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 production rate for the underground mine is assumed to be 2.16 Mtpa (6,000 tpd) and will be reached after 4.5 years of pre-production and 1.5 years of ramp-up with owner-operated mining equipment.

Four gold bearing mineralized lenses were delineated and modelled during the resource evaluation process, namely the Main Zone, situated below the old Wasamac Mine, Zones 1, 2, and 3. These mineralized lenses extend vertically for nearly 850 m and over 1.8 km east to west. A six-metre diameter concrete shaft will permit direct access to three main levels of mineralized lenses. All underground development will be trackless and two service ramps will provide access to all mining levels spaced every 30 vertical metres, for each zone. The main service ramp will also access the surface.

The underground mining method recommended by RPA is a longhole mining with transverse accesses from the deposit footwall through the hanging wall, considering zones geometry and targeted production rate. When the thickness of the zone is less than eight metres RPA recommends the longitudinal mining method for economical reasons.

Primary and secondary stope dimension will be 25 m along strike by 30 m floor-to-floor in the transverse method. The same strike length opening of 25 m will be considered for the longitudinal stoping approach.

The mining sequence will proceed upward from mining horizons following an inverted Vshape.

Mining will incorporate the following activities:
• Lateral development performed with hydraulic jumbos (drilling) and mechanical bolters (ground support).
• Vertical development performed with Alimak (vent raise/second egress, ore pass, waste pass/fill raise) and V30 type drill or Raise Bore (42-in dia.) (slot raises).
• Production drilling carried out with longhole drills.
• Blasting using ANFO (ammonium-nitrate fuel oil) in both development and production operations.
• Production blasting will be initiated with electronic detonators.
• Loading and hauling operations performed with load-haul-dump (LHD) units (scooptrams) and underground trucks for waste and gold bearing material (ore).
• Backfilling of primary stopes with cemented hydraulic fill and backfilling of secondary stopes with non-cemented hydraulic fill or unconsolidated rockfill.

Production drilling for the 100 mm diameter and for the 150 mm diameter will be carried out with a down-the-hole (DTH) drill.

Gel cartridges or pumpable slurries will replace ANFO under wet conditions or up holes. Production blasting will be initiated with electronic detonators. The better precision of electronic detonators permits to blast a whole stope with as few as three consecutive blasts, from bottom to top. This will reduce overbreak and dilution thereby increasing the productivity.

All stopes will be backfilled with either cemented or non-cemented hydraulic fill, or unconsolidated rockfill when waste rock from underground development will be available. Waste rock skipped at surface may be returned underground via a fill raise and hauled to open stopes with LHDs or trucks. Non-cemented hydraulic fill can be used when there is not enough underground waste production.

Cemented hydraulic fill will be placed in all primary stopes and some parts of the secondary stopes when the zone thickness exceeds 30 m from hanging wall to footwall and there is a tertiary stope filled with unconsolidated rockfill to complete the retreat sequence. Cemented hydraulic fill will also be placed in all stopes of longitudinal longhole mining method. The secondary stopes will be backfilled with non-cemented hydraulic fill or unconsolidated rockfill and surrounding stopes will be filled with cemented hydraulic fill.

Two processing options for the Wasamac material were investigated: (1) whole rock leaching and (2) flotation followed by leaching of both the tailings and reground concentrate products. Results of the metallurgical testwork program indicated that the combined flotation/leaching process resulted in higher overall gold recoveries over those obtained by whole rock leaching. Using the testwork results and the mining plan provided by Richmont, a technico economic trade-off study comparing the two options was conducted. The results of the trade-off study revealed that the whole rock leach option was in fact more economically favourable in spite of the higher recoveries obtained in using the flotation/leaching processing option due to lower capital and operating expenditures. The process flow sheet was therefore designed based on the whole rock leaching option.

The processing plant is designed for a capacity of 2.1 Mtpa with an average hourly throughput of 260 tph. The average head grade estimated from the mine plan is 2.25 g/t Au. A head grade of 2.75 g/t Ag was calculated as a weighted average of the silver assays for the metallurgical testwork samples.

Run of mine material from the crushed rock storage silo is conveyed to a 500 t capacity surge bin adjacent to the concentrator. From the surge bin, the crushed rock, with a P80 of 150 mm, is reclaimed and conveyed to a 3,730 kW (5,000 HP) semi-autogenous (SAG) mill fitted with a trommel screen for primary grinding to a P80 of 1,500 µm. The SAG mill discharge feeds the secondary closed-loop grinding circuit made up of a cyclone cluster and a 2,835 kW (3,800 HP) ball mill. The cyclone underflow feeds the ball mill for grinding to a P80 of 120 µm while the overflow reports to the tertiary closedloop grinding circuit. Two cyclone clusters split the material sending the underflow to a second 2,835 kW (3,800 HP) ball mill for a final size reduction to a P80 of 40 µm. The cyclone overflow reports directly to the pre-leach thickener, 23 m in diameter, to dewater the slurry to a solids density of 50% (w/w) prior to leaching. The thickener overflow is pumped back to the process water tank for recycling in the process.

The leaching circuit consisting of six tanks, 15.5 m in diameter each, and is designed to provide a total retention time of 48 hours. The leach discharge slurry is pumped to a carousel carbon-in-pulp (CIP) circuit where the leached gold is adsorbed onto carbon in eight 100 m3 cells. Upon exiting the CIP circuit, a screen recovers the loaded carbon and the gold depleted slurry is sent to the pre-detox thickener. The loaded carbon is sent to the stripping circuit where an acidic solution strips the gold and silver from the carbon. The coarse spent carbon is regenerated in a kiln for re-use in the CIP circuit while the fine carbon is bagged and sold for gold and silver credit. The gold and silver bearing pregnant solution is then pumped to a series of electrowinning cells where the precious metals are recovered on the cathodes in the form of a sludge. The gold and silver are subsequently washed from the cathodes, filtered and dried. The dried gold and silver is mixed with flux material before smelting into a doré bar.

The tailings slurry from the CIP circuit is pumped to the pre-detox thickener, 23 m in diameter, for dewatering. The overflow is pumped to the process water tank for re-use throughout the processing facility. The underflow is pumped to a cyanide destruction circuit in which the Inco SO2/air process reduces the residual cyanide concentration to below 5 ppm. The underflow slurry at 60% (w/w) solids is either pumped directly to the tailings pond, or is stored in agitated tanks to be used in the production of hydraulic backfill. Approximately 35% of the tailings produced will report to the hydraulic backfill plant where they will be mixed with a cement/flyash mixture prior to being pumped back to the mine. The production of backfill will be variable, with a maximum production of 3,000 tpd.

It is estimated that 70% of the water contained in the hydraulic backfill becomes available for recycling to the mine water storage basin. As the tailings deposited in the tailings pond continue to settle, an estimated 65% of the water sent to the tailings pond becomes available for recycling to the process water tank, while 35% remains locked within voids. It was estimated that 702 m3/h of process water and 59 m3 /h of fresh water are required. With recycled water being added to the process from the pre-leach and pre-detox thickeners, as well as that reclaimed from the mine water basin and tailings pond, approximately 28 m3/h of make-up water collected from precipitation and run-off will be required to satisfy process needs.

Mill reagents, including cyanide, lime, sulphur dioxide, caustic soda, nitric acid and carbon will be delivered to the site by transport truck and stored in the processing facility as required.

The overall gold recovery of the proposed circuit is estimated at 90.2% with additional silver recovery of 74.6%.