The Hardrock Project lies within the granite-greenstone Wabigoon Subprovince of the Archean Superior craton, in eastern Canada. The Wabigoon Subprovince, averaging 100 km wide, is exposed for some 900 km eastward from Manitoba and Minnesota, beneath the Mesoproterozoic cover of the Nipigon Embayment, to the Phanerozoic cover of the James Bay Lowlands (Card and Poulsen, 1998). The Wabigoon Subprovince is bounded on the south by the metasedimentary Quetico Subprovince, on the northwest by the plutonic Winnipeg River Subprovince, and on the northeast by the metasedimentary English River Subprovince. The Wabigoon-Quetico Subprovince boundary is a structurally complex, largely faulted interface.

The Hardrock Property is located within the Beardmore-Geraldton Greenstone belt that contains several narrow, east-west striking sequences of volcanic and sedimentary rocks of Archean age. The southern edges of these sequences are spatially related to the through-going, major structural discontinuities thought to be thrust faults that have imbricated the sedimentary sequences. A comprehensive description of the regional geology can be found in Smyk et al., 2005. In the Geraldton area, most of the gold mines and a number of gold showings occur within or in proximity to the Bankfield-Tombill Deformation Zone (also known as the Barton Bay Deformation Zone), a zone of folding and shearing up to 1 km wide. The southern limit of the Tombill-Bankfield Deformation Zone is marked by the Tombill-Bankfield Fault, a zone of intense shearing up to 12 m wide.

In the immediate Geraldton area, the dominant rock types are clastic sediments (greywacke and arenite), oxide facies iron formations (“BIF”) and minor mafic metavolcanics. There are a number of younger intrusives, including an albite-rich porphyry unit (Hard Rock Porphyry) that is spatially associated with much of the gold mineralization on the Hard Rock, MacLeod-Cockshutt and Mosher Mines. Significant gold mineralization is also often spatially associated with BIF. In the case of the Little Long Lac Mine, gold mineralization is primarily hosted by an arkosic unit.

Gold mineralization in the Hard Rock, MacLeod-Cockshutt, Mosher Mines and the Little Long Lac Mine generally occurs in association with subvertical structures associated with quartz veins or stringers, minor to semi-massive sulphides (associated with replacement zones in BIF), weak to moderate carbonate and weak to strong sericite alteration. The ore zones rake shallowly towards the west in the vicinity of the Hard Rock, MacLeod-Cockshutt and Mosher Mines (15-30° W) and slightly more steeply towards the west at the Little Long Lac Mines (50-60° W), indicative of a strong structural control that post-dates the tight folding of the primary lithological units.

The gold mineralization occurs in a variety of host rocks and the style of mineralization is partly a function of the host rock. While the location and overall orientation of the ore bodies appear to have been largely structurally controlled, the deformation of the ore bodies has not been as intense as that of the host rocks. Nevertheless, there are areas where local folding and boundinage of mineralized veins is apparent. Additionally, there are strong secondary controls that influence the extent and intensity of gold mineralization such as the competency contrast between host rocks (e.g. the Hard Rock Porphyry and its contacts with either wacke or BIF) and the chemical character of the host rocks (e.g. oxide facies BIF being replaced by sulphides).

The following discussion on mineralization was taken from Smyk et al. (2005). “Gold mineralization in the BGB has resulted from the introduction of hydrothermal fluids in zones of high crustal permeability (Smyk et al., 2005). Permeability was generated by prolonged, multiple periods of deformation, which focused not only fluids, but magmatic activity and intrusions. In the Hardrock Deposit area, a major zone of deformation in which the gold mines are located has been alternatively termed the Bankfield-Tombill Fault Zone (Pye, 1951; Horwood and Pye, 1951) or the Tombill-Bankfield Deformation Zone (Lafrance et al., 2004, and herein).

Most mineralized occurrences in the Hardrock Deposit area lie in a zone of deformation to the immediate north of, and genetically linked to, the Tombill-Bankfield Deformation Zone. This zone of deformation varies from 600 to 100 m in total width, while the crush zone of the Tombill- Bankfield Fault proper ranges from metres to hundreds of metres in width. Gold mineralization is associated with D3 brittle shear zones and folds overprinting regional F2 folds (Lafrance et al., 2004). The plunge of the mineralized zones is parallel to F3 fold axes and to the intersection of D3 shear zones with F2 and F3 folds. On a sub-province scale, regional folds cut by D3 dextral shear zones are promising targets for discovering the next generation of large gold deposits.”

The Project consists of developing an open pit that will mine through the historical underground workings of the MacLeod-Cockshutt and Hard Rock Mines. Furthermore, the proposed open pit location is bisected by the Trans-Canada Highway 11 and requires a new by-pass and the relocation of various surface infrastructures.

Mining of the Hardrock main pit will occur in five main phases preceded by a starter pit.

The phase designs introduce different geotechnical slope profiles for temporary pit walls. The temporary wall slope profile allows for wider catch benches to allow for overbank hazard management on pit walls. Overbank hazard results from muck from one phase spilling down the slope of the previous pit phase, filling the catch benches. This creates an increased rockfall hazard for workers and equipment at the bottom of the previous pit phase.

The objective of pit phasing is to improve the economics of the Project by feeding the highest grade during the earlier years and/or delaying waste stripping until later years. The starter pit and Phase 1 are designed to initiate mining before the Trans Canada Highway is relocated. With the mineralization plunging westward, the pit phases progressively expand to the west.

The starter pit phase is designed to avoid various surface constraints such as the Trans-Canada Highway 11. Mining during the pre-production period is concentrated in the starter pit which provides for a 50 m buffer with the Trans-Canada Highway 11. The highway will be relocated to the north during the initial construction period. This starter pit phase reduces risk with respect to the timing of the highway relocation.

The Phase 1 design continues the constraint of the 50 m buffer from the Trans-Canada Highway 11 but without the constraint of avoiding historical tailings. Phase 1 pit uses the same ramp from the starter pit and descends significantly deeper. On the west side of Phase 1 is a flattened bench. This area will act as a plateau and dumping grounds for the historical tailings. Moving the tailings will allow the mining to the north. This plateau will be preserved throughout Phase 2 until it is eventually properly removed.

Phase 2 is restricted by the historical high tailings of a buffer of 50 m. The Trans-Canada Highway 11 has been displaced by this point and is removed to allow mining underneath. The plateau is kept with an independent ramp as the material can be transported away in this phase. A ramp on the north wall is added to facilitate the future mining of Phase 2.5 without requiring an additional external ramp.

Phase 2.5 continues with the main ramp from Phase 2, utilizing the internal ramp to mine material not normally accessible. There are no immediate constraints limiting this design and it is driven by profitability.

Phase 3 begins a new temporary ramp for access to the ore body until the depth of the pit is more then Phase 2.5, that the haulage is shifted to the main haulage ramp from 2.5. This explains the seemingly abrupt end to the Phase 3 temporary ramp when the switch is made between ramps.

Phase 4 is the maximum depth of the pit. Phase 4.5 is an internal sub pit that is mined after Phase 4 is complete. This is to ensure that the main haulage lane always has a side wall during production and is less prone to geological faults and sloughing. Phase 4.5 also cuts the main haulage lane into a single lane in a part. This can only be done when the main haulage from the main pit is completed.

Phase 5 is the eastern extension of the pit. There is a new independent ramp that will service all the haulage from the phase. This ramp will connect to and replace the main haulage lane created in Phase 2. Phase 5.5 are two sub pits on the edge of the Phase 5 pit. There are no constraints on when they are mined and are driven by the optimized schedule.

Waste rock will be disposed of in five distinct waste rock storage areas (“WRSA”) of which four are located around the pit and one further to the south. The open pit generates 675.9 Mt of waste rock, 1.6 Mt of backfill, 3.16 Mt of historical tailings and 8.7 Mt of overburden that require storage. The tailings material will be transported for disposal within the Tailings Management Facility (“TMF”).

The design criteria of each waste dump has been adjusted based on foundation stability assessments. All rock waste dumps have 20 m high lifts to allow for wider catch benches to facilitate reclamation. Overburden dumps use 5 m high lifts to better control the material.

The stockpile is designed with a maximum capacity of 12.9 Mt of ore. The total capacity is never reached throughout the mine life with the max stockpile material peaking at 9,652 kt in Year 12 ensuring a safety factor to account for changes in mine plan or stockpiling. There are four grade bins of stockpiled material ranging from marginal material to high grade. These bins are to be stored within the main stockpile in separate piles to reduce dilution.

The objective of the crushing circuit is to reduce the size of the run-of-mine ore to the required particle size for the downstream HPGR and ball mill circuit. The crushing plant is a two-stage circuit consisting of a primary gyratory crusher and a secondary cone crusher. The crushing plant has a design availability of 67%. A 20% design factor has been selected such that crushing circuit equipment is sized to handle up to 2,025 t/h. The design factor is industry practice, providing extra production capacity to handle processing fluctuations due to changes in ore feed rate and ore hardness.

Primary Crushing
Run-of-mine ore is delivered by mine haulage trucks. The ore is dumped from the truck into a steel dump pocket that feeds the primary crusher. The pocket has two dump points and a live capacity equivalent to 1.5 times the haul truck payload. A rock breaker will be used to break oversized rocks. The 1,300 mm x 1,800 mm 450 kW gyratory crusher crushes the ore from a 1,000 mm top size (275 mm P80) to a P80 of 120- 160 mm product. The primary crusher operates with an open side setting (“OSS”) of 160 to 200 mm. The crushed ore falls into a steel discharge pocket with 1.6 truckloads capacity and is reclaimed via an apron feeder. The main discharge and dribbles from the apron feeder are discharged onto the sacrificial conveyor that feeds the secondary crushing and screening circuit.

An electromagnet and metal detector are installed on the primary crusher sacrificial conveyor to prevent tramp iron from entering the secondary cone crusher. The ore contains magnetite in sufficient quantities that would overwhelm a standard magnet. The metal detector is used to detect spikes in magnetic susceptibility from tramp metal, which will then activate the electromagnet to remove the tramp. This feature is utilized for all of the tramp removal systems.

Secondary Crushing and Screening
The secondary crusher is a 950 kW standard cone crusher with a 45-60 mm closed side setting (“CSS”). The secondary cone crusher is installed in closed-circuit with a double deck screen to control the top size feeding the HPGR. Secondary crusher screen undersize is conveyed to the covered crushed ore stockpile.

The 4,250 mm x 8,500 mm double deck secondary crushing screen is fitted with a top deck 75 mm closing screen opening, and bottom deck with 50 mm closing screen opening. Combined oversize from both decks is returned to the secondary cone crusher. The crushing circuit produces a final crushed product with a 50 mm top size and a 35 mm P80. Combined screen oversize flows onto the secondary screen oversize discharge conveyor and transfers to the secondary screen oversize return conveyor which feeds a mass flow bin with 17 minutes retention time. A second tramp metal electromagnet is installed before the secondary screen, and a metal detector-activated flop gate on the screen oversized recycle to protect the secondary crusher. In case tramp metal is not removed successfully, the secondary crusher retractable belt feed conveyor is equipped with a metal detector which will retract the chute and deposit the tramp steel into a separate chute. These measures prevent tramp iron from entering the secondary cone crusher.

Crushed Ore Stockpile and Reclaim
The crushing circuit product is stored in a 21,196 t live capacity stockpile which provides 20 hours of operation. The stockpile is located North of the secondary crushing and the process plants. There is a single stockpile reclaim tunnel with three apron feeders located in a reinforced concrete tunnel underneath the stockpile. These apron feeders feed the crushed ore stockpile (“COS”) reclaim conveyor which discharges into the HPGR feed bin located inside the HPGR building. Cartridge type dust collectors are installed in the transfer chutes between the apron feeders on the skirt boards of the COS reclaim conveyor. Spile bar isolating systems are installed on each of the apron feeder reclaim hoppers, to isolate the apron feeders for maintenance.

High Pressure Grinding Rolls (HPGR)
A detailed comminution trade-off study recommended a two-stage crushing circuit followed by HPGR and ball milling circuit over other typical comminution flowsheets such as crushing followed by semi-autogenous (“SAG”) milling and ball milling, to reduce throughput risk and increase energy efficiency from the high hardness levels measured in testing.

Wet screen oversize is recycled back to the COS reclaim conveyor, which discharges a blended product into the HPGR feed bin (mass flow bin equipped with a slide gate for isolation). A belt feeder reclaims ore from the HPGR feed bin and feeds the HPGR weigh bin, located above the HPGR. Belt feeder speed is controlled in order to ensure that the weight bin choke-feeds the HPGR. The HPGR is equipped with two, 2,650 kW motors for a total of 5,300 kW. The HPGR roll dimensions are 2.2 m in diameter by 2.0 m in length and have a rotating speed of 22 rpm. The HPGR discharge falls onto the wet screen feed bin conveyor and then into the wet screen surge bin feed (mass flow bin equipped with slide gates for isolation) that divides the HPGR product between two double-deck screens.

Grinding
Undersize from each HPGR screen product is discharged to their respective ball mill pump box. Each pump box is equipped with two slurry pumps, one pump feeding the ball mill cyclones and one feeding the gravity circuit. Approximately 60% of the fresh feed is pumped to the gravity circuit.

The cyclone overflow P80 ranges from 72 µm to 90 µm. The ball milling circuit recirculating load is estimated at 300%, with a design value of 350% for pump selection. The cyclones are 600 mm diameter installed in a radial distributor. There will be seven cyclones operational per mill at a pressure of 105 kPa. Each distributor also has two installed spare cyclones.

The cyclone overflow feeds the pre-leach thickener trash screen while the underflow (approximately 75% solids) is directed to the ball mill feed chute. Lime is added to the ball mill feed to raise the slurry pH to between 10 and 11. The grinding mills are twin pinion ball mills equipped with motors totaling 10,500 kW per mill. Both mills are 6.7 m in diameter (inside liners) by 12.3 m in length (EGL). The ball mills have discharge trommel screens to remove scats. The ball mill discharge flows into the ball mill pump box where it is combined with the HPGR discharge slurry and pumped to the cyclones for classification. The plant can be operated at a lower throughput, by operating only one ball mill and one HPGR screen, slowing the HPGR RPM to accommodate.

The gold recovery process for the Hardrock Project consists of a crushing circuit (primary gyratory and secondary cone), a HPGR and ball mill grinding circuit with gravity recovery, pre-leach thickening, cyanide leaching, carbon-in-pulp (“CIP”) adsorption, elution and regeneration circuit, electrowinning and refining, cyanide destruction and tailings deposition.

Gravity Concentration
The gravity feed pump transfers a portion of the cyclone feed to the gravity circuit to recover gravity recoverable gold. Two gravity screens and two gravity concentrators are installed to process the material.

The vibrating gravity feed screen prevents particles coarser than 3.36 mm from entering the gravity concentrator. The screen oversize flows into the ball mill pumpbox. Screen undersize flows to the gravity feed chute onto each gravity concentrator. A by-pass line is installed on the gravity feed chute to the gravity concentrators to operate when the concentrators are trans ........