The Gruyere gold deposit is an Archaean orogenic gold deposit. This deposit type is widespread in the greenstone belts of the Yilgarn Craton and in other greenstone belts around the world including in Canada, Africa and India.

Archaean orogenic gold deposits share a number of similar geological characteristics including location in greenstone belts, strong structural control of orebodies, relative timing of gold mineralisation with respect to peak metamorphism, consistent metal association and broadly uniform hydrothermal fluid chemistry. However individual deposits display a diverse range of depositional site characteristics including host lithologies, structural setting, alteration and mineralisation styles, and various aspects of fluid and ore chemistry (oxidation state, gold fineness).

Most orogenic gold deposits are hosted in mafic-ultramafic extrusive and intrusive rocks whereas the Gruyere deposit is hosted in a granitoid. As the Gruyere deposit is a recent discovery and the first major deposit in the Yamarna area, insufficient studies on the mineralised system have been completed to indicate how closely Gruyere compares with major orogenic gold deposits in other locations such as the better known Kalgoorlie Terrane in the central Yilgarn Craton.

Gold mineralisation at Gruyere is characterised by varying intensity albite-sericite-chlorite biotite-calcite alteration, with associated pyrite-pyrrhotite disseminations, and coarse arsenopyrite proximal to high grade zones. Grade commonly increases with the intensity of albite-sericite-chlorite-biotite-calcite alteration; high-grade zones are commonly overprinted with limited porphyry textural retention. Intense alteration and pyrrhotite>pyrite+arsenopyrite mineralisation is often observed proximal to sheared, recrystallised south-east plunging quartz veins.

Lower grade gold mineralisation (commonly < 0.3 g/t Au) within the porphyry is characterised by reddened (hematite dusted) albite-quartz muscovite-biotite-oligioclase-magnetite alteration with only minor pyrite disseminations. Visible gold is commonly observed within brittle-ductile chlorite±pyrite bearing fractures, which are common throughout the porphyry.

While fractionated gold-only porphyry analogues are uncommon, the Canadian Malarctic deposit, hosted along the Cadillac Fault Zone – Abitibi Greenstone Belt, is similar in host lithologies, intrusive lithology, mineralised volume and primary structural features. Both deposits are hosted on major shear or fault zones along secondary dextral events, and include similar intrusive hosts - quartz monzonite porphyry (Gruyere) and quartz monzodiorite porphyry (Canadian Malarctic), intruding into volcanic/sedimentary country rock sequences.

Mineralisation within the Canadian Malarctic is primarily hosted in the host Pontiac group clastic sediments, and with the remaining mineralisation within the quartz monzodiorite porphyry, whereas mineralisation at Gruyere is entirely hosted within the quartz monzonite porphyry. Porphyry geometries vary, with the Malartic porphyry showing multiple dykes extending from a deeper pluton-stoping up into the Pontiac sediments, while the Gruyere porphyry has intruded the host volcanics as a singular intrusive, possibly as a dyke in a higher relative position to a deeper pluton.

Gold mineralisation at Gruyere has developed in response to a complex reverse shearing structural event. The porphyry, a more competent and brittle body compared to the relatively ductile host rocks, responded to deformation with significant cracking and fracturing resulting in increased permeability. Gold-bearing mineralising fluids were able to flow freely through the rock mass, resulting in uniform and disseminated gold mineralisation ubiquitous to the porphyry.

The entire Gruyere Porphyry is variably altered and gold grade can be related to variations in style and intensity, of alteration, structure, veining and sulphide species. Zones containing higher grade gold mineralisation above 1.2 g/t Au generally have strong albite ± sericite ± chlorite ± biotite alteration and are associated with a sulphide assemblage of pyrrhotite + pyrite ± arsenopyrite, weak to moderate foliation, common micro-fracturing and steeply dipping quartz veining.

Sulphides are common throughout the zone of gold mineralisation, with pyrite dominant in the upper areas and pyrrhotite-arsenopyrite increasing with depth. The total percentage of sulphide minerals is generally in the range 0.5-2%. Quartz ± carbonate vein sets observed in diamond core and optical televiewer surveys show multiple character: early shear veins parallel to the shear foliation; late tabular veins (0.01 to 1 metre thick) at high angle to foliation with variable albite alteration halos; veins with strong chlorite margins; chlorite fractures ± albite halos; and fine stock work veins in areas of intense alteration.

Gruyere will be mined by open pit mining methods utilising conventional mining equipment.

Mining activities will be conducted by a mining contractor with technical and managerial direction provided by Gold Road.

The general mining method is summarised as follows:

- Clearing and stripping of suitable material from all disturbed areas into discrete stockpiles
- Drilling and blasting of ore and associated internal waste on 5 metre benches, while bulk waste which is outside the ore envelope is blasted on 10 metre benches. Trim blasts and pre-splits will be used to provide wall control in fresh rock as required. The majority (70%) of the explosives usage is bulk emulsion and the remainder is ANFO, with all explosives supply provided by a subcontractor
- Loading and hauling utilising 360 t excavators and 180 tonne capacity haul trucks mining on 3 metre high flitches in ore zones and three to 4 metre high flitches in bulk waste zones. Ore material is planned to be marked out by paint or tapes on the ground, supported by dedicated ore spotters as required. Ore will be direct fed to the crusher or placed on stockpiles for future rehandle as required
- Waste dumps will be developed in 10 metre lifts and progressively rehabilitated. Raising of the TSF embankment will be constructed with waste material from the mine as required
- Ancillary plant support for floor control, haul road construction and maintenance, rehabilitation, drill support, waste dump battering and the like provided by a fleet of dozers, graders and water carts
- Pit dewatering is expected to be minimal and will be managed by collection of water by in pit sumps for use within the mining operation
- Crusher feed is provided by a combination of direct tip from the mine (50% of crusher feed) and rehandle from ROM or long term stockpiles using either a front end loader (FEL) only or a FEL and 135 t capacity haul trucks (dependent on haul distance)
- Grade control will be provided by a subcontractor on a 25 metre x 25 metre pattern of RC drill holes, and is campaigned during the mine life.

It was recommended that ore should be blasted on five metre benches using 102 mm and 127 mm diameter holes to optimise fragmentation. Waste is generally drilled and blasted on a 10 metre bench height. All material with an in situ dry density greater than 2.0 t/m3 was classed as hard, otherwise it was categorised as soft. The transported cover material (27 Mt) was assumed to not require any drill and blast. AMC developed drill and blasting costs and equipment fleet requirements based on the parameters provided by the specialist consultant.

Pre-split blasting was assigned to fresh rock walls in Stage 3 and 4. The estimation of pre split to Stage 3 is a conservative decision based on limited geotechnical input and may overstate the pre-split drill metres required. Information gathered during the mining of Stages 1, 2 and 3 will be used to refine the final design for Stage 4. Presplit drilling is designed on a spacing of 1.3 metres, drilled on 10 metre benches with 0.7 metres sub-drill using 102 mm diameter holes.

The primary loading fleet will consist of a maximum of three hydraulic excavators in the 360 tonne class. This class of machine is considered the largest option that could also practically excavate bulk waste and more selective ore zones on 3.5 metres to 4 metres flitches. Smaller machines would result in a higher operating cost and introduce scheduling issues by the need to create additional dig locations. The excavator model was the Hitachi EX3600 unit.

The ore and waste haulage fleet will consist of 180 tonne mechanical drive haul trucks. The truck model was the Cat 789 haul truck which is the largest unit that could direct tip to the primary crusher and is well matched to the proposed excavator. Cat 785 trucks are selected to operate with the 992 FEL loading topsoil and on ROM rehandle. This fleet could supplement mining activities although not scheduled to do so.

The mining schedule has been constrained by setting a maximum vertical advance rate of 60 metres per annum in Stages 1 and 2 and 80 metres in Stages 3 and 4 (due to more bulk waste mining activity in Stages 3 and 4) to allow sufficient time for drill and blast, load and haul, dewatering and grade control. Stages 3 and 4 were split into north and south ends to allow a lag in bench advance between each end of the pit. The maximum vertical lag between benches was set at 20 metres.

The process plant will be designed to operate seven days per week at a nominal treatment rate of 1,100 dtph on Oxide ore, 1,000 dtph on Transitional ore and 937 dtph on Fresh ore at a grinding circuit utilisation rate of 91.3%. Availability is defined as the percentage of total time that the process plant is mechanically and electrically ready to operate while utilisation is defined as the percentage of the total time that the process plant is actually operated.

A single stage primary crush, Semi Autogenous Grinding and Ball Milling with Pebble Crushing (SABC) comminution circuit followed by a conventional gravity and carbon in leach (CIL) process is proposed.

The gravity circuit will consist of four centrifugal concentrators treating a portion of the cyclone underflow. Gravity concentrate will be leached using a vendor supplied intensive leach reactor to yield a pregnant solution from which precious metals will be recovered by electrowinning.

The cyclone underflow launder will have three separate compartments. Two of these compartments will feed the gravity circuit. The number of cyclones servicing each gravity compartment can be adjusted as required. Each of the two gravity feed compartments will feed a dedicated 2.4 metres wide by 6 metres long, horizontal, wet vibrating screen. The screen deck panels will have alternating rows of 2.4 mm by 6 mm and 2.4 mm by 18 mm slots. Screen oversize will return to the ball mill feed. Screen undersize will feed the centrifugal gravity concentrators. Each screen will supply two 1.219 metre (48 inch) diameter centrifugal concentrators. The concentrators will operate in a staggered discharge cycle so that while one unit is flushing the other units are collecting concentrate. The gravity circuit has been designed for a 40-minute collection cycle followed by a standard flushing cycle.

The tailings from the gravity concentrators will return to the ball mill feed. Concentrate from the gravity concentrators will discharge to the intensive leach reactor. The batch leach process will be initiated on a daily basis. The leaching sequence will be controlled by a programmable logic controller (PLC). After leaching, the residue will be returned to the mill discharge hopper by a centrifugal slurry pump and the pregnant solution will be forwarded to electrowinning located in the gold room.

Electrowinning will be carried out in a dedicated 800 mm by 800 mm electrowinning cell fitted with 12 cathodes and 13 anodes. Electrical current will be supplied from a 1,200 A rectifier. The cathodes will be stainless steel and the precious metal precipitate will be removed by washing loaded cathodes in a cathode washing station and filtering the resulting sludge. The filter cake will be dried in an oven and then combined with fluxes and smelted to produce gold doré.

After screening to remove trash, the cyclone overflow from the grinding circuit will be thickened using a 38 metre diameter Hi-rate thickener and then leached with cyanide in a hybrid CIL circuit that consists of a inglestage of leaching and six stages of leaching and adsorption. The total nominal pulp residence time in the hybrid CIL circuit will be 24 hours.

The cyclone overflow from the grinding circuit will gravitate to one of two duty trash screens. The trash screens will be 3.66 metres wide by 6.1 metres long, horizontal, wet vibrating screens. The screen aperture will be 0.8 mm by 18 mm. Trash screen oversize will discharge into the trash bunker and will be periodically removed for disposal by an integrated tool handler. Trash screen undersize will gravitate to the pre-leach thickener. The two trash screens provide a degree of redundancy.

Pre-leach thickener feed will be dosed with flocculant and thickened in the 38 metre diameter Hi-rate thickener to 50% solids (w/w). The thickener underflow (leach feed) will be pumped by one of two centrifugal slurry pumps (14 by 12 inch) with 315 kW drives, arranged in a duty/ standby configuration, to the CIL tanks. Dual transfer lines have been incorporated into the design to permit both pumps to be run for short periods of time in the event of high loads in the pre-leach thickener. Cyanide will be dosed into the suction of the duty thickener underflow pump, and oxygen will be injected into the leach feed line. The thickener overflow will gravitate to the process water pond via a sedimentation pond.

The leaching and adsorption circuit will consist of a 5,000 m³ leaching tank with a nominal pulp residence time for Fresh ore of four hours and six 4,200 m³ CIL tanks with a nominal 20 hour pulp residence time (leaching and CIL). For Oxide ore the residence time will be a total of 20.4 hours, for the Transition it will be 22.7 hours and for Fresh ore it will be 24 hours. The design will include the ability to bypass any tank in the train should this be required.

Cyanide will be stage dosed into the discharge of the leach tank and the first CIL tank as required. Oxygen will also be injected down the agitator shaft of the leach tank and the first two CIL tanks as required. The oxygen manifold could be extended down the leach train to CIL TK-06 but this is not necessary for Gruyere ore.

Each CIL tank will have two 20 m² mechanically wiped, inter-tank screens with 1 mm aperture stainless steel wedge wire to retain carbon. The design carbon concentration will be 9 g/L. Carbon will be advanced through the CIL circuit counter current to the pulp, on a batch basis, by recessed impeller pumps. Loaded carbon from the first stage of the CIL will be pumped to the loaded carbon screen. The loaded carbon screen will be a 1.5 metres wide by 3.6 metreslong, horizontal, wet vibrating screen. Loaded carbon from the loaded carbon screen will gravitate into the acid wash column. The design advance rate for the circuit is 15 t/d. Barren carbon from the kiln (or directly from the elution column) will be returned to the circuit via the barren carbon screen. The barren carbon screen will be a 1.5 metres wide by 3.6 metres long, horizontal, wet vibrating screen.

The pre-leach thickener area will be serviced by one vertical spindle centrifugal sump pump and the leaching and adsorption area will be serviced by two vertical spindle, centrifugal sump pumps for clean-up, with floor slopes appropriately graded to the relevant sumps to facilitate ease of cleaning.

The carbon handling system has been designed for (nominally) six elution cycles per week. However, the design of the acid washing and elution circuit will provide the flexibility to increase the elution frequency should this prove necessary. The flexibility of the design has been enhanced by the addition of a second pregnant solution tank into the elution system which will permit a new second elution cycle to be commenced whilst electrowinning of the pregnant solution from the previous elution cycle is still in progress. A barren solution tank has been incorporated into the design to permit the barren solution from electrowinning to be returned to the leaching and adsorption circuit over an extended period to minimise the disruption to the circuit. A split AARL circuit wasselected in preference to a conventional AARL circuit to reduce the requirement for fresh water.

The elution area and the gold room will each be serviced by a dedicated vertical spindle centrifugal slurry pump for clean-up, with floor slopes appropriately graded to the relevant sumps to facilitate ease of cleaning.

Gold security from production to transportation logistics has been considered in the design of the process plant. The gold room will be a secured building with remote security monitoring, motion detection, seismic monitors and electronic surveillance including closed circuit television (CCTV) recording. Industry standard procedures involving key holder systems, swipe cards and double lock systems for vault and safe access will be employed.