Halilaga is classified as a copper-gold porphyry system. Advanced argillic alteration and gold mineralization at Kunk Hill and Pirentepe are classified as high sulphidation epithermal mineralization. Copper mineralization in the Bakirlik Tepe area is classified as proximal copper skarn. All three types are related to magmatic-hydrothermal activity associated with intrusion of the Kestane porphyry stock and other intrusions in the area.

The geology of Halilaga is characterized by various lithological associations made up of: (1) Paleozoic and early Mesozoic basement metamorphic rocks; (2) Permian and Mesozoic sedimentary and ophiolitic rocks; (3) Tertiary volcanic and intrusive rocks; and (4) Neogene sedimentary rocks. Older rocks are affected by several collisional orogenic events. Tertiary rocks record mainly brittle extensional and transtensional deformation.

The property is located in a district with significant potential for both porphyry copper-gold and high-sulphidation-style gold deposits. Investigation of geological relationships in the field along with geochemical and geochronological data suggest that magmatic activity began with the intrusion of granitoids, coeval with an early phase of volcanic activity, and was superseded by a second phase of volcanism. The mineralized intrusives are thought to be part of a suite of 26 to 38 million year old (Ma) quartz monzonites and granodiorites.

The Halilaga district is mainly underlain by Oligo-Miocene volcanic and-sedimentary rocks overlying a basement of schist and carbonate rocks which outcrop to the southeast of the Bakirlik area. Granodioritic stocks intrude the basement rocks, giving rise to metasomatism and skarnification. The Kestane porphyry stock, the main host rock to the Halilaga porphyry system, has a pronounced hornfelsed halo.

The Halilaga Project is centered on a large mineralized system containing porphyry copper-gold, skarn and high-sulphidation gold mineralization with related alteration assemblages extending over an area of more than 4 km by 2 km. The Kestane porphyry stock is characterized by potassic alteration (K-feldspar-biotite-magnetite) overprinted by phyllic (quartz-sericite-pyrite) alteration. Most quartz veins are B-type that average 5% of the rock by volume. A-type veinlets are rare but locally present. Chalcopyrite and pyrite are the dominant sulphide minerals. The highest gold and copper grades in drill core are associated with early biotite + magnetite + chalcopyrite associated with dense array of A- and B-type quartz veins, overprinted by phyllic alteration. In places the mineralization has been subject to near surface oxidation and leaching to form supergene chalcocite.

At higher elevations to the south of Kestane, the Kunk-Kumlugedik hilltops are characterized by advanced argillic alteration (quartz-alunite-dickite) surrounded by zones of argillic and propylitic alteration. Skarn-related alteration (magnetite-epidote) is located to the southeast of Kestane in the Bakirlik and Bostanlikbasi areas.

The geology of the Halilaga area is affected by post mineral faults of the North Anatolian Fault System characterized by ENE-WSW strike-slip faults, with subsidiary WNW-ESE striking faults.

250 mm diameter blast hole drills are planned to perform the bulk of the production drilling in the mine (both mineralized and waste rock). The hydraulic drill with a 115 mm diameter bit would be used for secondary blasting requirements and may be used on the tighter spaced patterns required for pit development blasts. The main loading and haulage fleet is planned to consist of 40 t haul trucks, loaded primarily with the diesel powered 7 m3 front shovels or the 7 m3 wheel loader, depending on pit conditions.

As pit conditions dictate, the D9-class dozers are planned to rip and push material to the excavators and maintaining the waste dump.

The additional equipment is planned to be used to maintain and build access roads and to meet various site facility requirements, including stockpile maintenance and further exploration development.

The Halilaga process plant and associated service facilities are designed to process 25,000 t/d of ROM material, to produce copper concentrate, gold doré and tailings. The proposed process includes crushing and grinding of the ROM material, rougher and cleaner flotation, regrinding, cyanide leaching, cyanide detoxification, gold room and dewatering of copper sulphides and is amenable to the mineralization type at Halilaga. The flotation and cyanide destruction tailings would be thickened before placement in the TSF.

Primary Crushing
The gyratory crusher is proposed as a permanent installation that would take ROM material and produce a product of 80% passing 150 mm. Haul trucks are planned to supply ROM material to the primary crusher dump pocket, where they would unload into one of two dump aprons. The dump pocket would have a hydraulic rock breaker to reduce any oversize rocks that may clog the crusher feed. The gyratory crusher would process the ROM mill feed rock at a rate of 1,390 t/h. The crushed material would discharge from the underside of the crusher hopper onto the sacrificial primary crusher discharge belt conveyor. The material would then feed into the coarse mill feed stockpile belt conveyor which would elevate the material to deposit onto the coarse mill feed stockpile.

A dust collection and suppression system would be installed to control fugitive dust generated at the crusher, material transfer points and other relative operations.

Stockpile and Reclaim
The coarse mill feed stockpile would hold one day of live storage of the crushed material, or 25,000 t. Two apron feeders would reclaim the material with one operating, and one on standby, during normal operation. The apron feeders would meter the flow onto the SAG mill feed conveyor equipped with a belt scale, at a controlled rate.

Primary Grinding and Classification
The primary grinding circuit is proposed to incorporate a SAG mill and one ball mill. The process rate would be 1,132 t/h (25,000 t/d).
The SAG mill would be fed at a controlled rate by the reclaim apron feeders under the coarse mill feed stockpile. Lime would be added to the SAG mill feed belt conveyor to raise the pH of the slurry, which would aid copper flotation. A SAG mill ball bin and feeder would feed fresh grinding media onto the SAG mill feed belt conveyor to maintain the grinding charge.
The SAG mill discharge containing 70% solids by weight would pass over a screen to remove over- size pebbles. The pebbles would be conveyed outside the building to a discharge pile for manual re-entry into the process, storage for future processing, or disposal, depending on the mineralized rock characteristics.
The SAG mill screen underflow would combine with ball mill discharge into one common pump box. The ball mill would be in closed circuit with cyclone cluster and slurry underflow stream. The overflow slurry stream would feed the copper rougher/scavenger flotation circuit. The cyclone overflow particle size is proposed to be P80 150µm and contain approximately 32% solids by weight. Cyclone underflow to the ball mills would be approximately 72% solids by weight, and the circulating load would be approximately 250% of new mill feed. Ball charge systems would add grinding media as required for maintaining grinding charge.

Flotation and Regrind
Flotation feed is planned to be conditioned in a rougher conditioner tank where flotation reagents would be added. Frother would be added to the conditioner tanks overflow pipes feeding the first rougher flotation cell.
The rougher conditioner tank would overflow to the rougher flotation cells connected in series. Five 300 m3 forced air tank cells have been selected to provide the required residence time for the roughing flotation duty. The cells would be arranged with a step in level between each pair of cells.
Concentrate from the rougher cells would flow by gravity to a rougher concentrate launder and pumped to the regrind circuit.
Vertical spindle sump pumps would be provided in the rougher flotation area to facilitate clean up.
Concentrate from the rougher cells would be pumped to a regrind mill circuit to achieve the fine regrind size of P80 20 µm. Regrinding will be achieved in a 18’ dia x 37’ EGL – 7,500 hp ball mill (type of the regrind mill will be finalized after performing regrind test).
Reground concentrate is designed to be mixed with flotation reagents: collector and frother, before being pumped to the cleaner 1 flotation cells. Concentrate from the cleaner 1 flotation cells would flow via gravity to the cleaner 2 feed pump. Tailings from the cleaner 1 flotation cells are planned to flow by gravity to the cleaner scavenger flotation cells.
Concentrate from the cleaner scavenger cells would flow by gravity to the cleaner scavenger concentrate launder and would be pumped to the first cleaner feed box. Tailings from the cleaner scavenger cells are designed to flow via gravity to a leach thickener pumpbox.

Thickening and Concentrate Filtration
Copper concentrate from cleaner 3 would be pumped to a high-rate concentrate thickener via the concentrate thickener feed box. Thickener overflow would be pumped to the process water tank for storage and re-use in the circuit, while thickener underflow would be pumped to the copper concentrate filter feed tank. The copper concentrate filter feed tank would have approximately 12-hours storage capacity.
Thickened copper concentrate slurry would be delivered to the copper concentrate filter. The filter press would reduce the moisture content of the concentrate prior to transport. Filter cake would discharge through the floor of the filter building into a concrete area underneath the press. Filter cake would be removed from the bunker by front-end loader (“FEL”) and stored in the covered concentrate storage shed.
The concentrate filter press and concentrate storage stockpile are planned to be housed in a building, which would be fully sheeted on all sides for protection from wind and rain. The concentrate would be stockpiled in the shed by FEL to provide covered on-site storage capacity. Concentrate would be reclaimed by FEL from the stockpile and loaded into concentrate trucks for transport off-site.
The moisture limit has been assumed to be 12% w/w in the filter cake for the filter duty.

Gold Leaching and Recovery Circuit
The first copper cleaner tailings would be pumped to a dedicated thickener and thickened to about 50% solids, giving a density at which the carbon should have neutral buoyancy. Approximately 20% of the feed gold would be in the leach thickener underflow. Leaching is planned to take place in six carbon-inleach (CIL) tanks. All tanks would be arranged in a pattern to minimize the footprint, and would sit on a series of descending steps. Any one tank could be taken off line for maintenance. Average residence time for the CIL circuit would be 12-hours.
At the end of the CIL circuit, the pulp would enter into a two tank cyanide destruct module with 2-hours residence time using SO2 and copper sulfate as the detox reagents to treat leaching circuit tails prior to discharge to the tailings storage facility.
Loaded carbon from the leach circuit would be sent to a carbon plant where gold would be recovered. Water overflowing the leach feed thickener would be pumped to process water tank.
Loaded carbon would be sent to an acid wash vessel and treated by hydrochloric acid solution to remove scale and other impurities. After neutralization the carbon would be pumped to a Zadra Strip vessel. Gold would be stripped from the carbon by circulating a hot caustic solution through the vessel at about 135°C and a pressure of 345-480 kPa. The strip solution would be heated using a combination of plate and frame heat exchangers and an electric hot water heater. After reaching stripping temperature, the solution would flow upward through the strip vessel.