Overview

Deposit type

• Vein / narrow vein
• Placer
• Intrusion related
• Hydrothermal
• Breccia pipe / Stockwork
• Porphyry

Summary:

The Lemhi Gold Deposit is localized within a major low-angle shear zone (possible thrust fault) and is spatially associated with a high-level porphyritic intrusion. Precious metals mineralization at Lemhi has historically been classified as shear-hosted intrusion related (porphyry-style) mineralization. Both FMC and AGR recognized this deposit type and used a porphyry-related model to guide their exploration program. Key elements of the exploration model were major structures (structural permeability); high-level intrusions (source of heat and fluids); alteration consisting of silicification and sericitization; and gold, copper, and molybdenum geochemical anomalies.

An alternate deposit model has been suggested consisting of a structurally controlled hydrothermal deposit with varying amounts of sulfides in a quartz-carbonate gangue hosted by late-Proterozoic metasediments within the structurally complex Trans-Challis fault system. It has been suggested that gold mineralization was introduced during a tectonically active period and was likely temporally related to intrusive activity associated with the Idaho Batholith. The observed gold mineralization is strongly associated with base metal (Cu and Mo) mineralization and occurs as multiple hydrothermal (epithermal – mesothermal) silica replaced structures resembling multiple flat-lying veins.

The gold deposit on the Lemhi Gold Property shares many similarities with the Beartrack mine, 35 km to the southwest, and the Musgrove deposit, 25 km further southwest. Both Beartrack and Musgrove are quartz stockworks hosted within major shear zones cutting the Apple Creek Formation.

Sericite believed to be a product of hydrothermal alteration yielded an age date of 65.5 +/-2.5 million years ago (Ma). The age date is close to that of the Beartrack deposit (68 Ma) and Napoleon Hill porphyry molybdenum deposit.

Deposit Character and Geometry
The Lemhi Gold Deposit has a footprint of 1000 m east-west by 1100 m north-south and is defined to a depth of 240 m below surface, based on grade x thickness plots. Higher-grade mineralization in the northern part of the deposit has a strong west-northwest alignment. McCarter (1988) describes this high-grade zone as 395 m long by 75 m wide and up to 30 m thick. West-northwest high-angle structures were noted in trenches and road cuts during the fieldwork conducted at the time, and are probably responsible for this trend. A strong northeast trend (035°) and a weaker parallel northeast trend further to the east are also indicated by the grade-thickness contours. Both the west-northwest and northeast high-grade zones are interpreted to be mineralization concentrated at intersections of high-angle structures with the broad low-angle fault zone. In the core of the deposit, the low-grade envelope of mineralization is greater than 200 m thick.

The current developing geologic model for the gold mineralization at the Lemhi Project is of a structurally controlled hydrothermal deposit associated with varying amounts of sulfides in a quartz-carbonate gangue hosted by late-Proterozoic metasediments within the structurally complex Trans-Challis fault system. It is further suggested that gold mineralization was introduced during a tectonically active period and is likely temporally related to intrusive activity associated with the Idaho Batholith. Gold mineralization has a strong association with base metal (Cu and Mo) mineralization and occurs as multiple hydrothermal (epithermal – mesothermal) silica replaced structures resembling multiple flat-lying veins.

Mineralization
Gold deposits in the Dahlonega mining district consist of two types of mineralization: gold-bearing lodes (quartz veins and stockworks), and placer gold deposits derived from weathering of the veins, which were mined in drainages a short distance downstream from the lode deposits. Gold occurs both as lodes (Lemhi Gold Deposit) and placers on the Lemhi project.

Previous interpretation of the mineralization by Cuffney (2011) states that:
“…gold accompanied by minor silver and copper mineralization is spatially and likely genetically related to sub-horizontal dikes/sills that intrude quartzites and phyllites of the Lemhi Group (Gunsight and Apple Creek Formations) in the hangingwall of a low-angle (thrust?) fault. Mineralization occurs as swarms of gold-bearing quartz veins and silicified zones. Quartz veining, silicification, and gold mineralization occur in low-angle zones of sheared/cataclastic phyllite generally dipping gently up to 25° to the southeast. Mineralization more or less surrounds the quartz porphyry intrusions. Thicker and higher-grade gold mineralization occurs in the footwall of the low-angle dike/sill, whereas mineralization above the intrusion is thinner and lower grade. Mineralization is also concentrated along the western terminus of the main intrusive. Minor precious metals mineralization occurs within the intrusions, suggesting that they are pene-contemporaneous with the mineralization.”

Reinterpretation based on the results of the 2012 core drilling program suggested that the deposit is a structurally controlled hydrothermal deposit associated with varying amounts of sulfides in a quartz-carbonate gangue hosted by late-Proterozoic metasediments within the structurally complex Trans-Challis fault system (Brewer, 2019). It is further suggested that gold mineralization was introduced during a tectonically active period and is likely temporally related to intrusive activity associated with the Idaho Batholith. Gold mineralization has a strong association with base metal copper (Cu) and molybdenum (Mo) mineralization and occurs as multiple hydrothermal (epithermal – mesothermal) silica replaced structures resembling multiple flat-lying veins.

Mineralogy
Precious metals mineralization at the Lemhi Project occurs within a gangue of quartz and minor carbonate (magnesite or ferroan dolomite). Bartlett (1986) identified magnesite as the most abundant alteration mineral, occurring as veinlets cutting both quartz veins and wall rocks. Overall gold mineralization has a low sulfide content, normally less than 2%, but pockets of high sulfide concentration have been noted. Pyrite and lesser chalcopyrite and molybdenite are the dominant sulfide species. Bornite, digenite, and traces of galena, sphalerite, pyrargyrite, and arsenopyrite have also been identified. Silver is present in small amounts with silver to gold ratios usually less than 1:1.

Gold is nearly always associated with quartz veining or quartz flooding. McCarter (1985) observed that gold intercepts in drilling correlated with zones of >20% quartz. Sandefur et al. (1994) performed a statistical analysis of mean gold grade vs lithology and alteration codes (9,723 code entries) from the 1985-1994 drilling database. Gold grades were found to be closely related to quartz veining. Two peaks in gold grade were found at quartz concentrations of 35%-50% and at 85%-95%. The latter range averaged 2.65 g/t Au. However, high quartz intervals account for only a small percentage of the volume of the deposit and the bulk of gold mineralization contains 15%-45% quartz. Intervals with < 15% quartz veining contained little or no gold.

A 2023 review of assayed samples from all drilling generations with logged veining percentages agrees with the >20% veining correlation observed by McCarter (1985). A current study of samples >2.65 g/t Au supports the findings of Sandefur et al. (1994); however, the new lower peak is centered closer to 15-20% quartz vein concentration.

Oxidation generally extends 30-50 m below the surface (Bertram, 1996). Gold occurs largely as free gold in the oxide zone. Gold grade below the redox zone correlates with sulfide content, suggesting that gold in primary occurs as auriferous pyrite or within/on copper sulfides.

Mining Methods

• Truck & Shovel / Loader

Summary:

Mine planning is based on conventional drill/blast/load/haul open pit mining methods suited for the project location and local site requirements. The open pit activities are designed for two years of construction followed by twelve years of operations.

Ultimate pit limits are split up into six phases or pushbacks to target higher economic margin material earlier in the mine life. Upper benches will be accessed via internal cut ramps on topography, or via ramps left behind on phased pit walls. In-pit ramps will access material below the pit rim.

Pit designs are configured on 5 m bench heights, with minimum 8 m wide berms placed every four benches, or quadruple benching. Slopes of 25° are applied in the thin overburden layer above the deposit bedrock. Since there has been no geotechnical test work or analysis completed on the bedrock, the applied bench face and inter-ramp angles, 70-75° and 50-55° respectively, are scoping level assumptions based only on the rock type and overall depth of the open pit.

Resource from the open pit will report to a ROM pad and primary crusher directly northeast of the pit rim. The mill will be fed with material from the pits at an average rate of 2.5 Mt/a (6.8 kt/d), increasing to 3.0 Mt/a (8.2 kt/d) after four years of operation. Resources mined in excess of mill feed targets will be stored in a low-grade stockpile directly south of the ROM pad, and east of the open pit. This stockpile is planned to be completely reclaimed to the mill at the end of the mine life.

Waste rock will be placed in one of two facilities, each planned as a comingled facility with processed tailings. The north facility sits directly adjacent and uphill from the open pit, with its most northern point lying 1.2 km from the pit rim. The south facility sits 0.6 km southeast and downhill of the open pit, with its most southern point lying 2.0 km from the pit rim. The waste rock from the open pit has not been tested or analyzed for potential acid generation (PAG).

Topsoil and overburden encountered at the top of the pits will be placed in a dedicated stockpiles directly south of the open pit and kept salvageable for closure at the end of the mine life.

Mining operations will be based on 365 operating days per year with two 12-hour shifts per day. Owner managed operations are planned, utilizing a diesel-powered mining fleet.

In-pit dewatering systems will be established for the pit. All surface water and precipitation in the pits will be gravity drained, or directed via submersible pumps, to ex-pit settling ponds directly outside the pit limits.

The mine equipment fleet is planned to be purchased via a lease financing arrangement, with down payments occurring when the equipment is commissioned, and lease payments deferred for one year after the equipment is operational.

Maintenance on mine equipment will be performed in the field with major repairs and planned interval maintenance in the shops located near the process facilities.

Pit Designs
In-Pit Haul Roads
In-pit haul roads are designed 25 m wide to facilitate two-way travel for 90 t payload rigid-frame haul trucks. Haul road grades are limited to a maximum of 10%. Access ramps are not designed for the last bench (5 m) of the pit bottom, on the assumption that the bottom ramp segment will be removed using some form of retreat mining. The bottom five ramped benches (25 m) of the pit use one-way haul roads of 19 m width and 12% grade since bench volumes and traffic flow are reduced.

Pit Phases
Ultimate pit limits are generally split up into phases or pushbacks to target higher economic margin material earlier in the mine life. Minimum pushback distances of 50 m are honoured. The Beauty Zone pit is mined as a standalone single pit phase. The main Lemhi deposit pit is split into five phases with the higher-grade, lower strip ratio early pit phases mined ahead of lower grade, higher strip ratio pushbacks to the ultimate pit limit.

Beauty Zone Phase, P621 - This phase targets the high-grade mineralization of the Beauty Zone. The upper benches of this phase will be accessed via ex-pit ramps to the 1,790 masl on the south side of the pit, wrapping around the hill side to the main deposit area.

Starter Phase, P622 - This phase targets the higher-grade, lower strip ratio portion of the deposit outlined by the Case PF 0.41 pit shell. The upper benches of this phase will be accessed via in-pit cut ramps up to 1,680 masl developed during the construction period of the project.

SW Starter Phase, P623 - This phase targets two standalone southwest portions of the main deposit. The upper benches of this phase will be accessed via in-pit cut ramps up to 1,680 masl developed during the construction period of the project.

West Pushback, P624 - This phase targets deeper, higher waste mining ratio mineralization below and west of the P622 pit, outlined by the Case PF 0.56 pit shell. The pit highwall is pushed to the final limits in the west.

South Pushback, P625 - This phase targets deeper, higher waste mining ratio mineralization south of the P622/P624 pits, outlined by the Case PF 0.65 pit shell. The pit highwall is pushed to the south with room for further southwest and southeast pushbacks to the final pit limits.

Final Pushback, P626 - Final Pushback, P626 - This final pit phase targets several pit bottoms west, south, and north of the initial pit phases. A standalone pit to the northeast is developed off a one-way ramp from the pit exit at 1,585 masl, to the pit bottom at 1,550 masl. The P623 SW starter phase is extended north and west to a new pit bottom at 1,510 masl. The remaining pit is pushed out to the north, west and south, utilizing existing ex-pit and previous phase in-pit ramps located between the 1,710 masl and the pit exit at 1,565 masl.

The production schedule is based on the following parameters:
• The mineral resource and associated waste material quantities are split by pit phase and bench quantities.
• An annual mill feed rate of 2,500 kt/a (6.8 kt/d) is targeted.
• This is increased to 3,000 kt/a (8.2 kt/d) in Year 5 of the Project.
• Mill throughput ramp-up is assumed to occur in the construction phase, such that the first year of mill operations is at the target mill throughput. Low grade resources are planned to be stockpiled well in advance of the mill rampup period.
• Within a given pit phase, each bench is fully mined before progressing to the next bench.
• Pit phases are mined in sequence, where the second pit phases do not mine below the first pit phases.
• Pit phase vertical progression in mineralized area is limited to no more than 36 m in each year, or 9 benches; average annual phase progression is 28 m.
• Pre-stripping done in the construction period, Years -2 and -1, is done to open the pits sufficiently to supply mill feed at the target throughput rate in Year 1 of the Project.
• Resource tonnes released in excess of the mill capacity are stockpiled, including those mined in the construction phase.

Comminution

Crushers and Mills

Summary:

Crushing Circuit
Run-of-mine (ROM) material is hauled from the mine and stockpiled on the ROM pad with one day storage capacity or directly tipped into to the hopper equipped with a static grizzly. Material from the hopper is discharged by gravity to an apron feeder and fed into the primary jaw crusher where it is crushed to product size (P80) of 56 mm. The primary crushing plant has an operating availability of 75%. The jaw crusher discharge is then transported to the SAG mill feeder hopper via the overland conveying system. Material will be transferred into the mill via SAG mill feed conveyor. The hopper will be designed as an overflow bin such that overflow material from the hopper will be transported to an emergency stockpile by a conveyor should hopper capacity be exceeded. The crushing plant and associated materials handling equipment is sized at the outset for the Phase 2 throughput.

Major equipment in this area includes:
• ROM hopper with static grizzly
• Primary jaw crusher (160 kW)
• Primary crusher conveyor
• SAG mill feed hopper
• emergency stockpile and feed conveyor
• SAG mill feed conveyor.

Grinding Circuit
An overland conveyor will deliver the crushed material to the grinding circuit consisting of a SAG mill followed by ball mill in closed configuration with hydro cyclones. The circuit is sized based on a grinding circuit feed size (F80) of 58 mm and a circuit product size (P80) of 110 µm. SAG mill slurry will discharge onto a rubber-lined trommel screen with trommel oversize discharging to a bunker for regular collection and disposal. The trommel undersize will combine with the ball mill discharge in the cyclone pump box where the slurry will be diluted to the desired pulp density with process water and pumped to the cyclone cluster. Overflow from the cyclones at 44% solids w/w will report to a trash screen followed by the leach circuit. Cyclone underflow will return to the ball mill directly for further size reduction. In the Phase 2, the ball charges of both the SAG and ball mill will be increased, and the target circuit product size (P80) will be increased to 130 µm to accommodate for the throughput expansion.

Major equipment in this area includes:
• SAG mill (2 MW)
• ball mill (4 MW)
• cyclone cluster
• cyclone overflow trash screen.

Processing

• Crush & Screen plant
• Carbon re-activation kiln
• Smelting
• Agitated tank (VAT) leaching
• Carbon in leach (CIL)
• Carbon adsorption-desorption-recovery (ADR)
• Elution
• Dewatering
• Filter press
• Solvent Extraction & Electrowinning
• Cyanide (reagent)

Summary:

The process plant design incorporates a staged expansion approach allowing the throughput to be expanded. The process flowsheet for the Lemhi Gold Project was tailored to support the rampup of the plant throughput in Phase 2 starting from fifth year of operation. The unit operations selected are standard technologies used in gold processing plants.

The staged expansion of the process plant over the mine life is presented below:
• Phase 1 (Years 1 to 4) – The process plant is operated at a throughput of 2.5 Mt/a (6,849 t/d).
• Phase 2 (Years 5+) – The pre-leach thickener is added, and grind size is increased to 130 µm to process material at throughput of 3.0 Mt/a (8,219 t/d).

The process plant features the following:
• primary crushing of ROM material
• SAG mill followed by ball mill with cyclone classification
• leach and carbon-in-leach (CIL) adsorption, a pre-leach thickener will be added for the expansion
• acid washing and elution of loaded carbon
• electrowinning and smelting to produce doré
• carbon regeneration
• cyanide destruction and wet tailings disposal.

The crusher plant circuit design is set at 75% availability and the gold room availability is set at 52 weeks per year. The plant will operate two shifts per day, 365 days per year, and will produce doré bars.

Leach and Adsorption
Hydrocyclone overflow gravitates to the leach and carbon-in-leach (CIL) area via a trash screen. The trash screen will remove any debris or trash from the slurry before leaching. Trash screen undersize slurry will be fed into the leach/adsorption circuit consisting of one leach tank and six CIL adsorption tanks providing total residence time of 36 hours. Air is sparged to the tanks to maintain adequate dissolved oxygen levels for leaching. Hydrated lime is added to adjust the operating pH to the desired set point of 10.5-11 and cyanide solution is added to the first leach tank.

Regenerated carbon from the carbon regeneration circuit is returned to the last tank of the CIL circuit and is advanced counter-currently using carbon advance/transfer pumps from a downstream to an upstream tank. Slurry from the last CIL tank gravitates to the cyanide detoxification tanks. Each CIL tank has a mechanically swept carbon retention screen to retain the carbon while allowing the slurry to flow by gravity to the downstream tank. Loaded carbon is transferred from the first CIL tank to the loaded carbon screen followed by the carbon elution circuit using a recessed impeller pump. The leach-adsorption circuit is sized for the expansion throughput in Phase 1.

For the expansion, a pre-leach thickener will be added to dewater trash screen undersize slurry to 50% solids to maintain 36 hours residence time through the leaching circuit.

Major equipment in this area includes:
• one mechanically agitated leach tank
• six mechanically agitated CIL tanks with interstage screens.

Cyanide Detoxification and Tailings Disposal
CIL tailings exiting the last CIL tank passthrough a carbon safety screen and are pumped into two cyanide detoxification tanks in parallel. Carbon retained on the safety screen is removed into bulk bags. Cyanide detoxification will take place using the SO2/air process. In this process copper sulphate is used as a catalyst. Hydrated lime is used to maintain the pH of the reaction. The cyanide detoxification makes use of two tanks in parallel that have each been sized for a total residence time of 90 minutes. The cyanide destruction tanks are equipped with oxygen addition points and agitators to ensure that the oxygen and reagents are thoroughly mixed with the tailings slurry. The tailings slurry discharges into final tailings pumpbox and pumped to the North Co-placement Storage Facility (CPSF). A filter plant will be added in the second year of operation to produce filtered tailings for placement in the CPSFs. The details of the CPSFs are described in section 18.4. The cyanide detoxification circuit and associated material handling equipment is sized at the outset for the expansion throughput in Phase 2.

Major equipment in this area includes:
• carbon safety screen
• two mechanically agitated detoxification tanks
• tailings pressure filter (installed in Year 2).

Carbon Acid Wash and Elution.
Loaded carbon slurry is pumped from the first CIL tank to the loaded carbon screen. Screen undersize is pumped back to the first CIL tank, while screen oversize discharges to the acid wash column. Loaded carbon will be washed with a weak hydrochloric acid solution at 2 BV/h rinse rate to remove impurities and residual leaching reagents that could render the elution less efficient or become baked on in subsequent steps and ultimately foul the carbon. Entrained water will drain from the column and the column will refill with the hydrochloric acid solution from the bottom up. Once the column is filled with acid, it will be left to soak, after which the spent acid will be rinsed from the carbon and discarded to the final tailings pump box.

The acid-washed carbon will be hydraulically transferred to the elution column for gold stripping via a pressure Zadra system. Hot elution solution consisting of a mixture of water, sodium hydroxide, and sodium cyanide is passed through the carbon bed to desorb the gold and other adsorbed species from the carbon surface. Pregnant solution from the elution column is transferred to electrowinning. Electrowinning barren solution is then recirculated through the elution column via a heater. A heat exchanger preheats the barren eluate by recovering some heat from the pregnant solution. When an elution cycle is complete, the circuit is ready to initiate a new acid wash and elution cycle. The acid wash, elution and carbon regeneration circuits are sized at the outset for the expansion throughput in Phase 2.

Major equipment in this area includes:
• loaded carbon screen
• acid wash column
• elution column
• recovery heat exchanger
• elution heater.

Carbon Regeneration
The stripped carbon is dewatered by a screen over a feed hopper that feeds an electric rotary kiln via a screw feeder. The kiln is operated at 750°C in an atmosphere of superheated steam to restore the activity of the carbon. Carbon discharging from the kiln will be quenched in water and pumped over a carbon sizing screen to remove undersized carbon fragments. As carbon will be lost by attrition, fresh carbon is added to the circuit as needed in the carbon quench tank. Carbon sizing screen oversize reports to the last CIL tank while undersize slurry is discharge into final tailings pumpbox. Major equipment in this area includes:
• stripped carbon dewatering screen
• regeneration kiln
• carbon sizing screen.

Electrowinning and Gold Room
Gold is recovered from the elution pregnant solution by electrowinning process. The pregnant solution is pumped through electrowinning cells fitted with stainless steel mesh cathodes. An electrical current is applied across the cells, causing gold to deposit on the surface of the cathodes. Expected electrowinning plating time is 16 hours. Barren solution is recirculated to the elution columns with a periodic bleed to the leach circuit in order to prevent the buildup of impurities. The gold-rich sludge is washed off the steel cathodes in the electrowinning cells using high-pressure spray water and gravitates to the sludge hopper. The sludge is filtered, oven dried, mixed with fluxes, and smelted in a single pot furnace to produce gold doré. The electrowinning and smelting process takes place within a secure and supervised gold room.

Water Supply

Summary:

The wells on site will provide potable water for the site, as well as water for the building facilities and the process plant.

Plant
Freshwater
Fresh water will be provided to a freshwater storage tank, where it will be further pumped for various application points, including reagent preparation, gland seal, elution circuit, and general mill make-up water supply. Approximately 450,000 m3 /a of fresh water will be required for make-up to the process plant.

Potable Water
Potable water is produced by an on-site potable water plant which processes water from the freshwater tank and makes it fit for consumption and human use. Potable water is stored in a tank for distribution to the processing plant.

Process Water
Process water will be made up of tailings reclaim water, contact water and freshwater make-up. After a filter plant is added in the second year of operation, process water will be made up of filtrate and wash water from tailings filter and freshwater make-up. Process water will be stored in a process water tank and pumped to various circuits in the process plant.

Fire water
Fire water for the process plant is sourced from the freshwater tank. A dedicated pump skid consisting of an electrical pump, jockey pump, and diesel pump will supply water from the fire water reserve volume to a fire water reticulation system that services the plant. The fresh water tank level will maintain a minimum level of water for use by the fire water system.

Gland Seal Water
Gland seal water is taken from the fresh water tank and pumped to various pumps throughout the processing plant, including sump pumps.

Production

Operational metrics

Personnel

Job Title	Name	Email	Profile	Ref. Date
Chief Executive Officer	Bassam Moubarak	bm@bmstrategiccapital.com		Jan 15, 2025
Consultant - Infrastructure	Scott Elfen			Oct 13, 2023
Consultant - Infrastructure	Jonathan Cooper			Oct 13, 2023
Consultant - Mining & Costs	Marc Schulte			Oct 13, 2023
Consultant - Recovery Methods & Costs	Kevin Murray			Oct 13, 2023
VP Exploration	Dean Besserer			Jan 16, 2025