The Wasamac deposit is a gold deposit of replacement type with close coexistence of gold and pyrite disseminated in the altered Wasa Shear Zone. Hydrothermal alteration is well developed, and alteration minerals have distinct zonation from the orebody outward: albite-pyrite to carbonate-hematite to muscovite-chlorite. Gold mineralization is closely associated with these alterations, especially albite-pyrite alteration.

The Wasa Shear Zone runs through the centre of the property in an east-west fashion. This shear zone trends at an azimuth of 265°, has a 50° to 60° dip to the north and a maximum thickness of 80 m. It is characterized by the development of a strong mylonitic fabric and an intense hydrothermal alteration that destroyed the primary structures and textures of the protolith. Mineral assemblages of rocks within the shear zone consist of chlorite, carbonate, hematite, albite and sericite in the middle of the zone. Gold is associated with a dissemination of fine pyrite in the altered portions of the shear zone.

The Main Zone can be described as a well laminated mineralized zone. It is located near the centre of the property, within the Wasa Shear Zone and high grade areas display true widths of 10 m to 15 m (up to 25 m locally) over a strike length of 400 m. Gold mineralization is associated with quartz, carbonate, sericite, albite, pyrite and chlorite inside the shear zone. Visible gold is rare, and strong gold assays are generally associated with high silica content and a lot of fine grained pyrite (Gill, 1947). If the entire mineralized zone is considered, including lower grade parts, the width of the mineralized zone can be up to 50 m. At depth, in the western part of the Main Zone, there is also gold mineralization in the footwall of the shear zone.

Located some 400 m east of the Main Zone, Zone 1 has a similar mineralogical assemblage. The high grade part displays true widths of 4.5 m to 7.5 m over a strike length of 150 m. The thickness of the mineralized envelop can be up to 20 m.

The higher grade part of Zone 2 has an average thickness of three to six metres over a strike length of 225 m. This zone was partially developed from underground, but no production was recorded. This mineralized zone is located in the upper part of the shear zone, near the hanging wall.

The mineralization of the Zone 3 is located in the lower part of the shear zone, near the footwall, below the MacWin Zone.

The MacWin Zone is also associated with the Wasa Shear Zone. The gold mineralization went out of the shear zone, as irregular mineralized zones are found in the rhyolite located just at the hanging wall of the shear.

Located approximately 300 m south of the Main Zone, the Wildcat Zone was the first gold showing to be discovered on the property (1936). This gold bearing zone consists of a carbonate altered zone at the margins of a gabbroic unit. Gold mineralization is associated with quartz carbonate veinlets containing fine grained pyrite. The pyrite mineralization is also present throughout the altered halo as disseminations.

The Wasamac underground mine consist of five zones: Main Zone (“MZ”), Zone 1 (“Z1”), Zone 2 (“Z2”), Zone 3 (“Z3”), Zone 4 (“Z4”). It is located approximately 15 km west of Rouyn Noranda, Quebec, Canada, within the heart of the Abitibi gold mining camp. The mine design strategy is based on high speed development and Rail-Veyor® technology to complete the pre production infrastructures from years 2021 to 2023 and to achieve an average ore production rate of 6,000 tpd from years 2024 to 2031 when the mine is running at full capacity. The strategy is to recover higher grade zones in the early years of the Project.

The Rail-Veyor® Integrated Top Down Long Hole Stoping Method was selected for the Wasamac Project based on the geometry of the mineralized zones, vertical dip, as well as economic factors. There are two types of mining methods: one with 40 m interval Rail-Veyor® main levels and another with 20 m interval Rail-Veyor® main levels. The 40 m interval involves accessing two levels of stopes using three longitudinal rubber tire drifts. The drift above the top stope is used for drilling, loading, blasting of the top stope and backfilling of the two levels of stopes. The drift above the bottom stope drift is for down hole drilling, loading, blasting of the bottom stope. The drift below the bottom stope at the main level is for mucking material onto the reclaim feeder that loads the Rail-Veyor® situated 30 m to the south of the orebody in the hanging wall.

The 20 m interval method involves accessing one level of stopes using two longitudinal rubber tire drifts, one above the stope for down hole drilling, loading, blasting and backfilling, and one below the stope at the main level for mucking. The ore handling process for both types of mining methods is the same. The empty stopes are backfilled with paste backfill from the underground paste fill plant. Stopes will be mined following a retreating sequence towards the closest twin ramps. Many horizons can be mined at the same time to maintain productivity level.

In areas with wide ore thicknesses, there will be addition of sill drifts that are in between two rubber tire drifts but shifted 4 m below. This will allow the recovery of stope apexes.

The development sequence focuses on prioritizing the high grade mineralized zones, specifically MZ. The mining sequence requires the excavation of an opening a V-30 slot raise from Machines Rogers International. Once the development is completed, the mineralized zone is surveyed with precision for the preparation of the drilling and blasting pattern. Waste material generated from the drift development will be transported by Rail-Veyor® and used to create the paste backfill of the stopes.

The mining method requires the use of paste backfill to provide better ground stability and lower dilution while mining. As recommended by Outotec, the maximum slurry feed to batch will be 6,000 tpd, and 100% unclassified tails feed from the mill will be pumped to the backfill plant or tailings facilities. This approach would avoid the need of double systems and management of complex tailings. Classification has been omitted because it would add both complexity and risk to the pressure filtration for disposal, which likely outweigh the likely benefits from classification. Double panel topes will be backfilled in one shot from the top panel to the bottom panel, once the bottom panel is mined out. All stopes will be filled with paste backfill with an estimated 7% binder content, and when and where appropriate, with a blend of rock fill and paste backfill, to reduce the amount of waste development rock to be hauled to surface.

The process plant will consist of crushing, stockpiling, primary and secondary grinding, leaching prior to a carbon-in–pulp (“CIP”) circuit, followed by cyanide destruction. Based on the mine backfill scheduling, material will be pumped underground to a paste backfill plant. When the paste backfill plant will not be operating, tailings will be pumped to a filtration plant where the filter cake will be sent to a dry-stack tailings pile. An adsorption-desorption-recovery (“ADR") circuit and a gold room will recover gold from activated carbon and will produce doré. The plant will also include a reagent preparation area, two process water circuits (cyanide bearing and cyanidefree), a fresh water circuit, as well as low-pressure and compressed air distribution systems.

Ore will be transported from the underground mine by a Rail-Veyor® conveying system. The RailVeyor® carts will dump into a material handling system that will transport the ore into the crusher . The material will be withdrawn from the Rail-Veyor® discharge bin using an apron feeder that will feed a vibrating grizzly feeder. The grizzly oversize will be directed to a jaw crusher to reduce the material to a P80 of 120-140 mm. The jaw crusher product and grizzly screen undersize will be collected on a conveyor belt feeding the crushed ore stockpile, which will be covered by a dome.

The reclaimed coarse ore will be conveyed to the SAG mill. Water will be added to the mill feed chute to achieve the desired slurry density within the mill. Given the grindability results, a SAG mill size of Ø7.92 m x 4.72 m (Ø26’ x 15.5’) was selected. The mill will be installed with a single 5,500 kW motor. The mill drive train will include a single-pinion arrangement powered by a lowspeed synchronous motor of 5,500 kW (7,376 hp). This selection allows for the ore to be ground from an F80 of 140 mm to a P80 of 0.9 mm.

A ball mill, Ø5.49 m x 9.75 m (Ø18’ x 32’), fitted with a trommel screen was selected based on the design metallurgical energy requirement of 15.0 kWh/t. The maximum required power is equal to 5,500 kW (7,376 hp).

The ball mill will be operated in closed circuit with a cluster of Ø400 mm cyclones producing an average product P80 of 60 µm as trash screen feed, with a pulp density of 30% solids (weight basis). A sufficient number of cyclones will be provided to cover a circulating load of up to 350%, whereas the nominal flow is based on 250%. The ball mill may be charged to 35% of its volume with Ø38 mm steel balls added intermittently via a ball bin located at the ball mill feed chute. Water will be added to the cyclone feed pumpbox to control the cyclone feed slurry density.

The cyclone overflow will be sent to the linear motion trash screen ahead of the pre-leach thickener while the oversize material in the underflow will return to the ball mill for further grinding.

The ball mill area will be serviced by an overhead crane used for maintenance duties and for grinding media addition to the mill, from a ball bin located within the process plant building. One sump pumps, will be installed to collect spillage directed towards the sump with sloping basement floor.

Prior to leaching, the ground slurry will pass through a trash screen before being thickened to 60% solids (weight basis) in a Ø22 m pre leach thickener. The thickener was sized using dynamic settling test results. Flocculant will be added to the thickener to improve the solids settling rate. The underflow from the thickener will be fed to the leach tanks by two slurry pumps, one operating and one standby. The thickener overflow water will feed the Ø6 m cyanide-free process water tank. The process water pumps will be located inside the plant, in the grinding area. The thickener drive mechanism will be enclosed in a shelter to prevent snow, ice and other weather hazards from damaging the equipment and to facilitate operation and maintenance during winter.

Feed to the leaching circuit is diluted with cyanide-bearing process water in order to control a feed % solids of 40% w/w. Leach cyanide levels will be maintained using a sodium cyanide (“NaCN”) solution. Lime slurry will be added, as required, to maintain a solution pH between 10.5-11. One row of six Ø16.5 m agitated leach tanks will be used to provide a retention time of 48 hours, as established by the testwork program. The tanks will be 20.6 m in height and the slurry will overflow due to gravity between tanks. All leach tanks will be located outside, with slurry flowing from the outermost tank to the plant. Oxygen will be used in the leach tanks to improve reaction efficiency and reduce cyanide consumption.

The leach tanks will be installed on concrete rings resting on rock. A concrete slab and concrete wall, poured on granular backfill will serve as secondary containment around the tanks. Two sump pumps will empty the containment basin back to the leach tanks. Any seepage through the steel bottom of the tanks and concrete slab will be contained by a high density polyethylene (“HDPE”) membrane installed under the tank, in the backfill. Tanks, agitators and agitator drives will be serviced by a mobile crane.

The Wasamac Project process design plans to use a “PumpCell” style CIP circuit to achieve gold adsorption from the leaching slurry. A required total of eight 120 m3 tanks will be installed, with each tank containing 6 tonnes of carbon. The retention time for each tank will be approximately 12-15 minutes.

With the PumpCell system, the slurry will be pumped from one cell to the other, while the carbon remains in each tank until loading is completed. Once the carbon gold grade reaches the desired value in the first tank, the feed will be diverted to the second tank and the content of the tank will be pumped to the loaded carbon recovery screen located on top of the carbon stripping circuit. The carbon retained on this screen will feed the stripping circuit while the slurry and wash water undersize will be recycled to the CIP circuit. The CIP circuit was sized in order to limit the maximum carbon loading to approximately 8,000 grams of gold/silver per tonne of carbon (g/t). One tank from the CIP circuit will be emptied and transferred to the stripping circuit every day.

Carbon elution or stripping will be conducted using the high-pressure Zadra process. A solution of NaCN and caustic soda (“NaOH”) is prepared in a barren solution tank. The stripping solution will circulate in a continuous loop throughout the duration of the stripping sequence, first exiting the barren solution tank and passing through heat exchangers to pre-heat the solution as it enters the elution columns. From there, the stripping solution, having been converted from barren to pregnant solution, will be partially cooled through heat exchangers before reaching the electrowinning (“EW”) circuit. Then, the solution changes from pregnant to barren again and is returned to the barren solution tank, from where it will enter the stripping loop again.

The electrowinning circuit will be made up of two electrowinning cells with the purpose of recovering gold and silver from the pregnant strip solution. The cells will produce a gold/silver sludge, which will be dried and smelted into doré bars as a final product. The now barren solution exiting the cells will report to a discharge tank and from there will be pumped back to the barren solution tank in the stripping circuit. Although stripping is done on a batch basis, the gold recovery cycle is a continuous process.