxidation-CIL Maintenance

Costs.

Area

Annual Cost

(US$)

Unit Cost

(US$/tonne)

Unit

Cost

(US$/ton)

Comminution

366,900

0.44

0.40

POX and Utilities

3,622,000

4.38

3.97

CIL, Elution and Refinery

639,800

0.77

0.70

Tailings

831,800

1.00

0.91

Ancillary

1,049,600

1.27

1.15

Oxygen Plant

—

—

—

Total

6,510,100

7.86

7.13

Source: i-80 Gold, 2026

Consumables

Consumable costs are estimated based on a reagent list prepared using the process flow diagrams and design criteria. Reagent consumption rates

were obtained from the mass balance and used to calculate the annual cost. Most of the consumable unit costs were provided by i-80 via vendor quotes or actual pricing from the currently operating Lone Tree

plant, while others were estimated based on assumptions, online database, and reference projects.

Consumables costs are summarized in

Table 18-13.

Table 18-13: Lone Tree Pressure Oxidation-CIL Consumable Costs

Area

Annual Cost

(US$)

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Comminution

3,253,700

3.93

3.57

POX and Utilities

32,397,500

39.14

35.50

CIL, Elution and Refinery

320,000

0.39

0.35

Tailings

736,600

0.89

0.81

Ancillary

16,347,200

19.75

17.91

Total

53,055,000

64.09

58.14

Source: i-80 Gold, 2026

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The key consumables for the operation and their cost information are shown in Table 18-14.

Table 18-14: Key Consumables Costs for Lone Tree

Pressure Oxidation-CIL

Consumable

Annual Cost

(US$)

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Unit Price

(US$)

Sulfuric Acid (93%)

31,896,400

38.53

34.95

220/tonne

Quicklime

9,638,700

11.64

10.56

354/tonne

Propane

3,703,900

4.47

4.06

0.412/L

Ammonium Bisulfite (65%)

1,122,500

1.36

1.23

0.579/L

Sodium Cyanide

801,600

0.97

0.88

3.06/kg

Source: i-80 Gold, 2026

Facilities

Facility costs include operation of the oxygen plant under the ‘sale of gas’ agreement, mobile support equipment for the plant and

site services including allowances for health and safety supplies and training.

Facility costs are summarized in Table 18-15.

Table 18-15: Lone Tree Pressure Oxidation-CIL Facility Costs

Area

Annual Cost

(US$)

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Oxygen Plant

2,460,000

2.97

2.70

Mobile Equipment

2,175,300

2.63

2.38

Site Services

109,200

0.13

0.12

Total

4,744,500

5.73

5.20

Source: i-80 Gold, 2026

CIL at Lone Tree (Oxide Processing)

The operating cost estimate for oxide processing at Lone Tree was derived from the operating cost estimate for the sulfides with adjustments

made due to bypassing of the autoclave and lower demand on ancillary facilities. The modified operating costs are shown in Table 18-16.

Table 18-16: Lone Tree Processing Facility Operating Cost Summary

Category

Unit Cost

(US$/tonne)

Unit Cost

(US$/ton)

Power

9.30

8.44

Labor

11.11

10.08

Maintenance

3.14

2.85

Consumables

7.09

6.43

Facility

3.34

3.03

Total

33.97

30.82

Source: i-80 Gold, 2026

Third-Party Autoclave Facility Operating Costs

Costs for third-party processing of sulfide material are covered by the toll milling agreement and consist of costs incurred by the third party

for processing, a treatment charge that varies based on the gold price and any penalties for material that does not meet the specifications of the contract.

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18.2.3

General and Administrative Costs

General and Administrative costs for this analysis are based on current Granite Creek actuals and forecasts and are categorized as follows:

•

Environmental - material, licenses, training and professional services

•

G&A – G&A payroll expenses, material, property tax and other general costs

•

Health and Safety - material, licenses, professional services and training

•

Human Resources – recruiting and human resources

•

Information Technology – information technology material, licenses and professional services.

General and Administrative staff at the operation consists of five personnel. Corporate costs are allocated to site

where appropriate when support is provided.

A summary of the average annual G&A cost build-up

is presented in Table 18-17. The LoM total G&A spend is US$49.9 million.

Table 18-17: Average Annual G&A Spend

Area

Unit

Value

Environmental

US$

M

0.7

G&A

US$

M

4.6

Health and Safety

US$

M

0.3

Human Resources

US$

M

0.0

Information Technology

US$

M

0.3

Total

US$

M

5.9

Source: SRK, 2026

US$M: US$ million

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19

Economic Analysis

19.1

General Description

A cash flow model was prepared to evaluate the Granite Creek Project on a real basis. The model was developed on a monthly basis to match the

mine plan generated for this exercise from provided start date through the exhaustion of the Mineral Reserves. The indicative model results are summarized on an annual basis and presented in this section alongside the main assumptions used in the

model. The model results are presented in U.S. dollars (US$), unless otherwise stated. While the model inputs are primarily in imperial units, the model summary and results have been calculated in metric units. For the avoidance of confusion, both

unit systems are provided in the tables in this section.

The economic model is based on mine plans and other inputs that were prepared as

outlined in previous sections.

Capital and operating costs were developed in previous section and the

build-ups and the associated accuracy and contingencies can be found in those sections.

All

results and technical and cost information are presented in this section on a 100% basis reflective of i-80 Gold’s ownership unless otherwise stated.

As with capital and operating cost and pricing forecasts, the economic analysis is inherently a forward-looking exercise. These estimates rely

upon a range of assumptions and forecasts that are subject to change depending upon macroeconomic conditions, operating strategy and new data collected through future study and operation.

Basic Model Parameters

Key criteria used in the analysis are presented throughout this section. Basic model parameters are summarized in Table 19-1.

Table 19-1: Basic Model Parameters

Description

Value

TEM Time Zero Start Date

May 1, 2026

Operations Start

May 1, 2026

Mine Life

8.3 years

Discount Rate

5%

Source: SRK, 2026

All costs incurred prior to the model start date are considered sunk costs. The potential impact of these costs on the economics of the Project

is not evaluated. This includes contributions to depreciation and working capital as these items are assumed to have a zero balance at model start.

Closure costs are assumed to be incurred as a single cost at the end of the mine life.

The selected Project discount rate is 5%, as directed by i-80 Gold.

Foreign exchange impacts were deemed negligible as most, if not all, costs and revenues are denominated in US dollars.

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External Factors

Pricing

Modeled prices

are based on the prices developed in the Market Study section of this report. The prices are modeled as US$2,750/oz over the life of the Project.

All product streams produced by the operation are modeled as being subject to the price presented above.

Taxes and Royalties

As

modeled, the Project is subject to a federal income tax rate estimated at 21%. All expended capital is subject to depreciation. All capital depreciation in the model occurs via straight line method over an eight-year period.

Taxable income is adjusted by depletion which is calculated via cost depletion methodology and percentage depletion methodology appropriate to

the operation and varies depending upon the period of operation.

The property is subject to the Nevada Commerce Tax which is modeled at a

rate of 0.051% of the gross proceeds.

The Project is also subject to the Nevada Net Proceeds of Minerals Tax which is a sliding scale tax

applied to the net proceeds of the operation. The sliding scale is calculated from a ratio of net proceeds to gross proceeds. The applicable tax rate is presented in Table 19-2.

Table 19-2: Net Proceeds of Mineral Tax Sliding Scale

NP/GP Lower Limit

(%)

NP/GP Upper Limit

(%)

Tax Rate

(%)

10

—

2.0

18

10

2.5

26

18

3.0

34

26

3.5

42

34

4.0

50

42

4.5

50

4.5

Source: State of Nevada

Property tax has been included in the model and is captured in the G&A cost.

i-80 Gold has indicated that approximately US$46.7 million in existing loss carryforwards will

apply to the property. This amount is modeled as a loss carryforward opening balance.

The Project is subject to a number of royalties as

outlined in previous sections. These royalties vary in rate and area of influence. For modeling purposes a combined royalty rate of 7.5% is applied to all material through 2030. All material after that date is subject to a royalty of 9.0%. Royalty

costs total US$104 million over the life of the Project as modeled for this analysis.

Working Capital

The assumptions used for working capital in this analysis are as follows:

•

Accounts Receivable (A/R): 1 month delay

•

Accounts Payable (A/P): 1 month delay

•

Zero opening balance for A/R and A/P

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Technical Factors

Mining Profile

The

modeled mining profile was developed by SRK. The details of mining profile are presented previously in this report. No modifications were made to the profile for use in the economic model. The modeled profile is presented in Figure 19-1 and Figure 19-2.

Source: SRK, 2026

Figure 19-1: Granite Creek Mining Profile (imperial)

Source: SRK, 2026

Figure 19-2: Granite Creek Mining Profile (metric)

Table 19-3 is a summary of the estimated mine production over a

9-year mine life for the operation. Ore mined refers to Proven and Probable Mineral Reserves.

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Table 19-3: Life-of-Mine Production Summary

LOM Mining

Unit

Value

Autoclave Ore Mined

tons

2,358,273

Oxide Ore Mined

tons

36,171

Total Ore Mined

tons

2,394,444

Autoclave Ore Mined

tonnes

2,139,392

Oxide Ore Mined

tonnes

32,814

Total Ore Mined

tonnes

2,172,206

Average Mined Autoclave Ore Grade

oz/ton

0.23

Average Mined Oxide Ore Grade

oz/ton

0.29

Average Mined Ore Grade

oz/ton

0.23

Average Mined Autoclave Ore Grade

g/t

7.85

Average Mined Oxide Ore Grade

g/t

10.02

Average Mined Ore Grade

g/t

7.88

Contained Autoclave Gold

ounces

539,738

Contained Oxide Gold

ounces

10,573

Contained Ore Gold

ounces

550,311

Stockpile

tons

28,930

Stockpile

tonnes

26,245

Stockpile Grade

oz/ton

0.21

Stockpile Grade

g/t

7.33

Stockpile Contained Gold

ounces

6,181

Source: SRK, 2026

Processing Profile

Material from the Project is processed at either NGM facilities or i-80 gold facilities through an

autoclave circuit or an oxide ore circuit. As i-80 Gold completes their processing facilities, ore will stop being shipped to NGM facilities for processing and will instead be shipped to i-80 Gold’s facilities for processing. The i-80 Gold processing facility is expected to be operational in January 2028. In July 2027, Granite Creek will stop shipping

ore to NGM facilities and ore will instead be stockpiled at thei-80 processing facility. Once the facility comes on-line with an expected 2,500 tons per day capacity the

stockpile will be drawn down and processed alongside the direct feed ore. All ore mined or in stockpiles is processed as outline above. The resulting production profile is presented in Figure 19-3.

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SEC

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Source: SRK, 2026

Figure 19-3: Granite Creek Processing Profile

A summary of the modeled LoM processing profile is presented in Table 19-4.

Table 19-4: Granite Creek Processing Summary

LOM Processing

Unit

Value

Autoclave

Ore Ounces Processed

ounces

545,919

Recovery

%

87.6%

Recovered Ore Ounces

ounces

478,095

Oxide

Ore Ounces Processed

ounces

10,573

Recovery

%

68.9%

Recovered Ore Ounces

ounces

7,284

Total

Total Ounces Processed

ounces

556,492

Recovery

%

87.2%

Total Ounces Recovered

ounces

485,379

Source: SRK, 2026

Operating Costs

Operating

costs are modeled in US dollars and can be categorized as mining, surface, power, processing and G&A costs. No contingency amounts have been added to the operating costs within the model. A summary of the operating costs over the life of the

operation is presented in Figure 19-4. Operating costs total US$900.7 million over life of mine at a rate of US$371.67/ton processed (US$409.69/t processed).

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Source: SRK, 2026

Figure 19-4: LoM Operating Cost Summary

The contributions of the different operating cost segments over the life of the operation are presented in Figure 19-5.

Source: SRK, 2026

Figure 19-5: LoM Operating Cost Contributions

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Mining

The mining cost profile was developed external to the model and incorporated into the model as a fixed cost. The result of this approach is

presented in Table 19-5.

Table 19-5: Granite Creek

Mining Cost Summary

LoM Mining Costs

Unit

Value

Mining Cost (w/o power)

US$M

395.98

US$

/ton mined

165.38

US$

/t mined

182.30

Source: SRK, 2026

US$M: US$ million

Processing

Processing costs were developed external to the model and incorporated into the model as variable costs. Ore transportation costs were included

in the processing costs for modelling purposes. The result of this approach is presented in Table 19-6.

Table 19-6: Granite Creek Processing Costs

LoM Processing Costs

Unit

Value

Processing Cost

US$

M

304.21

US$

/ton processed

125.53

US$

/t processed

138.37

Source: SRK, 2026

US$M: US$ million

Surface

Costs

Surface Costs were developed external to the model and incorporated on a fixed basis. The result of this approach is

presented in Table 19-7.

Table 19-7: Granite Creek

Surface Costs

LoM Surface Costs

Unit

Value

Surface Cost (w/o power)

US$

M

117.85

US$

/ton mined

49.22

US$

/t mined

54.25

Source: SRK, 2026

US$M: US$ million

Power

Costs

Power consumption for surface and underground was calculated external to the model and incorporated into the model on a

fixed basis. A power cost of US$0.08/kwH was applied in the model. The result of this approach is presented in Table 19-8.

Table 19-8: Granite Creek Power Costs

LoM Power Costs

Unit

Value

Power Cost

US$

M

32.77

US$

/ton mined

13.69

US$

/t mined

15.09

Source: SRK, 2026

US$M: US$ million

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G&A

G&A was calculated external to the model and incorporated into the model on a fixed basis. The result of this approach is presented in Table 19-9.

Table 19-9: Granite Creek G&A Costs

LoM G&A Costs

Unit

Value

G&A Cost

US$

M

49.87

US$

/ton processed

20.58

US$

/t processed

22.68

Source: SRK, 2026

US$M: US$ million

Selling

Costs

Selling costs consist of the transport costs associated with moving the operation’s product to the selling point

refining charges incurred. Doré refining and freight costs are modeled as follows:

•

99.75% payable Au

•

US$1.77/troy oz Au refining cost

•

US$3.69/troy ounce transportation cost

Capital Costs

Sustaining capital estimates were developed external to the model as outlined in the previous sections. No additional contingency has been

included in the model. General closure costs are modeled as sustaining capital and are captured as a one-time payment the year following cessation of operations.

Total sustaining capital is presented in Table 19-10.

Table 19-10: Modeled Sustaining Capital

Capital Costs

Unit

Value

Mine Capex

US$

M

60.74

Autoclave Capex Allocation

US$

M

49.33

Surface Capitalized Costs

US$

M

22.19

Closure

US$

M

12.66

Total

US$

M

144.92

Source: SRK, 2026

The modeled capital profile is presented in Figure 19-6.

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SEC

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Source: SRK, 2026

Figure 19-6: Granite Creek Sustaining Capital Profile

19.2

Results

The economic analysis metrics are prepared on annual after-tax basis in US dollars. The results of the

analysis are presented in Table 19-11. The results indicate that at a gold price of US$2,750 the after tax NPV @ 5% is US$118 million. As the Project is in operation, IRR and payback timeframe are not

relevant metrics.

Table 19-11: Indicative Economic Results

LOM Cash Flow

Unit

Value

Total Revenue

US$

M

1,328.81

Royalty

US$

M

(104.47

)

Total Opex

US$

M

(900.68

)

Operating Margin

US$

M

323.65

Operating Margin Ratio

%

24

%

Taxes Paid

US$

M

(25.35

)

Free Cash Flow

US$

M

153.38

Before Tax

Free Cash Flow

US$

M

178.73

NPV @ 5%

US$

M

137.66

NPV @ 7.5%

US$

M

120.98

NPV @ 10%

US$

M

106.39

After Tax

Free Cash Flow

US$

M

153.38

NPV @ 5%

US$

M

117.55

NPV @ 7.5%

US$

M

102.97

NPV @ 10%

US$

M

90.19

Source: SRK, 2026

US$M: US$ million

The Project

does not generate positive free cashflow in 2026 and 2027 largely due to the higher processing costs incurred as a result of third-party processing of ore and a six-month stockpiling period in 2027. Cashflow

is forecast to be negative at the end of the operational life as the operation winds down, and closure costs are incurred.

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SEC

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This estimated cash flow is inherently forward-looking and dependent upon numerous

assumptions and forecasts, such as macroeconomic conditions, mine plans and operating strategy, that are subject to change.

The life of

mine AISC cost is estimated at US$2,375/oz Au.

The economic result and back-up chart information

for the charts withing this section are presented on an annual basis in Table 19-12 and Figure 19-7.

September 2026

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SEC

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Table 19-12: Economic Results – Tabular

Data

Cashflow Model Data Summary

Granite Creek

Period Start

1-May-26

1-Jan-27

1-Jan-28

1-Jan-29

1-Jan-30

1-Jan-31

1-Jan-32

1-Jan-33

1-Jan-34

Period End

31-Dec-26

31-Dec-27

31-Dec-28

31-Dec-29

31-Dec-30

31-Dec-31

31-Dec-32

31-Dec-33

31-Dec-34

Calendar Year

Total

2026

2027

2028

2029

2030

2031

2032

2033

2034

Cashflow

Income

Net Revenue

USD

1,328,812,296

97,853,880

64,222,781

275,823,877

198,447,769

181,837,035

204,880,336

167,489,136

104,443,793

33,813,689

Total

USD

1,328,812,296

97,853,880

64,222,781

275,823,877

198,447,769

181,837,035

204,880,336

167,489,136

104,443,793

33,813,689

Operational Expenditure

Fixed

USD

(563,700,553

)

(43,427,268

)

(59,957,658

)

(73,417,431

)

(78,752,715

)

(77,217,002

)

(79,732,922

)

(69,599,378

)

(53,205,072

)

(28,391,108

)

Variable

USD

(336,984,135

)

(42,858,496

)

(33,990,592

)

(56,739,659

)

(44,889,267

)

(39,861,250

)

(45,734,756

)

(37,837,640

)

(25,792,531

)

(9,279,945

)

Total

USD

(900,684,688

)

(86,285,764

)

(93,948,250

)

(130,157,090

)

(123,641,981

)

(117,078,251

)

(125,467,678

)

(107,437,017

)

(78,997,603

)

(37,671,052

)

Royalty

USD

(104,473,217

)

(7,251,737

)

(4,641,556

)

(20,029,635

)

(14,427,325

)

(13,253,919

)

(17,983,111

)

(14,697,023

)

(9,203,789

)

(2,985,125

)

Working Capital Adjustment

USD

0

(5,311,948

)

10,960,294

(7,793,197

)

(1,214,654

)

(6,589,214

)

(1,682,841

)

9,844,588

(1,211,660

)

2,998,633

Capital Costs

Mine Capex

USD

(60,739,069

)

(16,606,202

)

(11,785,494

)

(10,306,973

)

(6,693,201

)

(4,364,800

)

(4,364,800

)

(4,364,800

)

(2,252,800

)

—

Autoclave Capex Allocation

USD

(49,326,013

)

—

(2,952,940

)

(8,577,679

)

(8,568,711

)

(7,505,803

)

(8,743,674

)

(7,106,626

)

(4,568,008

)

(1,302,572

)

Surface Capitalized Costs

USD

(22,194,094

)

(15,617,998

)

(3,370,385

)

(1,708,365

)

(612,736

)

(524,157

)

(345,033

)

(15,419

)

—

—

Closure

USD

(12,660,000

)

—

—

—

—

—

—

—

—

(12,660,000

)

Total

USD

(144,919,176

)

(32,224,199

)

(18,108,819

)

(20,593,017

)

(15,874,648

)

(12,394,760

)

(13,453,508

)

(11,486,845

)

(6,820,808

)

(13,962,572

)

Cashflow Before Funding

USD

178,730,749

(33,219,769

)

(41,515,550

)

97,250,781

43,288,839

32,520,495

46,292,558

43,710,407

8,209,683

(17,806,693

)

Cashflow Before Tax

USD

178,730,749

(33,219,769

)

(41,515,550

)

97,250,781

43,288,839

32,520,495

46,292,558

43,710,407

8,209,683

(17,806,693

)

Tax Paid

USD

(25,352,526

)

(169,482

)

(167,309

)

(5,099,687

)

(3,232,502

)

(4,586,767

)

(6,543,604

)

(5,279,478

)

(184,483

)

(89,216

)

Net Cashflow

USD

153,378,223

(33,389,251

)

(41,682,859

)

92,151,094

40,056,337

27,933,729

39,748,954

38,430,929

8,025,200

(17,895,909

)

Ounces Payable

ounces

484,166

35,654

23,400

100,499

72,306

66,254

74,650

61,026

38,055

12,320

Gold Price

USD/

oz

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

Cash Cost

Royalty and Operating Cost

USD

1,005,157,905

93,537,501

98,589,806

150,186,725

138,069,306

130,332,171

143,450,789

122,134,040

88,201,392

40,656,177

Cash Cost

USD

1,005,157,905

93,537,501

98,589,806

150,186,725

138,069,306

130,332,171

143,450,789

122,134,040

88,201,392

40,656,177

Cash Cost

USD/

oz

2,076

2,623

4,213

1,494

1,910

1,967

1,922

2,001

2,318

3,300

All-in Sustaining Cost

Cash Cost

USD

1,005,157,905

93,537,501

98,589,806

150,186,725

138,069,306

130,332,171

143,450,789

122,134,040

88,201,392

40,656,177

Sustaining Capital

USD

144,919,176

32,224,199

18,108,819

20,593,017

15,874,648

12,394,760

13,453,508

11,486,845

6,820,808

13,962,572

All-in Sustaining Cost

USD

1,150,081,546

125,761,700

116,698,625

170,779,899

153,944,276

142,727,326

156,904,937

133,623,317

95,022,450

54,619,015

All-in Sustaining Cost

USD/

oz

2,375

3,527

4,987

1,699

2,129

2,154

2,102

2,190

2,497

4,433

Project Cashflow (unfinanced)

Revenue

USD

1,328,812,296

97,853,880

64,222,781

275,823,877

198,447,769

181,837,035

204,880,336

167,489,136

104,443,793

33,813,689

Operating Cost

USD

(900,684,688

)

(86,285,764

)

(93,948,250

)

(130,157,090

)

(123,641,981

)

(117,078,251

)

(125,467,678

)

(107,437,017

)

(78,997,603

)

(37,671,052

)

Working Capital Adjustment

USD

0

(5,311,948

)

10,960,294

(7,793,197

)

(1,214,654

)

(6,589,214

)

(1,682,841

)

9,844,588

(1,211,660

)

2,998,633

Royalty

USD

(104,473,217

)

(7,251,737

)

(4,641,556

)

(20,029,635

)

(14,427,325

)

(13,253,919

)

(17,983,111

)

(14,697,023

)

(9,203,789

)

(2,985,125

)

Sustaining Capital

USD

(144,919,176

)

(32,224,199

)

(18,108,819

)

(20,593,017

)

(15,874,648

)

(12,394,760

)

(13,453,508

)

(11,486,845

)

(6,820,808

)

(13,962,572

)

Tax Paid

USD

(25,352,526

)

(169,482

)

(167,309

)

(5,099,687

)

(3,232,502

)

(4,586,767

)

(6,543,604

)

(5,279,478

)

(184,483

)

(89,216

)

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SEC

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Project Net Cashflow

USD

153,378,223

(33,389,251

)

(41,682,859

)

92,151,094

40,056,337

27,933,729

39,748,954

38,430,929

8,025,200

(17,895,909

)

Cumulative Net Cashflow

USD

(33,389,251

)

(75,072,111

)

17,078,984

57,135,320

85,069,049

124,818,003

163,248,932

171,274,132

153,378,223

Operating Cost (LOM)

Mining Cost

USD

395,983,226

33,170,017

40,658,268

52,583,668

56,857,053

55,329,645

57,780,356

48,306,983

35,362,501

15,934,735

Surface Costs

USD

117,851,064

7,026,057

13,480,558

14,925,111

15,626,448

15,719,939

15,978,152

15,154,514

12,020,119

7,920,165

G&A

USD

49,866,263

3,231,194

5,818,831

5,908,652

6,269,214

6,167,417

5,974,413

6,137,881

5,822,452

4,536,208

Power Costs

USD

32,773,646

1,836,886

3,328,773

3,711,340

4,126,599

4,174,254

4,179,053

4,154,765

4,152,659

3,109,316

Processing Costs

USD

304,210,489

41,021,610

30,661,819

53,028,319

40,762,668

35,686,995

41,555,702

33,682,875

21,639,872

6,170,629

Total Opex

USD

900,684,688

86,285,764

93,948,250

130,157,090

123,641,981

117,078,251

125,467,678

107,437,017

78,997,603

37,671,052

Mining Profile

Autoclave Ore Mined

tons

2,358,273

178,372

211,605

364,078

363,697

318,582

371,124

301,639

193,888

55,287

Oxide Ore Mined

tons

36,171

8,531

12,885

5,147

3,623

2,764

2,862

359

—

—

Autoclave Ore Mined

tonnes

2,139,392

161,816

191,965

330,286

329,941

289,013

336,678

273,643

175,893

50,156

Oxide Ore Mined

tonnes

32,814

7,739

11,689

4,669

3,287

2,508

2,596

326

—

—

Autoclave Ore Grade

oz/ton

0.23

0.22

0.22

0.22

0.22

0.23

0.23

0.23

0.24

0.27

Oxide Ore Grade

oz/ton

0.29

0.34

0.24

0.35

0.26

0.31

0.28

0.09

—

—

Autoclave Ore Grade

g/t

7.85

7.52

6.11

6.31

6.35

6.63

6.48

6.62

6.66

7.57

Oxide Ore Grade

g/t

10.02

9.72

6.89

9.97

7.47

8.86

7.99

2.58

—

—

Autoclave Ore Ounces Mined

ounces

539,738

39,139

45,825

81,471

81,810

74,850

85,274

70,767

45,768

14,835

Oxide Ore Ounces Mined

ounces

10,573

2,939

3,145

1,818

959

868

811

33

—

—

Production Profile

Autoclave Gold Recovered

ounces

478,095

34,352

22,138

98,172

71,766

65,778

74,231

61,158

38,150

12,351

Oxide Gold Recovered

ounces

7,284

1,392

1,321

2,579

721

643

606

22

—

—

Gold Price

USD/

oz

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

2,750

Cash Cost

USD/

oz

2,076

2,623

4,213

1,494

1,910

1,967

1,922

2,001

2,318

3,300

AISC

USD/

oz

2,375

3,527

4,987

1,699

2,129

2,154

2,102

2,190

2,497

4,433

Capital Cost

Mine Capex

USD

60,739,069

16,606,202

11,785,494

10,306,973

6,693,201

4,364,800

4,364,800

4,364,800

2,252,800

—

Autoclave Capex Allocation

USD

49,326,013

—

2,952,940

8,577,679

8,568,711

7,505,803

8,743,674

7,106,626

4,568,008

1,302,572

Surface Capitalized Costs

USD

22,194,094

15,617,998

3,370,385

1,708,365

612,736

524,157

345,033

15,419

—

—

Closure

USD

12,660,000

—

—

—

—

—

—

—

—

12,660,000

Total

USD

144,919,176

32,224,199

18,108,819

20,593,017

15,874,648

12,394,760

13,453,508

11,486,845

6,820,808

13,962,572

Source: SRK, 2026

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

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391

Source: SRK, 2026

Figure 19-7: Annual Cash Flow Summary

19.3

Sensitivity Analysis

A sensitivity analysis was performed to determine the relative sensitivity of the Project’s NPV to a number of key parameters (Figure 19-8). This is accomplished by flexing each parameter upwards and downwards by 10%. Within the constraints of this analysis, the Project appears to be most sensitive to metal price recovery assumptions within the

processing plant and mined grades.

SRK cautions that this sensitivity analysis is for information only and notes that these parameters

were flexed in isolation within the model and are assumed to be uncorrelated with one another which may not be reflective of reality. Additionally, the amount of flex in the selected parameters may violate physical or environmental constraints

present at the operation.

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

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Source: SRK, 2026

Figure 19-8: Sensitivity Analysis

September 2026

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Technical Report Summary – Granite Creek

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393

20

Adjacent Properties

Most of the mineral rights surrounding Granite Creek are owned or controlled by Nevada Gold Mines. The only reported mining activities adjacent

to Granite Creek were conducted by Nevada Gold Mines at the Turquoise Ridge Complex (TRC) where approximately 28 Moz Au have been produced since 1938 to 2023 (Figure 20-1). The TRC is in Humboldt County,

approximately 64 km (40 mi) northeast of Winnemucca and approximately 10 km (6 mi) north of the Granite Creek Mine. Historic and current production at TRC includes open-pit and underground mining.

Source: Barrick, 2024

Figure 20-1: Summary of Gold Production at Turquoise Ridge (as reported by Barrick 2024)

Deposits that compose the TRC are Carlin-type, structurally and stratigraphically controlled and sediment-hosted. They contain disseminated micrometer-sized gold occurring on arsenic-rich pyrite rims, primarily within decalcified, carbonaceous rocks. Preferred host lithologies for gold mineralization are the Comus Formation, followed by the Valmy and

Etchart Formations. Sub-microscopic gold mineralization is associated with arsenian pyrite, quartz, calcite, realgar, and orpiment. Gold mineralization is likely Eocene in age, and it is overprinted in some

areas by a late stage of realgar, orpiment, and calcite. Gold-bearing zones can be located close to granodiorite and dacite dikes and beneath basaltic sills, evidencing the importance of rheologic contacts to mineralization (Barrick., 2024). The

most recent public disclosure of the Mineral Resources and Mineral Reserves for the Project are reported in the technical report entitled “NI 43-101 Technical Report on the Turquoise Ridge Complex

Humbolt County, Nevada, USA”, dated March 15, 2024, with an effective date of December 31, 2023. A summary of the Mineral Resource and Mineral Reserve statements are shown in Figure 20-2 and

Figure 20-3. The QP has been unable to verify the information on production,

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mineral resources and mineral reserves included for the adjacent properties. This type of adjacent property information is not necessarily indicative of the mineralization at Granite Creek. In

addition, the QP is not aware of any declared mineral resource that might have an impact on Granite Creek’s Mineral Resources, Mineral Reserves or mining operations.

Source: Barrick, 2024

Figure 20-2: Barrick Mineral Resource Statement for Turquoise Ridge (effective date

December 31, 2023) – reported on 100% basis

Source: Barrick, 2024

Figure 20-3: Barrick Mineral Reserve Statement for Turquoise Ridge (effective date December 31,

2023) – reported on 100% basis

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Other Relevant Data and Information

There is no other relevant data or information necessary to make this Technical Report Summary understandable and not misleading beyond what

has already been disclosed in the preceding items of this report.

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Interpretation and Conclusions

Drilling

A total

of 2,083 drillholes for 291,312 m (955,748 ft) have been completed at the Property since 1970 by successive operators, including PMC and its predecessors, Homestake, Barrick, and Atna. Since acquiring the Property in 2021, i-80 Gold has completed a further 1,083 drillholes for 123,804 m (406,181 ft) within the current property boundary, comprising 438 core holes for 93,552 m (306,929 ft) (including RC

pre-collars with core tails) and 645 RC holes for 30,252 m (99,252 ft).

The i-80 drilling has completed definition and infill of the underground Ogee, Otto, and SPZ mineralization, and has expanded substantially year on year: approximately 69 holes in 2021, 154 in 2022, 278 in 2023, 185 in

2024, and 379 in 2025, the latter marking the introduction of underground production drilling alongside continuing surface drilling. Drilling continued into Q1 2026 as the tail of the 2025 deep SPZ infill campaign together with the start of a new

exploration program targeting the Ogee, Otto, Adam Peak, and Range Front fault systems. All holes supporting the current estimate were drilled between April 2021 and January 2026, with complete assay results returned by February 17, 2026, the

database cut-off date for this study.

Surface holes are generally

pre-collared using RC to the water table and completed with HQ-diameter core; underground holes have been drilled entirely as HQ core. Two RC water wells and one RC

piezometer were also drilled and sampled. Short-term ore-control RC drilling completed with a Cubex rig is used by mine geologists for grade-control decisions only and is not merged into the resource database.

A separate geotechnical drilling program of approximately 6,100 m (20,000 ft) across 40 holes has been completed to support underground

mine design, incorporating Rock Mass Rating logging, acoustic televiewer surveys in 11 holes, and dedicated geotechnical holes for the VR1 and VR2 vent raises.

Core recovery across the i-80 programs is reported at 99% or greater, with recorded core loss limited

to short intervals where voids are present in the stratigraphy. Sample intervals are assigned by the project geologist and are not permitted to cross geological, alteration, or oxidation boundaries, and do not exceed 3.05 m (10 ft); average core

sample length from 2023 through Q1 2026 has been 1.25 to 1.46 m (4.1 to 1.5 m (4.8 ft)), with nominal 0.61 to 1.52 m (2 to 2 m (5 ft)) intervals through mineralized zones. Downhole surveys are completed on all holes at 15.25 m (50 ft) intervals by

an independent contractor using north-seeking gyroscopic tools, and all geological and geotechnical logging is recorded directly into the acQuire database management system. Sample security and chain of custody are maintained from the drill site

through to the laboratory, with no chain-of-custody breaches reported over the period.

No QA/QC data is available for the pre-2005 portion of the historical dataset; however, the volume of

subsequent drilling completed under the modern QA/QC framework is sufficient to support the MRE without exclusive reliance on the unverified historical data, and therefore it is the QP’s opinion that it is not material.

Drilling programs completed at the Property between 2005 and 2015 incorporated QA/QC monitoring through the insertion of CRMs, blanks, and

duplicates. These programs were reviewed by the prior QPs, most recently by GRE (2025), which found no material errors and concluded the historical assay database to be acceptable for Mineral Resource estimation, subject to observations on insertion

rates and duplicate precision typical of programs of that era. The QP has also reviewed this data and agrees with these conclusions.

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Since 2021, and formalized from the 2023 program onward,

i-80 has operated a four-stream QA/QC program comprising CRMs, coarse blanks, field duplicates, and preparation duplicates, supplemented by independent third-party check assays. The program targets an overall

control-sample insertion rate of 20% of submitted samples. The 2021 and 2022 programs operated under a transitional framework at aggregate rates of approximately 13% to 21%. Sample preparation and analysis have been performed throughout by

independent ISO/IEC 17025-accredited commercial laboratories, being ALS through 2024 and MSA Labs as principal laboratory from 2025.

CRM

performance, assessed at two and three standard deviations of the certified mean, showed no consistent bias across the principal standards used. i-80 2,353 blanks inserted over the i-80 era, 89.8% returned values below the 5x lower-limit-of-detection design threshold; applying a materiality floor at 10x detection

limit, consistent with the QP’s assessment of what constitutes material contamination, raises the effective pass rate to >92.5% with the majority of the failures concentrated in batches assayed at the Lone Tree in-house laboratory rather than distributed across the independent commercial laboratories.

Duplicate

performance was assessed against tolerance envelopes derived using the Thompson and Howarth (1978) framework, with preparation duplicates returning an R-squared of 0.99 across the program and 0.9997 in Q1

2026, which indicates a strong correlation between laboratories. There is, a gap in Third-party check assays programs between 2023 and early 2024, which the QP discussed with i-80, and it was noted as a

datatype labelling issue in the database versus a lack of submissions. This is recommended to be addressed in the database. The program recommenced, and the data labelling fixed from 2025; across the full population of 543 check samples from 49

holes, the mean bias is -6.5% and the median -3.5% relative to the primary assay, with 75.3% of pairs agreeing within plus or minus 10%, which the QP on review has

deemed acceptable.

The transition of the primary gold method from conventional fire assay to Chrysos Photon Assay in 2025 was

independently validated by the QP in a three-way comparison against conventional Fire Assay and Screen Fire Assay, returning coefficients of determination in excess of 0.994 against both reference methods with

a small positive bias in the Photon Assay results that is mechanistically explained by, and consistent with, the substantially larger effective sample mass analyzed. The QP separately compared the assay database against original laboratory

certificates across 138 certificate files and 14,127 sample rows, confirming a 99.90% match on gold values for the 7,188 directly comparable samples; the small number of residual items were resolved as laboratory

re-run reconciliations or database ingest gaps of negligible grade impact and have been provided to i-80 for correction.

It is the QP’s opinion that the drilling and sampling completed at Granite Creek are of a type, quantity, quality, and spatial

distribution appropriate to the style of mineralization, and that the drilling, sampling, sample preparation, analytical, and security procedures applied across both the historical and i-80 programs are

adequate and reasonable for use in the Mineral Resource estimation process. The QP further considers the resulting database adequate to support the Mineral Reserve estimate and the technical disclosures presented in this Report, consistent with

Society for Mining, Metallurgy and Exploration (2017) and the disclosure requirements of S-K 1300.

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Environmental and Permitting

The environmental and permitting information reviewed indicates that Granite Creek is well characterized and currently operating under an

established, in-good-standing permit base, supporting reasonable confidence that environmental and permitting factors will not materially disrupt continued operations or the reliability of the underlying

resource and reserve estimates. The principal uncertainty affecting projected economic outcomes is the 2025 discovery of petroleum-contaminated soil in the CX Pit backfill: because the extent of contamination is not precisely known, remediation

could require re-handling of up to 1.17 million tons of waste rock at a cost that is not yet quantified, and this remains an open-ended exposure until Osgood and NDEP-BMRR finalize a management plan. A

related, lower-order uncertainty is the site’s reclamation and closure cost estimate (US$8.64 million), which reflects a government cost model rather than Osgood’s actual self-performed costs, will be revised at the next three-year bond

review (2028 to 2029), and may shift further as site conditions evolve ahead of the anticipated 2032 closure. Ongoing management of elevated arsenic and antimony in site water, and the designated status of the Kelly Creek hydrographic basin under

Nevada water law, add continuing operating costs and a modest constraint on securing incremental water rights, respectively, but neither is expected to limit the project’s ability to operate under its existing permits. None of these risks and

uncertainties are considered reasonably likely to threaten the project’s overall economic viability, as they are cost and scheduling matters addressable through established treatment technologies and routine regulatory processes rather than

threats to continued permitting, access, or operation of the Project.

Metallurgy

Granite Creek underground samples were refractory with baseline CIL gold recoveries ranging from 9% to 46%, averaging 31%;

Shake flask tests with gold cyanide spikes were used to determine preg robbing index. The average preg-robbing index was 17.9%, ranging from

4.4% to 54.1%.;

Bench top autoclave batch pressure oxidation tests were completed on all samples with 2 sets of acid conditions and four

sets of alkaline conditions. Acid conditions resulted in higher sulfur oxidations and higher gold recoveries;

Three continuous pressure

oxidation runs were completed with two acid and one alkaline sets of conditions based on the batch results. The continuous results followed the results of the batch tests with the acid conditions producing higher sulfur oxidations and gold

recoveries. High acid additions required for high carbonate samples resulted in lower sulfide oxidation and pressure oxidation viscosity issues;

Overall gold recoveries increased with increasing sulfur oxidation;

Cyanide destruction tests on CIL tailings using the SO2/air process reduced weak acid

dissociable cyanide concentrations to below 5 mg/L using established reagent addition rates and retention time;

Thickening and filtration

tests on CIL tailings showed high thickening unit area requirements and low filtration rates.

Arsenic concentrations in the samples

averaged 0.29%, largely occurring as arsenian pyrite with only trace amounts of arsenopyrite.

Sulfide minerals were predominantly pyrite

with some marcasite.

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Mercury concentrations ranged from 31 ppm to 138 ppm, averaging 81 ppm. These concentrations

will require mercury capture and abatement equipment in the process flowsheet.

Sample Representativity

Samples were selected to provide spatial and grade representation of three zones in the Granite Creek underground deposit as seen in Figure 10-16 and Figure 10-17. The samples’ metallurgical response is likely to represent the zone’s general behavior, but additional sampling of each zone to confirm the

metallurgical response will reduce uncertainty. The 33 samples tested in the FLS and SGS programs provide the required coverage for this study. Grade distributions of the samples as outlined in section 13.2.2, reflect grade distributions with the

resource.

Test Work on Underground Samples

The two recent testwork programs and one historical program confirm that refractory Granite Creek samples are amenable to pressure oxidation

followed by CIL to provide acceptable gold recoveries. While the samples respond to both alkaline and acid pressure oxidation, superior recoveries were achieved with acid pressure oxidation. The results confirm that Granite Creek production can be

processed through the Lone Tree acid pressure oxidation conditions once restarted. The testwork also confirmed that Granite Creek production can be processed through acid pressure oxidation at third party processors.

Mineral Resources

The Mineral Resource for the Granite Creek deposit comprises both open pit and underground components. These have been estimated separately by

GRE (open pit) and SRK (underground) using block modeling methods appropriate to each mining environment.

The open pit Mineral Resource

was estimated by GRE and remains unchanged in terms of input technical information, drilling and sampling from the previously disclosed estimate; however, GRE has updated the assessment of reasonable prospects for economic extraction (RPEE) to

reflect the updated resource gold price. The Mineral Resources for the Granite Creek open pit mine project were estimated in conformity with the Society for Mining, Metallurgy and Exploration (2017) guidelines and are reported in accordance

with the S-K 1300. This mineral resource estimate includes Inferred Mineral Resources. Inferred resources are defined as resources that are “…reasonably expected that the majority of the Inferred

Mineral Resources could be upgraded to indicated.” by additional drilling. There is also no certainty that the Inferred Mineral Resources will be converted to the Measured or Indicated categories through further drilling or into Mineral

Reserves, once economic considerations are applied. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the mineral resource will be converted into Mineral

Reserves. The project presently has no Mineral Reserves. Whittle Pit optimization was applied to the open pit mineral resource estimate to assess the reasonable prospects for economic extraction for the resource.

In the opinion of the QP, the Mineral Resource model presented in this report is representative of the informing data, which is of sufficient

quality and quantity to support the Mineral Resource estimate to the classifications applied.

The underground Mineral Resource has been

estimated by SRK Consulting. SRK completed the MRE process using the updated exploration database which targeted the underground operation completed by i-80.

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The open pit Mineral Resource was estimated by GRE, and remains unchanged in terms of input

technical information, drilling and sampling from the previously disclosed estimate; however, GRE has updated the assessment of reasonable prospects for economic extraction (RPEE) to reflect the updated resource gold price. The effective date of

both the open pit and underground MREs is March 31, 2026.

For the underground Mineral Resource, which is the focus of current mining

and near-term development, mineralization is modeled within three principal structurally controlled domains: Ogee, Otto, and the SPZ, with a fourth smaller domain defined in the SPZhw, located above the CX West Fault. Each of these is subdivided

into high-grade (HG) and low-grade (LG) sub-domains.

Domain boundaries were defined using indicator grade shells at a nominal 0.5 g/t

(0.015 oz/st) Au threshold for the low-grade and 1.5 to 2.5 g/t (0.073 oz/st) Au for the high-grade (refined using an ISO-value of approximately 0.44 to 0.45 probability). The domains have been constrained

geologically by key structural elements including the CX West Fault, Range Front Fault, and the interpreted domain boundaries A, B, and C. These interpretations were cross-checked against underground grade-control data and verified during the

QP’s site inspection of active mine headings.

Gold grades were estimated using capped 1.53 m (5 ft) composites and a combination of

Ordinary Kriging and Inverse Distance Weighting. Search orientations and variogram models were aligned with the principal fault and vein orientations that control each domain. Gold grade continuity supports the use of orientated domain-specific

variogram models for the Ogee, Otto, and SPZ (HG and LG) domains. Whereas sample density in the SPZhw sub-domain was insufficient to support a robust directional variograms and therefore omni-directional variograms were applied. Estimation

parameters were refined using Kriging Neighborhood Analysis (KNA), resulting in a first-pass search using 9–12 composites (maximum 4 composites per hole, reduced to 3 in the final pass, which equates to a minimum of 3 holes in the first pass

and 2 holes in the second and third passes).

Geometallurgical parameters (cyanide-soluble gold, total organic carbon [TOC], preg-robbing

indices [PregRob], sulfide sulfur, and carbonate [CO₃]) were estimated within the same domain framework and using the same orientation concepts as gold. These estimates were supplemented with underground muck-sample data to improve local

coverage, particularly in the upper Ogee and Otto areas where exploration-drilling coverage for these parameters is limited. Oxide and sulfide boundaries have been defined based on visual logging of the drilling which were interpretated using

indicator values for the highest oxidation state, which was then combined with additional oxide domain models provided by i-80, with the results compared to the logging and sulfur grades for confirmation.

Block model validation, including visual checks, statistical comparisons, and swath-plot comparisons against composites and Nearest Neighbor

(NN) estimates, indicates no material local bias existed based on the comparison of the declustered composite mean, nearest neighbor and block estimates. High-grade domain estimates are within approximately 5% of declustered composite means.

Mineral Resources are classified as Measured, Indicated, or Inferred in accordance with the SME Definition Standards (SME, 2017), based on

drilling density, the number of composites and drillholes informing each block estimate, and geostatistical confidence measures (estimation variance and slope of regression).

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Resources are reported on a diluted basis within stope shapes generated MSO (Deswik MSO), at

an average cut-off grade of 3.38 to 9.81 g/t (0.286 oz/st) Au, based on a gold price of US$3,250/oz, variable recoveries (range between 37.2 % to 97.0 % for the autoclave and 30.0 % to 86.5 % for the oxide material) and

reasonable assumptions for mining, processing, and general and administrative costs.

It is the QP’s opinion that the underground and

open pit MREs for Granite Creek are based on adequate exploration and production data, utilize estimation methods and classification criteria appropriate for the style of mineralization, and are reasonable for public disclosure as Mineral Resources

under NI 43-101. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

Mineral

Reserves

Mineral Reserves were classified using the Society for Mining, Metallurgy and Exploration (2017) Standards for

Mineral Resources and Mineral Reserves. Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the appropriate modifying factors, as described herein, to potential mining shapes created during the mine design process. In

the same manner, Measured Mineral Resources were converted to Proven Mineral Reserves.

Mineral Reserves are subject to risks typically

associated with drift and fill operations. These risks could affect the reserves and include, but are not limited to, the following:

•

Differences between the Mineral Reserves metal price assumption and actual metal prices.

•

Changes in the interpretation of the deposit geometry and continuity.

•

Changes to the estimated gold grades for the deposit.

•

Changes to mining, processing, and G&A costs used to determine the

cut-off grade.

•

Changes to metallurgical recovery rates.

•

Changes in the geotechnical assumptions leading to potential impacts to schedule, dilution and/or mining

recovery.

•

Changes to mining method in portions of the deposit.

•

Assumptions related to ongoing access to the site, retention of mineral tenure, obtaining necessary

environmental, mining, and other regulatory permits, and maintaining a social license to operate with relevant stakeholders.

Mining and Infrastructure

Mining

Mine

infrastructure has been completed. Production ramp up has reached approximately 600 ore tons per day. The mining contractor is in place with the full complement of equipment and personnel.

Decline development has accessed 700 vertical ft of mineralization of the Otto and Ogee zones. Development has reached the top of the SPZ

allowing additional active production stopes.

The drill lateral drift over the SPZ has been completed.

Reconciliation of the model to mill indicates process head ounces exceed model by 19%. This appears to be from mining in a larger low-grade halo around high grade core.

Processed grade is lower than the

life-of-mine planned grade due to extensive mining of marginal mineralization below the economic cut-off grade.

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Mine Ventilation

After a comprehensive review of ventilation data and analysis, WSP concludes that the recommendations in this report are adequate to support

pre-feasibility level design. WSP’s experience indicate that these methods and design parameters are in line with best practices at other successful operations in the region. i-80 has been operating the

Granite Creek Mine since 2022 without major complications. In addition, the technical staff and mining workforce draw from extensive experience in the region, further increasing the confidence in successfully mining the South Pacific Ore body.

Outside adjustments made within normal mining practices, WSP does not foresee any major complications that would drastically impact the pre-feasibility estimates proposed in this report.

Infrastructure

Previous

operators of the Granite Creek Project constructed a large portion of the infrastructure required to operate the Granite Creek underground mine. This includes:

•

Connection to the NV Energy grid

•

5 MW substation

•

13.8 kV distribution network

•

Administration building

•

Dry facilities

•

Three of the active dewatering wells

•

Two of the four permitted rapid infiltration basins

Since acquisition of the project, i-80 has added to the existing infrastructure with completion of:

•

Dewatering well GCW5

•

Deepening of well BPW5 to 677 m (2,220 ft)

•

Expansion of the dry facilities to accommodate 176 personnel

•

Installation of a rental Water Treatment Plant to treat 800 gpm of dewatering water for discharge to the ribs

Infrastructure construction in progress includes:

•

Drilling and completion of dewatering wells GCW14 and GCW15

•

Construction of a 3,300 gpm water treatment plant

•

Construction of the two remaining permitted RIBs

•

Expansion of the contact water handling system and tie-in into the water

treatment circuit

After completion of the infrastructure construction currently underway, the electrical power

distribution, dewatering, water treatment, administrative and related support infrastructure will be adequate to support the planned underground mine operations.

Economics

The

Granite Creek operation consists of an underground mine with no onsite processing facility that mines and transports gold ore to offsite processing facilities. The operational life as modeled for this analysis is approximately 9 years. Under the

forward-looking assumptions modeled and documented in this report, the operation is forecast to generate positive cumulative cash flow. However, the Project does not generate positive free cashflow in 2026 and 2027 largely due to the higher

processing costs incurred as a result of third-party processing of ore and a six-month stockpiling period in 2027.

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Cashflow is forecast to be negative at the end of the operational life as the operation winds down, and closure costs are incurred. This estimated cash flow is inherently forward-looking and

dependent upon numerous assumptions and forecasts, such as macroeconomic conditions, mine plans and operating strategy, that are subject to change.

The results demonstrate that at a gold price of US$2,750/oz the after-tax NPV @ 5% is

US$118 million.

The sensitivity analysis performed indicates that the operation’s NPV is most sensitive to metal price,

recovery assumptions and mined grades.

Risks and Uncertainties

Risks

Environmental and Permitting

•

Petroleum-contaminated soil discovered in 2025 within CX pit backfill – location not precisely known, so

remediation could require re-handling of up to 1.17 million tons of waste rock at a cost that isn’t yet quantified.

•

Reclamation / closure cost estimate (US$8.64 million) is based on a government cost model, not Osgood’s

actual self-performed costs, and is due for revision at the next three-year bond review (2028 to 2029) – actual future liability could differ.

•

Ongoing reliance on water treatment for dewatering water discharged to the RIBs could impact operating cost that

continues for the operational life of mine.

•

Kelly Creek hydrographic basin is designated for closer administration by the Nevada State Engineer, which could

make it harder to secure additional water rights if the project needs more in the future.

•

Regulatory reference values can tighten over time (arsenic/antimony standards were lowered in 2006), which is a

precedent for future standards potentially affecting compliance costs.

Recovery Methods

While acid pressure oxidation provides higher gold recoveries and alkaline pressure oxidation, testing of Granite Creek samples showed high

acid additions required for high carbonate samples results in low sulfide oxidation and slurry viscosity problems. Careful blending of Granite Creek production will be required to ensure these problems do not occur at the Lone Tree pressure

oxidation facility during operations.

Lone Tree tailings management will rely on filtered tailings disposal. Testing has shown low

filtration rates for tailings samples. Efficient operation and maintenance of the tailings filters will be required to ensure design operating time is achieved to avoid unplanned downtime.

Opportunities

Mining

•

Upgrade of some or all the Inferred Mineral Resources to higher-confidence classification categories, with

additional drilling and supporting studies, such that this higher confidence material could potentially be converted to mineral reserves.

•

Exploration and step out drilling for the potential to expand the mineral resource inventory, with potential

resource to reserve conversion.

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•

There is potential for a positive impact to the operating cost by considering owner operated operation instead of

the current contractor operated operation.

Environmental and Permitting

•

All permits required for current operations are already held and in good standing, with no permitting actions

pending before the BLM or NDEP – low near-term permitting risk.

•

Legacy tailings storage facilities have already been reclaimed and released from financial assurance, removing a

common source of closure liability.

•

Attenuation study to support site specific discharge requirements to reduce WTP operational costs.

•

Improvement of productivity/production rates as dewatering operations are implemented, and passive inflow to the

underground workings is decreased.

•

Arsenic treatment technology (WTP, ferric sulfate dosing) is already proven at this site, de-risking the main long-term water quality contingency.

•

No onsite processing or new tailings generation under the current plan, which limits the project’s

environmental footprint and closure complexity relative to a milling operation.

•

Nearly three decades of consistent geochemical data (1998 to 2025) show waste rock is reliably non-acid-generating, supporting confidence in the mine plan without unexpected ARD-related costs.

•

Established, positive community relations track record (town halls, regional partnerships) with no unresolved

stakeholder conflicts.

•

An additional 1,364 AFA of water rights is currently ‘Ready for Action’ with NDWR and could be

secured to support future expansion.

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23

Recommendations

23.1

Recommended Work Programs

Metallurgical Testing

•

Conduct additional variability sampling using the most recent mine plan to select samples to provide additional

recovery information to further populate recovery models.

•

Testing should attempt to establish head grade and extraction relationships for use in more detailed resource

modelling;

•

Mineralogy impacts need to be established and geologic domains within each resource need to be determined;

•

Testing of additional samples to provide comminution data to assess hardness variability within the zones and any

potential impacts on throughput in the Lone Tree process plant.

•

The estimated cost for the suggested next phase metallurgical program is US$150,000 based on current market

pricing.

Geology

The QP considers that the drilling and sampling information is sufficiently reliable to interpret the boundaries of the mineralized structures

and domain interpretations, and that the sample grade data are sufficiently reliable to support the MRE.

The 2027–2028 exploration

program targets five zones at Granite Creek: the CX Fault depth/north extension, Rangefront and Adam Peak, the Ogee Zone, SPZ infill and depth extension, and the SPZ/ Rangefront /Mag structural intersection.

The combined drill program comprises 128 planned drillholes totaling 32,614 m (107,000 ft). plus associated CX Fault underground development

and a gravity survey at the SPZ/Rangefront/Mag intersection. The proposed work program has an estimated cost of US$16M based on current conditions and is defined in Table 23-1.

In the opinion of the Qualified Person, the proposed program is technically reasonable and appropriately sequenced, with Rangefront and Ogee

prioritized ahead of active and planned underground development to reduce the risk of mining ahead of drill data, while the CX Fault program advances the known

oxide-to-sulfide transition at depth. If successful, the program has the potential to add additional mineral resources at Rangefront, Adam Peak, Ogee, and the SPZ /

Rangefront / Mag intersection, and to upgrade a portion of the existing SPZ inferred resource to the indicated category. In addition, mineralization drilled at the CX fault target could result in this area being brought into the mine plan in the

near future, allowing flexibility in the mine plan for more active faces and mined tonnage.

Upon completion of the above items, an update

to the geologic model and MRE should be conducted, along with updated metallurgical updated recovery assumptions.

Environmental and

Permitting

•

The ongoing metals attenuation study should be completed, and its findings used to support site specific

discharge requirements to optimize operational WTP costs.

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•

The unpublished LRE post-closure Mag Pit Lake modeling should be finalized to firm up long-term closure water

quality predictions.

•

Given the Kelly Creek basin’s designated status, Osgood should evaluate whether to proactively secure its

1,364 AFA of “Ready for Action” water rights before basin administration tightens further.

•

Ahead of the next three-year bond review (2028 to 2029), SRK recommends Osgood develops an independent,

self-performed closure cost estimate, incorporating updated PCS remediation costs once known, to better anticipate the Project’s actual future reclamation liability rather than relying solely on the government SRCE model used for financial

assurance.

None of these programs are expected to be material in cost relative to the Project, and none are

prerequisites to continued operation under existing permits.

Infrastructure

The following recommendations are offered for improving costs and optimizing operations of the Granite Creek Underground Mine.

•

Perform a cost / benefit analysis for installing power factor correction at the Granite Creek electrical

substation.

•

Install power meters to assist with optimization of costs by at a minimum distinguishing between underground and

surface electrical consumption.

•

Evaluate the installation of a real time monitoring system with adequate flow meters, piezometers, and any other

required instrumentation to optimize mine dewatering efficiency and costs.

•

Investigate the acquisition of a real time highwall monitoring system to provide early warning of a potential

fall of ground, protecting workers, equipment and infrastructure.

Granite Creek Open Pit Technical Study

The Granite Creek open pit Mineral Resources were not included in the Mineral Reserve estimate or economic analysis presented in this

Pre-Feasibility Study. Although the Open Pit was evaluated as part of the 2025 IA, the engineering and supporting technical information available as of the effective date of this Technical Report were not sufficiently advanced to support development

of a PFS- or FS-level mine plan.

Since completion of the

IA, additional trade-off work has been completed to evaluate alternative processing routes, production capacities, mine sequence, capital requirements, and development strategies for the open pit. This work

was intended to support selection of a preferred development pathway and subsequent project advancement rather than establish a feasibility-level mine plan. The recent trade-off study specifically describes

its purpose as evaluating realistic development pathways and relative performance rather than defining a feasibility-level mine plan.

Additional technical work is ongoing to advance the open pit towards a PFS level of definition. This work includes geotechnical

characterization and design of proposed process and mine waste management facilities, evaluation of pit slope stability and development of appropriate geotechnical design criteria, further development of the selected processing configuration,

additional drilling and testing, an updated MRE, mine planning and continued environmental and permitting activities. The current infrastructure program includes PFS-level HLF and TSF design supported by site

characterization and geotechnical investigation.

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It is recommended that the ongoing technical studies be completed and integrated into an

updated open pit mine design, production schedule, capital and operating cost estimates, and economic analysis. Completion of this work will provide the technical basis required to evaluate the potential conversion of applicable open pit Mineral

Resources to Mineral Reserves and potential incorporation of the open pit into a future Granite Creek development plan. The estimated cost of the recommended work program is summarized in Table 23-1.

The estimated costs presented in Table 23-1 are budgetary and are intended to provide an order-of-magnitude estimate of the expenditures required to advance the Granite Creek Open Pit to the next level of study. The PFS engineering and Technical Report estimate is

based on recent comparable study proposals and excludes the separately identified geotechnical investigations, pit slope evaluation, and additional drilling and testing and metallurgical testing.

Ventilation

It is

recommended that i-80 should monitor conditions and implement controls as necessary based on actual results to comply with MSHA regulations and provide a safe working environment. WSP recommends, in accordance

with the cost specified in this report, that all headings are ventilated directly from the intake raises for better control of DPM and silica dust exposure, with the ductwork passing over air doors placed in the crosscuts that connect the decline to

the intake raises. This method provides fresh air directly to working faces for an improved working environment. Additional investment could be made in enclosed cabs for the equipment items which currently have open cabs and in a study for the

optimization of future raise locations. The investment in closed cab equipment would be the responsibility of the contract miner SMD, and it is assumed costs would not be passed through to i-80. WSP recommends

a raise optimization ventilation study and estimates the cost to be approximately US$50,000. Further recommendations related to the geotechnical and construction risks and costs of the raise development are outlined in the Geotechnical portion of

this section.

Geotechnical

Opportunities may exist to optimize the primary ventilation system and reduce capital costs through evaluation of alternate primary vent raise

alignments. Potential alternatives could include shorter raise lengths and alignments located within more competent ground conditions, which may reduce construction complexity, ground improvement requirements, and long-term support needs.

Advancement of an alternative raise alignment would require additional geotechnical investigation, including drilling of a dedicated geotechnical cover hole, televiewer logging, and completion of a geotechnical stability assessment to confirm

feasibility and support design development.

The estimated cost of evaluating a 1000 ft alignment is US$210,000, comprising approximately

US$150,000 for drilling, US$10,000 for televiewer surveying, and US$50,000 for geotechnical engineering assessment and reporting. This investigation would provide additional information to support optimization of the ventilation raise design and

associated project capital costs.

23.2

Recommended Work Program Costs

Table 23-1 summarizes the costs for recommended work programs.

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Table 23-1: Summary of Costs for Open Pit and

Underground Recommended Work Programs

Recommended Work

Estimated Cost

(US$M)

Estimate Basis

Open Pit Focused Work

Open Pit PFS Engineering and Technical Report

2.00

Budgetary

estimate based on

recent

comparable study

proposals

Infrastructure and Facility Geotechnical Studies

0.54

Current proposal

Pit Slope Geotechnical Evaluation

0.25

Budgetary

allowance

Additional Drilling and Testing

3.00

Estimated footage

and unit rates

Additional Metallurgical Testing

0.35

Budgetary

allowance based

on comparable

metallurgical test

programs

SubTotal

6.14

Underground Focused Work Programs

Exploration Drilling and Infill

16.0

Estimated @

US$4 M per

annum

Raise optimization ventilation study

0.05

Ventilation Raise Alignment Geotechnical Evaluation

0.21

Budgetary

estimate

SubTotal

16.26

US$M: US$ million

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24

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25

Reliance on Information Provided by the Registrant

In compliance to the SEC S-K regulation Subpart 1302 (f)(1), SRK may rely on information provided by

the registrant in preparing its findings and conclusions regarding the following aspects of modifying factors:

(i) Macroeconomic trends,

data, and assumptions, and interest rates;

(ii) Marketing information and plans within the control of the registrant;

(iii) Legal matters outside the expertise of the qualified person, such as statutory and regulatory interpretations affecting the mine plan;

(iv) Environmental matters outside the expertise of the qualified person;

(v) Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans; and

(vi) Governmental factors outside the expertise of the qualified person.

The Consultant’s opinion contained herein is based on information provided to the Consultants by

i-80 throughout the course of the investigations. Table 25 1 of this section of the Technical Report Summary will:

(i) Identify the categories of information provided by the registrant;

(ii) Identify the particular portions of the Technical Report Summary that were prepared in reliance on information provided by the registrant

pursuant to Subpart 1302 (f)(1), and the extent of that reliance; and

(iii) Disclose why the qualified person considers it reasonable to

rely upon the registrant for any of the information specified in Subpart 1302 (f)(1).

The Consultant’s opinion contained herein is

based on information provided to the Consultants by i-80 throughout the course of the investigations. SRK has relied upon the work of other consultants in the project areas in support of this Technical Report.

The Consultants used their experience to determine if the information from previous reports was suitable for inclusion in this technical

report and adjusted information that required amending. This report includes technical information, which required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding

and consequently introduce a margin of error. Where these occur, the Consultants do not consider them to be material.

These items have not

been independently reviewed by SRK and SRK did not seek an independent legal opinion of these items.

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Table 25-1: Reliance on Information Provided by

the Registrant

Category

Report Item/ Portion

Portion of

Technical Report

Summary

Disclose why the Qualified

Person

considers it reasonable to rely upon the

registrant

Macroeconomic trends, data, and assumptions, and interest rates

Item 19 – Economic Analysis (External Factors: Pricing, Taxes and Royalties)

Gold price assumptions and discount rate used in the economic model

Corporate pricing decks, tax structuring, and discount rate assumptions are set at the registrant’s corporate finance level and are outside the QP’s mining/geological expertise.

Marketing information and plans within the control of the registrant

Item 16 – Market Studies and Contracts (Contracts and Status)\

Toll Milling Autoclave Contract, Ore Purchase Agreement, and doré/byproduct offtake arrangements

These are registrant-negotiated commercial agreements; the QP relies on the registrant’s own account of contract terms rather than independently verifying them.

Legal matters outside the expertise of the qualified person, such as statutory and regulatory interpretations affecting the mine plan

Item 3 – Property Description (Encumbrances / royalty burdens)

Mineral tenure status and the 5 disclosed royalty interests

Interpretation of property tenure and royalty legal standing is outside the QP’s expertise; consistent with the source report’s own statement that SRK did not seek an independent legal opinion of these items (see

verbatim text above).

Environmental matters outside the expertise of the qualified person

Item 17 – Environmental Studies, Permitting and Social or Community Impact (Environmental Management and Planning; Mine Closure)

Environmental compliance status, closure cost/bonding basis, and the Petroleum Contaminated Soils remediation approach

Environmental regulatory compliance assessment relies on the registrant’s environmental staff and consultants, outside the mining QP’s direct expertise.

Accommodations the registrant commits or plans to provide to local individuals or groups in connection with its mine plans

Item 17 – Environmental Studies, Permitting and Social or Community Impact (Plans, Negotiations, or Agreements with Local Individuals or Groups)

Community agreements and accommodations

Commitments to local stakeholders are registrant-controlled business decisions, not independently verifiable technical facts.

Governmental factors outside the expertise of the qualified person

Item 17 – Environmental Studies, Permitting and Social or Community Impact (Project Permitting)

Item 19 – Economic Analysis

Permitting status, current permitting activities, and performance/reclamation bonding. Application tax and the impact of existing loss carryforward balances.

Interpretation of governmental regulatory requirements and permitting status relies on the registrant’s own compliance representations. Registrant control over the applicability of tax losses

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

417

26

Date and Signature Page

Signed on this 21st Day of September, 2026.

Prepared by

/s/ SRK Consulting (U.S.), Inc,

SRK Consulting (U.S.), Inc,

Prepared by

/s/ Global Resource Engineering

Global Resource Engineering (GRE)

Prepared by

/s/ TR Raponi Consulting Ltd.

TR Raponi Consulting Ltd.

Prepared by

/s/ WSP

WSP

Prepared by

/s/ Practical Mining

Practical Mining

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

418

Prepared by

/s/ i-80 Gold Corp.

i-80 Gold Corp.

Prepared by

/s/ Hatch Ltd.

Hatch Ltd.

All data used as source material plus the text, tables, figures, and attachments of this document have been reviewed and prepared in

accordance with generally accepted industry practices.

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

419

Appendices

September 2026

SRK Consulting (U.S.), Inc.

SEC

Technical Report Summary – Granite Creek

Page

420

Appendix A: Consent Letter

September 2026

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