The principal area of mineralization on the project is the Yellowhead Copper Deposit (the Deposit). The northeast trending Harper Creek Fault separates the deposit into a west domain and an east domain. In the west domain, chalcopyrite mineralization is primarily in three copper bearing horizons. The upper horizon ranges from 60 m to 170 m in width and is continuous along an east-west strike for some 1,300 m, dipping approximately 30º north. Mineralization within this horizon occurs within felsic and mafic volcanics and volcaniclastic rock units. The middle horizon is not as well developed and is often fragmented. It primarily exists within a graphitic and variably silicified package of rocks that range from 30 m to 40 m in width at the western extent, increasing up to 90 m locally eastward, gradually appearing to blend into the upper horizon. Of the three horizons, this contains strong to intense silicification and localized tension fractures filled with mineralization. The lowest or third horizon has less definition mainly due to a lack of drill intersections. Commonly hosted within mafic to intermediate volcaniclastics and fragmental rocks, it can range from 30 m to 90 m in width although typical intersections are in the 30 m range. These horizons host within felsic and mafic metavolcanics and metavolcaniclastics and generally contain foliation-parallel wisps and bands as the dominant style of sulphide mineralization.

In the east domain, mineralization characterized by high angle, discontinuous, tension fractures of pyrrhotite, chalcopyrite ± bornite is frequently associated with quartz carbonate gangue. This style is common within, but not limited to, the metasedimentary rocks and areas of increased pervasive silicification. Mineralization is not selective to individual units and frequently transgresses lithological contacts throughout the area. Locating mineralized horizons in this area has proven difficult due to multiple east-west trending and northward dipping interpreted thrust faults (or possible reverse faults). At the near surface areas in the south and down-dip to the north, widths of mineralization typically range from 120 m to 160 m. In the central area of the east domain where thrust/reverse fault stacking has been interpreted, mineralization thicknesses typically range from 220 m to 260 m with local intersections of up to 290 m. Mafic metavolcanics and coarse-grained quartz-rich metasedimentary rocks generally contain higher grade copper mineralization.

Interpretation of the deposit type is that of a remobilized polymetallic volcanogenic massive sulphide deposit, comprising lenses of disseminated, fracture-filling and banded iron and copper sulphides with accessory magnetite. Mineralization is generally conformable with the host-rock stratigraphy as is consistent with the volcanogenic model. Observed sulphide lenses measure many tens of metres in thickness with kilometer-scale strike and dip extents. In 2009, YMI conducted a program of field mapping, sampling, relogging, petrographic examination of existing thin sections and re-assessment of the total digestion geochemical dataset that confirmed the deposit type hypothesis for the deposit (Armstrong and Hawkins, 2009).

Location and Infrastructure
The Yellowhead project is located approximately 150 km northeast of Kamloops and consists of 131 mineral claims covering 42,500 hectares.

The Yellowhead Development Plan 1 envisions an open pit mine utilizing conventional truck and shovel mining techniques. The equipment utilized in this operation would be typical of that found in today’s large open pit operations. Open pit operations are planned to supply a conventional copper concentrator with 90,000 tpd of ore at a cutoff grade of 0.17% copper. Ore would be delivered to a primary crusher located at the southwestern rim of the ultimate pit. An ore stockpile would be built during the first five years of operation to maximize ore grade delivered to the mill during that period and provide an operating contingency. Potentially acid generating (PAG) waste rock would be stockpiled inside the tailings storage facility while non-acid generating (NAG) waste and overburden would be stockpiled in conventional waste storage locations proximal to the open pit.

The pit design is based on the selected Lerchs-Grossmann pit shell. A single-bench configuration of 15 m high benches is used based on the scale of mining equipment selected. Steeper inter-ramp slopes up to 150 m high are used with enlarged berms or haul roads breaking up larger slopes to honor overall slope requirements. Haul roads are designed 40 m wide to allow for double-lane hauling including allowances for berms and ditches. Single-lane, 27.5 m wide roads are used to maximize ore extraction and mining width at pit bottoms. Road grades are limited to 10% with flat switchbacks.

The total pit waste rock produced would be 1.1 billion tonnes. This includes
• 50 million tonnes of overburden type waste;
• 560 million tonnes of non-acid generating (NAG) waste rock; and
• 500 million tonnes of potentially acid generating (PAG) waste rock.

Overburden waste consists of the unconsolidated materials located above bedrock. Overburden of sufficient quality for reclamation use would be segregated from NAG waste rock and stockpiled in several locations surrounding the pit.

For scheduling purposes, the ultimate pit has been divided into five interim phases. The mine schedule considered the following objectives in order to ensure efficient and practical mining operations:
• Target areas of higher copper grade to maximize copper production early in the mine plan;
• Maintain sufficient mining width on each bench for efficient operations in each phase;
• Limit vertical bench mining rate to no more than 6 benches per year;
• Supply enough non-acid generating (NAG) waste rock to meet material requirements for ex-pit infrastructure construction activities; and
• Provide an efficient ramp system that minimizes haul distances to ore and waste destinations.

Ore is classed into the following three categories: PAG ore, high-grade NAG ore and lowgrade NAG ore using a cut-over grade of 0.25% copper. PAG ore and high-grade NAG ore mined during the pre-production period would be stockpiled within the ultimate pit footprint and processed in year 1. Excess low-grade NAG ore mined would also be stockpiled within the ultimate pit footprint and west of the ultimate pit adjacent to the primary crusher for processing in years 6 through 11.

Pre-production mining focuses on pre-stripping of pit phases 1 and 2, establishment of an ore stockpile to support mill start-up, construction of the main haul roads to the various material destinations, construction of the starter dam for the TSF main embankment and filling the primary crusher pad. Ore mined during the period would be stockpiled inside the ultimate pit footprint.

For the first 5 years of operations, ore supply is planned from pit phases 1 and 2 which would be completed in years 4 and 5 respectively. Mining of both ore and waste from phase 3 would begin in year 4.

Mining in years 6 through 10 is planned from phases 3 and 4 with mining in phase 4 beginning in year 6. Ore would be supplied from phase 3 continuously and from phase 4 starting in year 8. The ore supply would be supplemented with low-grade ore mined and stockpiled in the previous five years.

In years 11 through 15, mining is planned in pit phases 3, 4 and 5 with phase 5 starting in year 11 and phase 3 concluding in year 14. Ore would be supplied from phases 3 and 4 throughout this period with phase 5 mining in waste only. Mining would continue in pit phases 4 and 5 throughout years 16 to 20. Both phases supply ore continuously through this period but only minor amounts of ore are supplied from phase 5 until year 19.

From years 21 through 25 mining would predominately occur in phase 5 with mining in phase 4 completed in year 21.

The crusher facility is designed with a single gyratory crusher with a double-sided dump pocket for the mine haulage trucks. The facility would be located on south-west edge of the open pit to minimize ore haulage distances. The crusher will be serviced by a fixed hydraulic crane and a rock breaker. The crusher and conveyor system have been sized to process ROM ore at design rate of 7,500 tonnes per hour (tph), which is an excess capacity of approximately 45 percent more than process plant throughput. This excess crushing capacity provides operating and maintenance flexibility while minimizing feed disruptions to the process plant. The 80% passing product size generated at the discharge of the crusher is expected to range between 160mm to 250mm, depending on the crusher gap setting. The crusher product would discharge into a surge bin sized to hold approximately two truckloads of material. From the surge bin the crushed ore would discharge via an apron feeder which meters the crushed material onto the conveyor system that transports the ore onto the coarse ore stockpile. The crushing facility would also be equipped with a dust suppression/collection system to control any fugitive dust that is generated during crushing, material loading, and related operations.

The major equipment in this area includes:
• One 1,200 kW gyratory crusher: 1,524 mm x 2,794 mm (60 ft x 110 ft);
• One apron feeder: 2,438 mm x 10,100 mm;
• One hydraulic rock breaker;
• One fixed hydraulic crane;
• One 930 kW 1,828 mm (72 ft) x 360 m sacrificial conveyor;
• Two 1,500 kW 1,524 mm (60 ft) wide overland belt conveyor with a total length of 1.7 kilometers;
• One 1,500 kW 380 m long and 1,524 mm wide stacking conveyor to transport ROM to the coarse ore stockpile;
• Dust suppression systems.

The coarse ore stockpile is designed with a live storage capacity of 45,000 tonnes. The crushed ore would be reclaimed from the stockpile via two parallel conveying systems with three apron feeders installed on each conveyor line.

The apron feeders for each grinding line have been sized to allow nominal design throughput rates to be attained by operating only two out of the three feeders. The reclaimed ore from the apron feeders would discharge onto a belt conveyor, transporting the crushed ore to the SAG mills.

Each SAG mill feed conveyor has been designed with 30 percent excess capacity compared to nominal plant throughout, and would be equipped with a belt scale to measure and meter the SAG mill throughput at a controlled rate. The reclaim area would be equipped with a dust collection system to control fugitive dust generated during loading and transport of the crushed material.

The major equipment in this area includes:
• Six 22 kW 1,219 mm (48 ft wide) x 7,000 mm apron feeders;
• Two 447 kW 1,828 mm (60 ft wide) and 243 m long conveyor belts;
• Dust suppression system.

Primary grinding consists of two parallel SAG mill and ball mill circuits. Variable speed dual pinion driven SAG and ball mills have been incorporated in the circuit design. The grinding mills are driven by water cooled low speed induction motors, and the electrical
drive systems for all of the mills are identical to keep equipment standardized with interchangeable parts.

Each grinding line is designed with a SAG mill which discharges onto a vibrating doubledecked screen equipped with spray bars to wash down any entrained fines on the screen oversize. The screen oversize would be returned to the feed of the SAG mill via a pebble conveying system. Consideration has been made in the design for installation of a future pebble crusher. Screen undersize would report to a primary cyclone feed pump box where it would be combined with the ball mill discharge and pumped via a single centrifugal pump to a hydrocylone cluster. The underflow from the cyclones would be fed to the ball mill and the cyclone overflow advanced by gravity to the rougher flotation circuit. The ball mills have been designed for a circulating load of 350 percent and to produce a product size of 80% passing 180 µm. Reject steel from the SAG mills would be recovered via belt magnets installed on the pebble recycle conveying system, while reject steel from the ball mills would be collected via trommel magnets installed on the ball mill discharge. Steel media would be loaded into the mills via skips from steel media storage bins located on the south wall of the grinding circuit.

The major equipment in this area includes:
• Two dual pinion 17 MW SAG mills – 11 m x 6.2 m (36 ft x 20.25 ft) driven by variable frequency low speed induction motors;
• Two dual pinion 17 MW ball mills: 7.9 m x 13.4 m (26 ft x 44 ft) driven by variable frequency low speed induction motors;
• Two pebble recycle conveying systems, consisting of three conveyors including a high-angle conveyor;
• Two hydrocyclone clusters containing fifteen– 700 mm hydrocyclones per cluster;
• Two vibrating double-deck Screens: 3.6 m x 7.3 m with 7.5 o incline;
• Two Primary Cyclone Feed Pumps: 28x36 – 1,565 kW Slurry Pump.

The equipment used in the regrind circuit includes:
• Two 3,355 kW stirred mills;
• One hydrocyclone cluster containing eighteen (18) 400 mm hydrocyclones (15 operating/3 standby);
• Two 12 x 10 -220 kW hydrocyclone feed pumps (one operation and one standby).

The proposed process plant for the Yellowhead ore is a conventional sulphide concentrator utilizing three stages of comminution, three stages of flotation and concentrate dewatering. The concentrator has been designed for simplicity of operations and maintenance and to meet the project metallurgical targets.

Flotation and Regrinding Circuits
The ground ore from both grinding lines would be combined and processed in the flotation and regrind circuits to recover the valuable minerals. The recovery process would consist of rougher flotation, concentrate regrind, and two stages of cleaner flotation.

The rougher flotation circuit is designed with a single bank of forced air flotation tank cells fed the cyclone overflow product from both primary grinding lines. The rougher flotation circuit would produce a concentrate which would be pumped to the regrind circuit and a tailing stream which would gravity flow to the TSF. Flotation reagents added to the rougher flotat ........