The Morrison deposit is classified as a calc alkaline copper-gold porphyry with an alkalic trace element signature (Ogryzlo et al., 1995), which may reflect a mixed alkaline/calc alkaline parentage for the Babine Igneous Suite. The geologic settings of the host rock relationships, structural development, and the general style of the hydrothermal alteration and mineralization at Morrison are similar to other porphyry deposits in the northern Babine Lake area.

Hydrothermal alteration at Morrison is similar to that at other Babine porphyry copper deposits (Carson and Jambor, 1974). Alteration is concentrically zoned with a central biotite (potassic) alteration core surrounded by a chlorite-carbonate zone. A third alteration facies, clay-carbonate alteration, is considered retrograde and associated with major faults and shears and subsidiary fracture zones. No well developed phyllic zone has been identified.

Sulphide mineralization at Morrison shows strong spatial relationships with the underlying intrusive (BFP) plug and associated alteration zones. The central copperrich core is hosted mainly within a potassically altered BFP plug with intercalations of older siltstone. This plug was initially intruded into the siltstone unit as a near-vertical subcircular intrusion approximately 700 m in diameter. It was subsequently disrupted by the East and West Faults and now forms an elongated body extending some 1,500 m in the northwest direction.

Chalcopyrite is the primary copper-bearing mineral and is distributed as fine grained disseminations in the BFP and siltstone, as fracture coatings or as stockworks of quartz veinlets in which the chalcopyrite occurs as coarse grains (1-3 mm) within veinlets that range from 1.0 mm to approximately 15 mm in width. Minor bornite occurs within the higher grade copper zones as disseminations and associated with the quartz-sulphide stockwork style of mineralization.

The open pit mine has been designed for four phases of pit development by conventional truck-shovel methods. The overall life-of-mine (LOM) strip ratio is relatively low at 0.825:1.0 waste to ore.

The current development concept for the mine is as follows:
- Phase I: West pit starter pit with a ramp that commences on the north side counter clockwise
- Phase II: East pit with stand alone access road from the north clockwise
- Phase III: West pit expansion with multiple switch backs
- Phase IV: East pit expansion.

When the pit bottom of Phase I and Phase II are mined, the pit bottom will be approximately 40 m wide.

Phase III and Phase IV pit expansion phases are typically no less than 60 m in width, which should provide adequate room for efficient mine operations.

Processing of the ore will be performed on-site using crushing, grinding, and flotation to recover copper and molybdenum concentrates.

The flowsheet design is based on 30,000 t/d mill capacity with an overall mill availability of 92%, which allows for scheduled and non scheduled maintenance.

Run-of-mine (ROM) material from the pit is delivered by 227 t capacity haul trucks that dump directly to the 1,372 mm x 1,905 mm primary gyratory crusher. The gyratory crusher discharge is collected by a 1,830 mm apron feeder under the crusher. The feeder then transfers the crushed ore onto a 1,524 mm conveyor, which in turn discharges onto the crushed ore stockpile. The total live capacity of the stockpile is 30,000 t. A dust collection system is installed to minimize dust emissions.

Three 1,200 mm x 5,500 mm apron feeders withdraw ore from the stockpile onto the 1,372 mm wide x 201,300 mm long primary sizing screens feed conveyor at an average rate of 1,359 tonnes per operating hour. The primary sizing screens are 2,700 mm wide x 8,000 mm long with grate apertures of 75 mm and 45 mm. The oversize material (+45 mm) from the both screens is feed to the cone crusher, powered with 750 kW motor.

The undersize product from the primary sizing screens (-45 mm) is fed to the HPGR surge bin. The material is fed to the HPGR through a belt feeder and then to the secondary sizing screens. The secondary sizing screens are 4,000 mm wide x 8,000 mm long with grate apertures of 20 mm and 6 mm.

The oversize product (+6 mm) from the secondary sizing screens is returned to the HPGR surge bin through series of conveyor belts.

The undersize product (-6 mm) is mixed with water before being fed to the grinding cyclones feed pumpbox. Water is added with the ore to form slurry in the cyclone feed at about 55% solids.

The two 6.1 m x 10.2 m ball mills are powered with 6,700 kW motor. The ball mills are in a closed circuit with a one stage cyclone package.

Aero 3302 and PEX, as the flotation collector reagent, are added to the ball mill pumpbox. Other reagents if required may also be added to the pumpbox.

The pumpbox slurry is pumped to the two sets of 660 mm primary cyclones (5 operating, one stand-by each), with the underflow returning to the ball mill. The primary cyclone overflow, at approximately 35% solids and a pH of about 8, is advanced to the copper flotation circuit for further recovery. The particle size range is 80% less than 150 µm.

The primary grinding cyclone overflow (from the grinding section) is directed to the copper rougher flotation. The slurry from the cyclones is fed to the four 200 m 3 mechanical cells. The copper rougher tailings are scavenged using four 200 m 3 mechanical cells prior to discharge to the tailings impoundment. MIBC (a frother) and possibly more reagents are added as required.

The copper rougher/scavenger concentrates are collected in a pumpbox and further reagents and makeup water are added as required. This is then sent to the regrind mills cyclopac. The underflow drops into the regrind mills feed chute and returns for regrinding. Each regrind mill processes about 140 to 150 t/h. Lime is also added to the regrind mill to maintain a pH of about 11.5.

The regrind cyclone overflow, at approximately 25% solids, is directed to the first cleaner flotation cells. The particle size P80 is less than 25 µm. The tailing of the first copper cleaner stage is scavenged using conventional mechanical cells with the concentrate being re circulated to the regrinding mill feed. The tailings product is added to the copper rougher/scavenger tail as final tailings.

The concentrate from the first copper cleaner stage is then advanced to the copper second cleaner flotation. The second cleaner is using a 3.1 m Ø x 10.0 m column cell with concentrate advanced to the molybdenum flotation circuit and tailings recycled to the copper first flotation.

The concentrate from the copper second flotation is fed in an 11 m Ø coppermolybdenum concentrate thickener. The thickened concentrate is fed to the three 5 m 3 mechanical cells. The molybdenum rougher tailings together with the coppermolybdenum and final molybdenum concentrate thickeners overflows are advanced to the copper dewatering circuit. The molybdenum rougher concentrate is further processed to the single cleaner stage circuit (column cell). The final molybdenum cleaner concentrate is predicted to be achieved with a grade of 45 to 50% Mo.

The final copper concentrate is thickened in an 11.4 m Ø concentrate thickener with the aid of flocculent. The thickener overflow is recycled to the process water tank for reuse. The underflow, at 60% solids, is pumped to a pressure filter.

The copper concentrate filter cake is expected to have between 8 and 10% moisture. The cake is conveyed to a stockpile. The water discharged from the filter is recycled to the concentrate thickener to recover the fine solids in the water.

The final molybdenum concentrate is thickened in a 2.5 m diameter thickener with the aid of flocculant. The thickener underflow at 60% solids is pumped to a filter. The molybdenum concentrate filter cake is expected to have between 10 and 15% moisture. The cake is conveyed to a standard rotary dryer to further reduce the moisture content to less than 5%. The final molybdenum concentrate product
discharges into a bin prior to packaging.

An automated bagging machine will package the dried molybdenum concentrate into two-tonne tote bags at a rate of one bag per hour. These bags are weighed and sampled before moving them to a storage area. The molybdenum concentrate production is approximately 0.01% of the mill feed tonnage averaging 8 t/d.