A variety of genetic models have been postulated for the formation of gold deposits in the Pine Creek Geosyncline. Gold and base metal mineralization is commonly associated with granite intrusions and are often been classified as high temperature contact aureole deposits. A secondary host rock control has also been suggested due to the association of gold mineralization with carbonaceous metasedimentary rocks. More recently, authors have argued that gold mineralization is structurally controlled; occurring in brittle ductile structures at the greenschist-amphibole facies boundary and hence has an epigenetic origin.

Maud Creek aligns with the gold-quartz veins, lodes, sheeted veins, stockwork deposit type.

The Maud Creek deposit is hosted within the Tollis Formation of the Finniss River Group. Mineralization is associated with a north-south trending Gold Creek Fault Zone (GCFZ) that forms the contact between mafic tuffs of the Dorothy Volcanics to the east and sedimentary rocks of the Tollis Formation to the west. The GCFZ and primary Maud Creek mineralization dips steeply to the east (65-75°). The GCFZ is characterized by intense deformed and brecciated to catallactic zone up to 10 to 15m width (AngloGold Report, 2000). The GCFZ and Maud Creek mineralization is associated with stockworks and massive quartz veining, silica flooding and brecciation as well as intense graphitic and chloritic alteration (Ahmad & Hollis, 2013). Additional alteration recognised includes silica, carbonate, fuchsite and haematite. The contact zone deposit geometry has been defined as lenticular in shape with a steep plunge (70-80°) to the south-east. This principal mineralized zone extends approximately 250m north-south, and ranges in width from several meters to up to 50m width. The deposit remains open at depth (Snowden Report, 2008).

Mineralization is recognised to extend beyond the contact vein lodes, with dispersion up to 25m into the hanging wall tuff and 5m into the footwall sediments. Outside of the primary vein deposit minor hanging wall micro breccia zones are recognised predominantly occurring proximal to minor faulting parallel to the GCFZ (AngloGold Report, 2000).

Away from the main contact fault zone, gold is recognised within a sub-vertical shear zone which lies proximal to the contact of the Maud Dolerite (Snowden Report, 2008). Mineralization is less continuous in this zone, with the absence of any major vein lode systems evident.

Gold occurs within the deposit as both free gold and as refractory gold in pyrite and arsenopyrite (Snowden Report, 2008). Sulphides can constitute up to 5% of the deposit with pyrite and arsenopyrite and gersdorffite recognised. These sulphides form as disseminations as well as massive intervals containing up to 50% pyrite (Ahmad & Hollis, 2013). Quartz makes up the remaining gangue mineral of the deposit assemblage.

The open pit design work carried out by SRK was based on the preferential allocation of potential mill feed to the proposed underground operations and oxide mineralization being treated as waste. With the proposed treatment of mineralization at the Union Reefs Processing Plant, the Oxide or contained within the pit design was report and included in the proposed mill feed.

Considering oxide as income generating provides the potential to increase the size of the open pit. The impact of the project economics are potentially large, particularly if the constraints of the waste dumps or underground preferences are relaxed.

Should the project proceed SRK recommends that a range of sensitivities/ scenarios are considered at a high level to inform the project team of the implications of these constraints. The sensitivities should include combinations of considering the open pit oxide mineralization as a mill feed source with both the underground preferences retained and removed. Conventional open pit mining techniques are proposed.

The underground is designed using conventional bottom-up open stoping mining methods utilising a combination of cemented rock fill and waste fill. The stopes will be accessed via a decline.

The open pit mining operation proposed is short term with a short operating life of 9 months. The mining equipment considered is small scale. A conventional 80 t class excavator would be ideal as the loading unit and can be matched to 85 t class rigid frame mine trucks for haulage. The length of haul is anticipated to vary from a 2 to 3 truck haul.

The productive mining fleet is anticipated to be supported by a combination of a water truck, grader and bulldozer. This support fleet will maintain the haul road, pit floor, waste dump and drill and blast pattern preparation. Other minor equipment such as IT loaders, support trucks and explosives trucks will support the drill and blast and mobile equipment maintenance activities.

The bench configuration is anticipated to be 5 m drilling benches, mined in 2 x 2.5 m flitches, with the material types being defined by mark-out tape and paint as designated by the site geologists.

Underground mining methods can be categorized into three categories:
• Caving methods.
• Unsupported methods.
• Supported methods.

Given the type of mineralization and the width of the deposit caving methods are not appropriate for the project.

The unsupported and supported mining methods identified were appropriate to the deposit, based on the geotechnical review were:
• Long Hole Open Stoping; and
• Benching

These two methods are suitable given the deposit geometry and predicted ground conditions. Further design work and cost analysis determined that Open Stoping is the preferred mining method for the deposit due to the lower development requirements.

The proposed open stope mining method has sublevel spacing of 25 m and stope strike lengths between 10 m and 30 m depending on the ground conditions and ore continuity. Cemented rock fill will be placed in the stopes as backfill to allow the mining of the next stope in the sequence. Two mining front on each level are possible.

Crushing and Screening Circuit
Run of Mine material is reclaimed from stockpiles by a wheel loader and tipped into a ROM bin. A vibrating feeder transports material from the ROM bin to a C140 Nordberg Single Toggle Jaw Crusher. Any oversize rock is broken at the Jaw Crusher by a hydraulic rock breaker. Jaw Crusher product is transported to a double deck “banana” 3.1 m x 7 m Nordberg Product Screen with a 40 mm aperture top deck and a 14 mm aperture bottom deck by conveyor belts. Tramp steel is removed after the jaw crusher by a fixed magnet to protect the secondary cone crusher and conveyor belts. Screen undersize (minus 14 mm) product is conveyed to a Fine Ore Bin/ Stockpile of 3000 tonnes live capacity, while screen oversize (plus 40 mm product) is transported to a Nordberg Omni-Cone 1560 secondary cone crusher. Secondary crusher product returns to the Product Screen. Intermediate product of minus 40 mm plus 14 mm from the screen is transported to a Nordberg HP500SX tertiary cone crusher. Tertiary crusher product is also returned to the product screen. No changes are envisaged to the existing crushing circuit.

Milling Circuit
Fine crushed ore is reclaimed from the fine ore bin by a slot type belt feeder. Stockpiled fine ore from outside the bin can be reclaimed using a wheel loader which tips to a day (emergency) feeder bin. Both the slot feeder and day feeder discharge to a single mill feed conveyor. The mill feed conveyor discharges to a Mill Feed hopper. The Mill Feed hopper has a split discharge onto two variable speed ball mill feeder conveyors, one feeding the 3 MW, 4.7 metre by 8.2 metre ANI No.1 Ball Mill, the other feeding the 4 MW 5 m x 9.1 m ANI No.2 Ball Mill. Each mill is able to be operated independently and/or in isolation, and each is in closed circuit with separate hydrocyclone clusters.

Mill discharge slurry overflows through a trommel screen into a discharge hopper, from where a portion is pumped to the hydrocyclones, a portion to the centrifugal gravity concentrator and the oversize scats material reports back to the ball mill.

Cyclone underflow is returned to the ball mill feed chute while the cyclone overflow reports to the flotation feed tank.

The study and process design is focussed on the Maud Creek sulphide resource. However, the Maud Creek deposit also includes around 300 kt of oxide and 230 kt of transitional material. The oxide mineralization is generally amenable to direct cyanidation and will be treated through the existing Union Reefs Processing Plant. The transitional mineralization is less amenable to conventional cyanide leaching (is more variable) so alternatively, this transitional material may be treatable by flotation. Flotation recovery could be further improved following a sulphidization pre-treatment process. Controlled potential sulphidization (CPS) technology has been successfully applied on several gold and copper/gold mines, and may be beneficial at Maud Creek.

The Union Reefs Processing Plant currently treats free milling ores from the Cosmo underground mine by conventional crushing and grinding, gravity recovery and CIL processing. It has excess crushing capacity and two underutilised grinding mills at the current throughput rates on a hard underground mineralization. From 1999 to 2003, annual processing rates of 2.8 Mtpa were achieved, but it currently treats only ~750-850 ktpa, so it can handle the additional 500 ktpa of feed from Maud Creek. The addition of a flotation plant enables campaign treatment of sulphides and is flexible in respect to treatment of oxide and transitional material.

At the Union Reefs Processing Plant, the ore is crushed and then milled in closed circuit with hydrocyclones. A separate gravity circuit recovers gravity recoverable gold from the ball mill discharge. The fine milled product from the cyclone overflow is pumped to a new flotation circuit where a high grade low tonnage gold bearing concentrate is produced from the sulphide minerals leaving a gangue non-sulphide residue that can be disposed of, to tailings. The concentrate is dewatered through thickening and filtration before being bagged, stored in shipping containers and transported by road, then ship to customers in China.

Based on a five year LoM, a 500 ktpa plant is recommended as the base case.

The 500 ktpa throughput rate is further supported by the likely underground mining rates. Since the deposit is relatively narrow, a 500 ktpa production rate is a reasonable assumption. A higher mining rate would be challenging given the likely method and geometry of the deposit. The process plant can be simply debottlenecked (for example equipment at this throughput is often oversized) to achieve additional throughput if mining rates and LoM allow it in the future.

The main stages of the process consist of:
• Three stage crushing and product screening circuit (existing)
• Two closed circuit ball mills, each with a hydrocyclone (existing)
• Gravity circuit including Knelson concentrators (existing)
• Flotation circuit (new)
• Concentrate thickening and filtration circuit (new)
• Concentrate storage and loading (new)
• Tailings thickener (existing)
• Utilities (largely existing)
• Reagents (largely existing).

It is expected that there will be some additional overlap as well as some additional modifications required to the existing plant. Confirmation is required to whether a separate process water circuit is required due contamination of the CIL circuit with flotation chemicals and vice versa. For example cyanide is often used as a depressant in pyrite flotation and this could impact gold recovery. It has been allowed for in the capital cost estimate.

Each Ball Mill hydrocyclone cluster is fitted with a direct off-take on the hydrocyclone feed pot, which directs a bleed stream of ball mill discharge to a Nordberg 1.2 metre by 1.5 metre scalping screen to remove coarse scats. Fine screen product reports to one of two automatic discharge 30 inch Knelson concentrators. Rough Knelson concentrate containing coarse gold is automatically discharged to a secured hopper located within the gold room. On a batch basis, the rough Knelson concentrates are transferred from the secured hopper to an Acacia intensive leach reactor (ILR). The concentrates are deslimed using water before caustic cyanide leach solution containing a proprietary “Leach Aid” is recirculated through the concentrate bed to rapidly dissolve the contained coarse gold. At the completion of the leach cycle, the concentrate residue solids are washed and discharged back to the grinding circuit. The high grade pregnant leach liquor containing the gold is then recirculated through an electrowinning cell, with the gold won onto stainless steel mesh cathodes. The high-quality won gold is then direct smelted to saleable bullion product.

Cyclone overflow is first screened to remove trash, then flows to the conditioning tank, where copper sulphate, frother and SIBX (xanthate collector) are added and given time to mix. Copper sulphate and SIBX can also be added upstream to the mill circuit if this is found to be beneficial. Conditioned concentrate overflows to the rougher flotation cells. Rougher concentrate reports directly to the final concentrate thickener. Rougher tails flows to the scavenger cells.

Scavenger cells flotation concentrate is pumped back to the cleaner flotation cells, while scavenger tails is pumped to the tailings thickener. Cleaner concentrate reports to the final concentrate thickener. Cleaner tailings are usually returned to the rougher cells, but may also be returned to the scavenger cells. The next stage of design will also enable tailings to be pumped to the CIL circuit to allow for low recovery Maud Creek transitional mineralization.