From a mineralogical perspective, the Gediktepe deposit is characterised as a massive sulphide skarn, in which processes such as weathering, leaching by the acidic (pH 5.5) groundwater, and oxidation of the sulphides in the upper regions have depleted sulphur and base metals, leaving an oxide zone.

Disseminated pyrite mineralisation, or veins of massive sulphide in the host rock, have sometimes been referred to as transitional sulphide, only because the massive sulphide mineralisation abundance is diminishing to waste grades as a result of phenomena other than post-emplacement alteration of the minerals.

The mineralisation at Gediktepe is associated with greenschist facies schist units, with the main mineralisation host rock unit being chlorite–sericite schist of the Upper Paleozoic. The mineralisation is thought to be developed syn-genetically in sedimentary units elongated along a north-east / south-west trending structure zone and metamorphosed to schist. Greenschist minerals are generally actinolite, chlorite, albite, and epidote.

The tabular mineralised zones, particularly within the sulphide horizon, dip gently to the west. In the north-eastern portion of the deposit, mineralised zones may be shallower dipping. In several locations the overall trend is abruptly terminated, and the tabular mineralised zones are displaced downwards to the north-east, indicating post-mineralisation activity. Progressing south-west to north-east across the deposit, this displacement geometry has been identified three to four times, and these features have been recognised as abrupt breaks or offsets during interpretation of mineralised bodies.

Massive sulphide-type mineralisation occurs as lens shaped units trending north-east / south-west and dipping at approximately 20° to 40° to the north-west. Minerals include pyrite, sphalerite, tetrahedrite, tenantite, chalcopyrite, galena, and magnetite. The units are cut by later north-west / south-east trending post-mineralisation structures within the oxide zone, in which the sulphide mineralisation has been completely leached out, leaving gold and silver relatively intact. Potentially-economic gold–silver–copper–zinc mineralisation is present to varying degrees in the sulphide zone.

The mineralisation at Gediktepe has been is divided by Polimetal into five main types:
- Gossan;
- Massive Pyrite;
- Massive Pyrite–Magnetite;
- Enriched;
- Disseminated Sulphide.

Recent review of interpretations revealed that, in the northern part of the deposit and in the
vicinity of the enriched mineralisation, areas within the sulphide horizon show high gold and
silver and low base-metal (< 0.1% copper and zinc) concentrations.

Gossan (oxide)
The natural acidity is due to the presence of sulphides, particularly pyrite, within the oxide zone, and the sulphide mineralisation has been completely leached out, leaving gold and silver relatively intact. Relic ‘lenses’ of high-gold mineralisation remain in the oxide zone. There is often an increase in gold grade just above the oxide–sulphide contact.

Copper and zinc grades are typically less than 0.10% within the oxide zone but increase to values typically around 1.40% Zn and 0.80% Cu immediately below the oxide horizon. Gold and silver follow the reverse trend, with gold in the range of 3.0 g/t in the oxide zone and often less than 0.7 g/t at the top of the sulphide zone.

The Gediktepe oxide-type mineralisation is characterised by yellow-to-red leached zones of intense iron oxide gossan material. Near surface, is a leached cap, locally containing elevated gold values.

Massive Pyrite
The massive pyrite zone consists of fine to medium-grained pyrite, with massive-to-banded, vuggy textures, and locally sandy textures near structural features. The sphalerite–chalcopyrite–galena and weak covellite are observed as vug fracture fill and replacement mineralisation within a pyrite matrix. Locally, magnetite fragments are observed. The massive pyrite zone hosts high gold and copper mineralisation.

Massive Pyrite–Magnetite
Massive pyrite–magnetite has been distinguished from massive pyrite based on the presence of magnetite. Massive pyrite–magnetite shows the same textures as the massive pyrite. Quartz– magnetite fragments can be seen conformable with the schistosity, or primary bedding structures, within the massive pyrite–magnetite. The massive pyrite–magnetite characteristically shows lower gold–silver–copper–zinc–lead grades than the massive pyrite.

Enriched
The enriched zone consists of mainly chalcocite–covellite within fine to medium-grained pyritic mass. Occurring near or along structural features, the enriched zone is generally intensely fractured. Relative to other sulphide mineralisation zones, the enriched zone contains higher grade gold–silver–copper–zinc mineralisation.

Disseminated Sulphide
A lower grade sulphide mineralisation (gold–silver–copper–zinc–lead) is present within the rich disseminated (pyrite > 10%) chlorite–sericite schist. The total sulphide content in this zone exceeds 8.5%. Bands of 1–50 cm thickness appear parallel to bedding in this host rock below and above the sulphide mineralisation.

Open pit mining is planned to be carried out on 2.5 m flitches using small excavators (3–4 m3 capacity) and trucks. Drilling and blasting will be required. All mining services will be performed by a suitably qualified and experienced Turkish mining contractor. It is currently anticipated that the same mining contractor will provide initial construction services, particularly construction of the tailings storage facility (TSF).

Grade control to determine material types and ore boundaries will be performed based on blasthole sampling and assaying, and under the control of the mine geologists. Feed to the process plants is expected to be a combination of both direct tipping and reclaim from ROM stockpiles to ensure optimal feed to the process plant, particularly for sulphides.

On the eastern side of the pit, if the bench faces are cleanly developed and scaled along the foliation such that the bench face slope is formed by the foliation at an average angle of 40°, rockfall hazards will be mostly removed and a minimum 6 m-wide catch bench would likely provide adequate rockfall protection in most cases. Leaving a 6 m-wide catch bench in the slope at 10 m vertical intervals would result in an inter-ramp slope angle of 29°. In the overburden, it is recommended that 5 m-high production benches with bench faces cut at 45° and a minimum 5.7 m-wide catch bench be developed at 5 m vertical intervals (single benches). This bench configuration results in an inter-ramp slope angle of approximately 25°.

On the west side of the pit, where the structural conditions are more favourable, bench face angles in phase slopes will mostly be limited by rock quality and the mining methods used to develop steep bench faces in highly fractured rock. It is recommended excavating bench faces at 63.5° in this sector. For phase slopes, where trim blasting to a free face is not used and bench faces are formed by cushion blasting in conjunction with standard production blasting, single benching (10 m-high benches) is recommended. Assuming 6.5 m- wide catch benches are left at 10 m vertical intervals results in a 41° inter-ramp slope. For final slopes, where cushion blasting is used in conjunction with trim blasting to a free face and scaling, double benching can be accomplished by stacking two 10 m-high production benches so that an 8.5 m catch bench is left in the slope at 20 m vertical intervals. This results in a 47° inter- ramp slope.

In the overburden, the geotechnical drillhole data indicates that the depth of highly weathered rock conditions on the west side of the pit is less than approximately 10 m. It is recommended to excavate the first bench at a bench face angle of 45° and leave a 6 mwide catch bench on top of sound bedrock at the crest of the pit.

It was assumed that effective depressurisation of all pit slopes will be feasible, and that groundwater will not be a control on stability. Achieving this may require that drainage enhancements such as wells and horizontal drains be installed in less-permeable geotechnical units and where locally perched groundwater occurs in pit slopes.

The configuration of the recommended 47° west wall inter-ramp slope (20 m bench stack height with an 8.5 m catch berm) was reviewed in the context of Turkish practices and regulations regarding maximum bench stack heights. After discussion with relevant parties, and technical assessment of the effectiveness of a narrower catch berm, an alternative west wall inter-ramp slope configuration of a 15 m-high stack height with a 6.5 m catch berm was adopted for both intermediate pit stages and for the ultimate pit. This revised configuration achieves the initial 47° inter-ramp slope target.

After review of the initial pit eastern (footwall) inter-ramp slope design, slight flattening (2° to 3°) was recommended to achieve acceptable factors of safety (FOS) for the rock types intercepted. Additionally, the southern portion of the eastern pit slope incorporates a permanent creek diversion. Review of the risks around this critical infrastructure recommended that:

• The berm the diversion is located on should be wider, and

• The overall local pit slope should be reduced in order to achieve a FOS of 1.5 (1.2 for standard slope design) to ensure longevity of this critical infrastructure.

The project will therefore be installed and commissioned in two stages:

• Stage 1 oxide ore – comprising a two-year period for processing gold and silver ore, which will be treated in a single stage semi-autogenous grinding (SAG) mill circuit.

• Stage 2 sulphide ore – the oxide processing plant will be expanded to process copper and zinc- bearing ore by flotation. A 5.5 MW secondary grinding ball mill will be added to the grinding circuit.

The crushing circuit, incorporating ROM ore handling, crushing, and storage, has been designed on the basis of 6,570 operating hours per annum, or 75% operational availability. This has been derived from first principles, taking into consideration planned and unplanned downtime events.

Oxide ore treatment
The ROM ore will be loaded into the ROM bin by a front-end loader or direct tipped by 25 t dump trucks. A 750 mm static grizzly will be fitted to the ROM bin to protect it, and all downstream processing equipment from oversize material. The static grizzly will be inclined and hinged to allow easy removal of oversize material or in case of a blockage or hang-up. Mining will be required to supply ore at a P100 of 750 mm to minimise grizzly cleaning requirements. Any oversize ore will be scalped from the screen and stockpiled adjacent to the ROM bin.

The ROM ore will be drawn from the ROM bin at a controlled rate by a variable speed apron feeder and discharged onto the vibrating grizzly feeder equipped with 90mm bar spacing’s. Oversize material from the vibrating grizzly feeder will discharge into the primary crusher. The undersize material from the vibrating grizzly feeder will gravitate onto the primary crusher discharge conveyor. The primary crusher is a 160 kW Metso C-120 single toggle jaw crusher single with a 1,200 mm by 870 mm gape and will operate with a closed side setting (CSS) of 80 mm to give a product with a P80 < 90 mm. The crusher product will discharge onto the primary crusher discharge conveyor and be transferred to the crushed ore bin feed conveyor that then discharges into the crushed ore surge bin. A dust collector will be positioned at the end of the primary crusher discharge conveyor for dust control and management.

The speed of the variable speed apron feeder will be controlled by a PID controller to maintain an overall circuit throughput rate as measured by the weightometer on the crushed ore bin feed conveyor. Process spillage in the crushing area will be pumped to the primary mill discharge hopper by a sump pump.

The milling circuit will consist of a single stage SAG mill in closed circuit with a hydrocyclone cluster. Process water and lime slurry will be added to the mill feed chute to control the mill discharge density and slurry pH respectively.

The primary SAG mill will be of the grate discharge type with a diameter of 6.4 m (inside
shell) and an effective grinding length of 4.23 m. The mill will operate with a nominal ball load of 6% by volume and an operating critical speed ranging from 60% to 78%. The SAG mill will be powered by a 3,000 kW motor with variable speed capability. The mill power draw and product size will be controlled by the periodic addition of grinding media. The slurry will be separated from oversize pebbles and undersized mill balls, with the washed oversize material exiting the trommel onto the pebble transfer conveyor and recycled back to the SAG mill feed chute via the pebble recycle conveyor. Trommel screen undersize material will discharge into the primary mill discharge hopper.

The combined slurry will be pumped from the primary mill discharge hopper to the primary mill cyclone cluster by the variable speed primary mill cyclone feed pumps. The cyclone underflow stream will be returned to the primary SAG mill for further grinding, while the cyclone overflow (target P80 size of 125 µm) will be directed to the tank leach trash screen to prevent the introduction of oversize material and debris into the leaching and adsorption circuit. Oversize material from the trash screen will report to a trash bin, whilst the trash screen underflow will report to the leach feed distribution box. Process spillage in the milling area will be controlled by two sump pumps.

Sulfide ore treatment
A two stage, SAG and ball mill grinding circuit has been proposed to reduce the crushed material to a P80 of 38 µm for feed to flotation. The ball mill will operate in closed circuit with a cluster of 250 mm diameter cyclones. The cyclone overflow will report to a horizontal, vibrating trash screen 2.4 m wide by 4.8 m long. The trash screen will be fitted with polyurethane screen panels having an aperture of 1.0 mm. Cyclone underflow will be directed to the ball mill feed chute.

The specific grinding energy required for the concentrate regrind duties was determined by generating signature plots in tests conducted to grind rougher concentrate samples using bead mills. A regrind P80 size of 15 µm was used in the flotation testwork for copper rougher concentrate and a P80 size of 20 µm for zinc rougher concentrate.

The project will therefore be installed and commissioned in two stages:

• Stage 1 oxide ore – comprising a two-year period for processing gold and silver ore, which will be treated in a single stage semi-autogenous grinding (SAG) mill circuit, followed by sodium cyanide leaching, carbon-in-pulp (CIP) and elution and electrowinning techniques to recover the gold and silver;

• Stage 2 sulphide ore – the oxide processing plant will be expanded to process copper and zinc- bearing ore by flotation.

During the period when both oxide and sulphide ore are available for treatment, parcels of ore will be campaign treated. This will coincide with the ramp-up period for the sulphide ore and will allow development of operating knowledge for the treatment of the sulphides allowing time to analyse data during oxide campaigns. Processing of sulphide ore will be undertaken whenever two weeks of ore supply is available (50–85 kt during ramp-up), which will generally a ........