The Cape Ray Gold Deposit (CRGD) consists of electrum-sulphide mineralisation that occurs in boudinaged quartz veins within an auxiliary shear zone (the “Main Shear”) of the Cape Ray Fault Zone (CRFZ). The boudinaged veins and associated mineralisation are hosted by chlorite sericite and interlayered graphitic schists of the Windsor Point Group (WPG), with sulphides and associated electrum occurring as stringers, disseminations and locally discrete massive layers within the quartz bodies.

The 04, 41 and 51 Zones occur along a northeast-trending fault within the CRFZ system, which dips moderately (50-60°) to the southeast. These zones consist of complex tabular zones of quartz veins, fault gouge and wall rock fragments, range from several cm to a few m in width, and correlate laterally for up to 700 m along strike. In section, the 04 and 41 Zones show eastsoutheast to southeast plunges and locally show down-dip extension of up to 300 m (Dubé & Lauzière, 1997). The 51 Zone is not as consistently defined but both sub-horizontal and steeply plunging grade trends are present.

Within the CRGD, the A vein typically consists of milky white breccia veins with various proportions of angular clasts of wall rock (chlorite and graphitic schist) set in a matrix of quartz, and containing up to 40% sulphides (averaging 5-10%). The vein is typically up to 2 m wide with local wider pods. Fault gouge commonly borders mineralised veins and is superimposed on already highly strained rocks. The nature of the host rocks and superimposed deformation make this mineralised zone complex and variable in terms of vein continuity, width and grade.

Underground investigations by Dolphin (Arnold, 1988) within the 41 Zone noted that the A vein pinches and swells, and commonly branches out into smaller discrete veins or lenses. Similar observations were made by Wilton and Strong (1986) in a trench that exposed the A vein within the 51 Zone for over 61 m. They observed numerous quartz veins (up to 2.4 m thick and 21 m long) sub parallel to the main fabric that show well developed pinch and swell structures. Wilton and Strong (1986) suggested that the podiform nature of the veins resulted from boudinage of a formally more continuous quartz vein. Outside of the main mineralised zone (30-50 m laterally), the A vein structure consists of either uneconomic quartz veining or fault gouge with little to no veining with traces of disseminated pyrite. Brittle deformation is superimposed on the veins as cataclastic features.

The C vein is typically located up to 30 m down-dip on the footwall of the A vein and is generally less deformed. Historical drill hole data indicate that the C vein is up to 15 m thick, and gold values are not as consistent or as high grade as those within the A vein (Dubé & Lauzière, 1997). Underground investigations of the 41 Zone showed the C vein to consist of a series of fault-fill mineralised quartz breccia intervals measuring up to 1.5 m in thickness. The transition between mineralised and barren quartz breccia veins is locally gradual, with the barren veins located in the central part of the mineralised quartz breccia. Fragments within the quartz breccia veins are typically angular and composed of graphitic schist and sericite-chlorite schist set in a matrix of hydrothermal quartz.

In the 51 Zone, the mineralised quartz is at the top of the graphitic schist unit within a gouge zone. The A vein is broken up into large lenses and breccia fragments of varying size. The graphitic schist in this zone varies from less than 15 m thick in the west to nearly 61 m thick in the east. Locally mineralised quartz veins and/or quartz stockwork intervals are found within the footwall schists. High grade mineralisation of good continuity is commonly restricted to the top of the graphitic unit (Ford, 1985).

Minerals of economic interest in the CRGD consist of varying amounts of galena, chalcopyrite, sphalerite and pyrite with associated electrum and free Au. Pyrrhotite occurs as minor inclusions in other sulphide grains and the primary iron oxide is magnetite, mainly found in wall rock fragments rather than veins. In places, deformation has recrystallized galena and chalcopyrite, shattered the pyrite and removed sphalerite. All sulphides contain complex intergrowths of euhedral pyrite crystals, and sphalerite typically contains rounded chalcopyrite inclusions. Electrum occurs as minute grains generally less than 0.02 mm in size, but up to 0.8 mm and dominantly inter-grown with pyrite, and to a lesser extent with chalcopyrite, galena and sphalerite or some combination of these sulphides. Electrum also occurs as solitary grains within the quartz veins in regions of high sulphide concentrations. Paragenetic features indicate that pyrite formed first followed by simultaneous deposition of the other sulphide minerals (Wilton, 1983).

The relative location and geometry of the 04, 41, 51 and WGH deposit to surface warrant an open pit mine scenario. The 04 deposit proposed plan incorporates a pit with the remainder of the deposit mined underground using a longhole stope method with cemented and uncemented rock fill.

The 04 underground deposit would be accessed by one ramp from pit bottom, with broken rock transported to surface in underground haul trucks.

The access ramp for the 04 zone would be developed from the 04 pit bottom. The ramp would be 509 metres long and driven at a grade of 18%.

Remuck bays would be excavated every 150 m along the ramp. The ramp would have dimensions of 4.5 m by 4.5 m to accommodate the underground haul trucks, and also to provide a reasonable cross-sectional area for ventilation. Safety bays would be spaced at 30 m intervals, as required. Services located in the main ramp would include compressed air line, water line, discharge water line, electrical power cables, blasting line and a leaky feeder cable.

The 4.5 m wide by 5.5 m high muck bays would be located on each mucking horizon. These loading areas would accommodate a Load-Haul-Dump (LHD) loading a 30 tonne truck.

Levels would be developed with access from the main ramp to the LH stopes. The levels would consist of a haulage drift in waste running the entire strike length of the zones to be mined.

Stopes would have a nominal dimension of 15 m along strike by 3m wide with sub level intervals of 15m.

Mineral extraction would commence from the bottom stope in each area and continue upwards. Primary stopes would be filled with cemented rock fill (CRF) and secondary stopes with uncemented rock fill.

Longhole open stoping requires an undercut and top sill drift to be developed for each stope. The undercut sill is the extraction level from which stopes are mucked out. The top sill is initially the drilling horizon. After a stope has been backfilled, the top sill is used as a mucking horizon for the longhole stope directly above the backfilled stope.

Mucking and backfilling activities would be accomplished utilizing remote controlled load-hauldump (LHD) units.

The proposed Cape Ray production schedule would begin with the open pit mining of the 04 deposit. The 04 open pit would take approximately 1 year to complete and generate approximately 270,000 tonnes of mill feed and 5.7 MT of waste that would be made available for underground backfill. The 04 pit would supply initial feed to the mill and surface stockpile.

Once the 04 open pit is complete, towards the middle of year 1, the 04 pit crew and equipment would move to the 51 deposit and an underground contractor crew would commence driving the access for the underground portion of the 04 Zone. The 51 open pit would take approximately 2 years to mine producing approximately 475,000T of mill feed and generating 10.6 MT of waste rock. The 51 open pit mining would be complete early in year 3. The 04 Zone underground production would start in year 2 and take approximately 1 year to complete mining approximately 151,000T of resource.

Upon completion of mining in the 51 pit, crews would commence mining in the 41 open pit. The 41 open pit would begin early in year 3 and finish in the later part of year 5. It would generate approximately 630,000 tonnes of mill feed and 5.9 MT of waste rock.

When the 41 pit is complete, the contracting crews would transfer to the largest pit on the property, Window Glass Hill. Mining would commence in the later part of year 5, and finish in year 9. Total mill feed produced would be 1.4 MT while generating 8.1 MT of waste rock. Once completed, the contractor pit crews and equipment would demobilize.

The proposed plan is to operate the open pits for 6 months of the year, from May to November producing approximately 2,000 t/d of mill feed with the surplus stockpiled beside the mill. The annual open pit mill feed is anticipated to be approximately 336,000 tonnes per year.

The underground deposits would be mined yearround with an anticipated production rate of 315 t/d for an annual production rate of approximately 113,500 tonnes.

The total feed to the mill over the 9 year mine life would be 2.9 MT and would be processed by the mill at a rate of 1000 t/d or 336,000 tonnes per year.

Run of mine Potential Mining Resource would be delivered to the mill feed stockpile at a rate of 1000 t/d for six to seven months of the year from the open pit while underground production would continue throughout the year. The proposed processing rate is 46 t/h or 1000 t/d with the mill operating at 92 % availability 365 days per year. ROM (Run of Mine), at– 203.2 mm, is first reduced to – 25.4mm in a single jaw crusher. Further size reduction to–12.7mm is achieved in a series of two gyratory cone crushers operating in closed circuit with a double deck screen. The crushed product is stockpiled ahead of the grinding circuit and is then fed to grinding via a primary grinding mill feed conveyor.

Size reduction from – 10 mm to 80 % passing 100 microns is achieved in the grinding circuit. Crushed product is fed to two parallel primary ball mills at 46 t/h. Primary mill discharge is directed by a screen and the screen underflow reports to a pair of gravity concentrators for recovery of any coarse free gold. The screen oversize reports back to the primary mill feed. The gravity concentrator tailing reports to the secondary grinding circuit for further size reduction before reporting to the CIL circuit thickener.

Prior to cyanide leaching of the contained gold in the PMR, the pulp is thickened to 45 % solids by weight using the CIL pre-leach thickener. The thickened pulp is advanced in the CIL circuit at the full process flow rate of 46 t/h solids. In the CIL circuit gold is recovered from the solids using whole ore cyanidation and the gold is transferred to the surface of the activated carbon. The activated carbon is separated from the main pulp flow by a combination of pumping and screening. The loaded carbon is washed with hydrochloric acid solution to remove the carbonates. Gold is then removed from the loaded carbon using pressure stripping. The gold is transferred from the surface of the loaded carbon to solution and from there is recovered to marketable form using the electrowinning process.

The stripped carbon is regenerated in a reactivation kiln before being reintroduced to the process. Fine carbon is constantly eliminated (and recovered) from the process to avoid gold loss, with fresh carbon being continuously added to the process.

The cyanide contained in the tailings from the CIL circuit is eliminated in a cyanide destruction tank with SO2-air process. Once the cyanide is destroyed, the tailings are transported to the tailings pond for disposal.