The Abcourt-Barvue silver-zinc deposit is classified as disseminated volcanogenic sulphide deposit. It shares numerous similarities with the Mattabi-type volcanogenic sulphide deposit. The zinc and silver sulphides mineralization is composed of disseminated and bedded sphalerite and pyrite found in close association with felsic volcanoclastic rocks.

The Abcourt-Barvue mineralized zone has a thickness ranging from 2 to 30 metres. The deposit has a known east-west strike length of 2.2 km. Mineralization has been delineated by diamond drilling to a maximum vertical depth of 425 m below the surface. Mineralized horizons are dipping at 75° to the north.

Zinc and silver are the main metallic constituents of the Abcourt-Barvue deposit and these metals are associated with sphalerite, native silver and argentite. The presence of some gold and copper has been observed. Gold is usually associated with silver and copper occurs mostly at the eastern limit of the Barvue deposit.

The main mineralization which is mostly made up of disseminated and bedded sphalerite and pyrite (up to 10% of the rock) is described as - iron poor, honey brown and dark colour sphalerite (ZnFeS2), finely disseminated in the tuff agglomerate horizons concentrated generally to the north of the marker bed, with additional mineralization in the marker bed and in the shear zone. - 1-2% disseminated pyrite (some encompassed within the sphalerite); - minor galena (PbS), chalcopyrite (CuFeS2), native silver (Ag) and proustite (Ag3AsS3).

The thickness of the mineralized horizon tends to increase eastward, most of the sulphides being located in the Barvue area. Generally speaking, we can observe a zinc enrichment in both footwall and hanging walls of the mineralized zone which are generally separated by lower grade marginal material. These two sub-parallel zinc-enriched units are currently known over a total length of 2.2 km to a vertical depth of 600 metres (Hole AB-92-05) in the mine sector (Vachon, 1994).

The gangue minerals include siderite, quartz, chlorite, chloritoïd, sericite, illite and rhodocrosite (Cornwall, 1955).

Open pit mining:

Mining will be carried out using conventional open pits method. Drilling will be performed by conventional production drills. Blasting operations will use an emulsion-ANFO (ammonium nitrate fuel oil) and a down-hole delay initiation system. Hydraulic backhoe shovels and a front-end loader will be used to load rigid haulage trucks of 62-t capacity.

For most of the open pit tonnages (ore and waste), haulage trucks will exit from Abcourt and Barvue pits onto a common haul road approximately 0.5 km west of the plant site. Waste rock will be hauled to two (2) different storage areas, one located 0.9 km from the exit of the Barvue pit on its north side and the other located 0.2 km from the Abcourt pit exit to the south of the Abcourt pit.

Open pit mining will occur at a maximum annual production rate of 6 Mt of rock per year of which 650 000 tonnes (1 800 tpd) will be ore and the remaining 5 350 000 tonnes will be marginal ore and waste rock. More details are given on the following pages.

Underground mining:

For the remaining measured resources (not exploited by open pit) and part of the indicated resources, the Abcourt-Barvue ore is found generally in the hanging wall of a marker tuff where ground conditions are good. For this zone, the Avoca cut-and-fill method will be used for the 6th to 10th year of production. From three main declines, cross-cuts will be driven into the ore zone and lateral development will proceed on both sides to the limits of the lenses. Production will start at the bottom of the indicated resources with sub-levels developed at 15 to 20-metre vertical intervals. All mining will be trackless.

Jumbos will be used for decline, access cross-cut and sub-level development. In the Avoca stopes, the ore will be removed with remote controlled 5 or 2-cubic-yard scoops. The benches will be drilled with long hole machines. The main declines will be driven with a minus 15 to 17 % slope.

The main declines, haulage drifts and access cross-cuts will be 5.5-m wide x 4.5-m high. The sub-drifts in ore will be 4.5-m high over the full width of the ore zone. Each stope will have an access cross-cut for each sub-level. The 1.5-m x 1.5-m slot raises will be drilled with a long hole machine. The Avoca method produces 100 % recovery of the ore.

The stopes will be backfilled with waste and paste or cemented backfill. The backfill will consist of mill tailings and/or pyrite concentrate.

The underground production will come from three Avoca cut-and-fill stopes which will be named «west stope», «central stope» and «east stope», and from mine development. The stopes extend over a distance of about 1 125 metres from 5010E on the Abcourt side to 135E on the Barvue side. All the ore tonnage which is planned to be mined into the aforementioned limits is located west of the limit separating Barvue from Abcourt and is therefore not subject to royalties.

The west and central stopes will start at a depth of about 150 metres and finish at a depth of about 55 metres in pit bottoms. The east stope will start at a depth of about 200 metres and will be completed at an elevation of about 65 metres, also in pit bottoms. Each stope will have two faces in ore, so it will be possible to alternate the drilling, blasting and mucking of ore cycle at one end of the stope with the filling cycle at the other end. The objective is to have a steady flow of muck from each stope.

Development will produce about 5 000 tonnes of ore per month. Once fully developed, each stope will produce about 7 350 tonnes per month for a maximum of 22 000 tonnes of ore per month. In total, the underground mine will have a capacity to produce 27 000 tonnes of ore per month.

The main process steps for treating the Abcourt and Barvue ores are primary crushing and stockpiling, SAG/ball grinding, primary thickening and double stage cyanidation of the ores, gold/silver solution recovery by double stage filtration and washing, gold/silver solution clarification, precipitation of dissolved gold & silver by zinc dust, refining of gold/silver precipitates, aeration of cyanide tails, zinc flotation from aerated cyanide tails, and dewatering of a zinc concentrate by thickening and filtration. There is also a pyrite (or sulphide) concentrate produced as the last stage of differential flotation. This pyrite (or sulphide) concentrate is disposed of in a safe sulphide containment cell.

Barvue and Abcourt ores are mixed on outside stockpiles located near the crushing facilities. The ore is picked up and carried by a front end loader over a feed hopper equipped with an incline 500 mm square openings grizzly with the undersize dropping into a lined 50 t hopper. Mixing the ore will regulate the grade fed to the processing plant. The ore is extracted from the feed hopper by a reciprocating feeder equipped with a slotted 140 mm openings grizzly section at its discharge end. This feeder discharges in a jaw crusher. The crusher will be a single toggle jaw crusher with a 150 kW motor and a 1220 mm by 810 mm (48” x 32”) feed openings.

A 1800 t coarse ore bin will provide a buffer of 24 hours of plant operation between the crusher and feed to the SAG mill. The 9 m (29.5’) diameter by 15 m (49’) high coarse ore bin will be covered and space heaters will be installed to reduce air moisture in the winter. The bin walls will be covered with spray-on insulation to prevent freezing.

The ore is withdrawn from the coarse ore storage bin by a 1000 x 5000 mm apron feeder. The feeder discharges on a 914 mm belt conveyor carrying the ore to a semi-autogenous (SAG) mill 5490 mm diameter x 2135 mm long driven by a variable speed 600 kW motor.

The SAG mill discharges onto a single deck vibrating screen. The screen oversize is returned to the SAG mill feed via two 610 mm oversize return belt conveyors in a row, the last one discharging onto the SAG mill feed belt conveyor. The fist oversize return belt conveyor is equipped with a belt magnet to remove metals and the last one with a belt scale. The screen undersize is directed by gravity to the ball mill discharge pump box. The ball mill is 3960 mm diameter x 5500 mm long and is driven by a synchronous 1500 kW motor.

The dilute cyclone overflow is gravity fed to a 20 meter diameter thickener. A flocculent solution is added to the thickener feed to promote high underflow densities and clean overflow water. The thickener underflow at a density of 50 % solids is pumped to the primary cyanidation circuit. The thickener overflow proceeds by gravity to a surge tank and then is pumped to the pregnant solution tank.

The primary cyanidation circuit is composed of two 10000 mm diameter x 10500 mm high mechanical agitators in series. The discharge from the second agitator is the feed to the primary filtration circuit which is composed of two 3660 mm x 4880 mm long drum filters equipped with snap blow discharges. Filter cakes are repulped to 50 % solids with barren solution and pumped to the secondary cyanidation circuit. Primary filtrates are returned to the primary thickener to be recovered as pregnant solution.

The secondary cyanidation circuit is composed of three 10000 mm diameter x 10500 mm high mechanical agitators in series. The discharge of the third agitator feeds the secondary filter circuit composed of two 3660 mm diameter x 4880 mm long drum filters. Filter cakes from secondary filters are repulped with water and pumped to the zinc flotation aeration tank. Filtrates from secondary filters are pumped to the mill solution tank and become the primary make-up water for the grinding circuit. All stages of filtration employ a ‘flood’ wash of barren solution for washing the soluble precious metals from the solids. The wash displacement ratio on the filters is 1.5. Each twin drum filter station is serviced by a complete vacuum system including filtrate receivers, filtrate pumps and a wet vacuum pump with silencer.

The precious metals are recovered from solution by zinc precipitation. The pregnant solution is first clarified in a 30 bag clarifier using diatomaceous earth as the filter medium then de aerated in a vacuum Crowe 2000 mm diameter x 5000 mm high tower tank. The zinc dust is added to the precipitation pump suction pipe. The precipitation pump pumps the precipitate slurry to one of two pre-coated 914 mm x 914 mm x 10 plate and frame filter presses. The barren solution going out of the in-service filter-press is directed by gravity to the barren tank. The barren solution is used for washes on primary filter station, cyanide solution make up, grinding solution make up and for slurry pump gland water service everywhere excepted in flotation and dewatering areas. A bleed to tailings via the pyrite tails pump box.

The filter press cake is melted to buttons with standard fluxes in an electrical arc tilting smelter furnace. The primary slag will be returned to the grinding ball mill feed after hand crushing. Buttons are remelted with additional borax and bullions are poured into 1000 and 400 oz. molds. The slag from remelting buttons is remelted in the smelter furnace with the next charge of precipitates.

The oxidized cyanide pulp from the 60 minute retention aeration tank is pumped by one of two horizontal slurry pumps to the first of two zinc flotation conditioning tanks. Based on the required conditioning time, an active volume of 27 m3 will be provided via two tanks. The first will be 3050 mm in diameter and 3350 mm high and will be used for lime conditioning. The second will be of same dimensions and will be used for activation with copper sulphate. Both tanks will provide 0.3 m of freeboard.

The last conditioning tank feeds by gravity the combined zinc rougher and scavenger flotation bank (5 x 8.5 m3 cells + 3 x 8.5 m3 cells). The scavenger rejects are pumped by one of two pumps to the pyrite (or sulphide residual) flotation conditioning tank. An inline sampler is delivering a continuous scavenger tails sample stream to the on stream analyzer. The scavenger concentrate is pumped by an integrated tank pump to the first cell of the rougher bank for refloating and upgrading. The rougher concentrate is pumped by an integrated tank pump to the feed of the first cleaner flotation bank (2 x 1.4 m3 cells) which is followed by the cleaner scavenger bank (2 x 1.4 m3 cells).

The cleaner scavenger rejects are pumped to the zinc flotation second conditioning tank. The cleaner scavenger concentrate launder exit is piped back to the first cleaner flotation cell feed. The first cleaner rejects are transferred to the first cell of the cleaner scavenger bank by a cell to cell displacement.

The first cleaner concentrate launder exit is piped back to the first cell of the 2nd cleaner bank (2 x 1.4 m3 cells). The 2nd cleaner rejects are transferred to the first cell of the first cleaner bank by a cell to cell displacement. The 2nd cleaner concentrate is pumped by an integrated tank pump to the Zn concentrate thickener as a final concentrate.

The rougher flotation feed, the 2nd cleaner concentrate and the cleaner scavenger tails are sampled by inline samplers which deliver continuous sample streams to the on stream analyzer.

The 8-minute retention flotation conditioning tank (3050 x 3050 mm) feeds by gravity the rougher flotation bank (10 x 8.5 m3 cells). The rougher rejects are pumped by one of two pumps to the tailings pond. An inline sampler is delivering a continuous rougher tail sample stream to the on stream analyzer.