The Skouries deposit is centered on a small (less than 200 m in diameter), pencil-porphyry stock that intruded schist and gneiss of the Vertiskos unit. The mineralized porphyry intrusion plunges steeply to the south southwest and obliquely crosscuts the moderate to steeply northeast dipping limb of a district-scale F2 antiform. Ore has been tested to a depth of 920 m from surface. Surface exposures and drill data indicate that the porphyry stock has a subtle northeast elongate geometry. The porphyry is characterized by at least four intrusive phases that are of probable quartz monzonite to syenite composition (Kroll et al., 2002; Frei, 1995), but contain an intense potassic alteration and related stockwork veining that overprints the original protolith. Potassic alteration and copper-gold mineralization also extends into the country rock; approximately two thirds of the measured and indicated tonnes are hosted outside the porphyry with about a 50:50 split in gold-equivalent ounces. The potassic alteration is characterized by K-feldspar overgrowths on plagioclase, secondary biotite replacement of igneous hornblende and biotite, and a fine-grained groundmass of K-feldspar quartz with disseminated magnetite. Four main stages of veining are recognized:
• Early stage of intense quartz-magnetite stockwork.
• Quartz-magnetite veinlets with chalcopyrite ± bornite.
• Quartz-biotite-chalcopyrite ± bornite apatite-magnetite veinlets.
• Localized, late stage set of pyrite ± chalcopyrite-calcite-quartz veins.

Skouries is typical of a gold-copper pencil porphyry; mineralization occurs in stockwork veins, veinlets and disseminated styles typical of a porphyry, and has a subvertical, pipe-like shape. The multi-phase monzonite to syenite porphyries intruded into metamorphic basement rocks. Both igneous and metamorphic rocks contain high temperature potassic alteration (K-feldspar-biotite) and stockwork quartz magnetite-chalcopyrite-bornite veins. The potassic zone in the surrounding country rock is surrounded by a high temperature inner propylitic alteration characterized by amphibole. The deposit however lacks extensive phyllic or argillic-advanced argillic zones typical of many porphyry systems. This may in part reflect a deeper level of erosion and the focused nature of the magmatic-hydrothermal system.

The Skouries Project is designed as a two phase mining operation. Phase 1 consists of a combined open pit and underground mine, operating over 10 years. Phase 2 consists of mining from the underground mine for further 13 years. The total life of mine (LOM) is 23 years.

Phase 1 total mill feed is 8.0 Mtpa, consisting of a nominal 5.5 Mtpa from the open pit mine combined with 2.5 Mtpa from the underground mine. At the start of the mine life, during the initial two year underground mine ramp up period, the open pit feed rate is 6.4 Mtpa in order to maintain the 8.0Mtpa mill feed. During Phase 1 approximately 6.5 Mt of oxide ore is stockpiled to be rehandled for mill feed during Phase 2. Phase 1 is complete at the end of the open pit mine life in Year 10. Phase 2 mine production, from Year 11 to the end of the LOM, is provided from the underground mine. Phase 2 mine development begins in Year 4 in order to allow a seamless ramp up from the Phase 1 production of 2.4 Mtpa. During the first three years of Phase 2 the mill feed rate of 8.0 Mtpa is maintained by reclaiming oxide ore stockpiled during Phase 1, at a rate of approximately 1.6 Mtpa. From Year 14 on, Phase 2 mill feed is maintained at a nominal feed rate of 6.2 Mtpa, solely from underground mine production.

Open pit mining will be by conventional truck shovel operation, with an ore production rate of approximately 5.5 Mtpa, at a waste to ore stripping ratio of approximately 0.88:1. The mining sequence will consist of drilling, blasting, loading and hauling of ore and waste materials for processing and waste disposal. Based on the modelled rock types, approximately 17% of the mined material is amenable to free digging, this material will not be blasted.

Direct feed ore from the open pit will be hauled to the Skouries processing plant. A portion of oxide ores will be hauled directly to the plant, and an additional portion will be hauled to the oxide ore stockpile (OOS) where it will be re-handled during the Phase 2 of the Project. Waste material will be hauled to a transfer point adjacent to the OOS area where it will be re-handled by a fleet of smaller contractor trucks and placed in one of the material management structures within the IWMF. The structures internal to the IWMF are the J5, OOS pad, KL embankment, upstream (US) waste zone and the capping stockpile pad.

The Skouries orebody that extends below the bottom of the open pit is amenable to a bulk underground mining methods and has been evaluated under several different design approaches since the late 1990’s including block caving, sub-level caving and sub-level open stoping. Sub-level open stoping (SLOS) has been confirmed as the most appropriate underground mining method for a number of reasons including: the geo-technical stability of the final reclaimed land after closure of the Project, the minimization of land-take needed for the surface tailings, and the ability to backfill the depleted open pit.

The majority of the stoping is considered to take place in reasonable quality rock mass. The geotechnical analysis has indicated that, for stoping in the porphry, a 60 m sublevel interval (65 m stope height) can largely be viable without significantly compromising stope wall stability if the length of the stope does not exceed 30 m. For primary stoping in schists, preliminary stope lengths have been set at 20 m. Stope widths have been limited to 15 m in consideration of stability of the stope back. Thus, the standard stope dimensions were set to 65 m high x 30 m long x 15 m wide in porphyry, and 65 m high x 20 m long x 15 m wide for primary stope design in schist material.

The parameters shown are for the evaluation of the stope sidewall height, which determines the sublevel interval. The hydraulic radius of 10.3 is for a stope 65 m high (floor to floor) and with a distance along strike of 30 m. Of the stopes that will be extracted in the schist, only half of these excavations will expose schist in the stope sidewalls as secondary stopes will expose the paste backfill within the primaries. The potential impact of exposing schist sidewalls is an increase in dilution.

Anticipated ground conditions have played a major role in sizing the stopes. All stopes are designed at 60 m high by 15 m wide, in a primary-secondary transverse arrangement. Primary stopes will be developed and then backfilled in segments of 20 m or 30 m in length depending on the rock type: 30 m long in porphyry and 20 m long in schist material due to expected less favourable ground conditions. Secondary stopes will lag the primary stopes by at least 60 m transversely and will be mined and then filled in 30 m long segments regardless of the rock type. Unlike the test stopes, production stopes will be backfilled with paste fill. The stoping methodology is the same for both Phase 1 and Phase 2.

PRIMARY CRUSHING AND STOCKPILE
The ore is delivered by 90 t haul trucks from the open pit and 50 t haul trucks from the underground, to the primary crushing station dump pocket. The ore is then crushed by the gyratory primary crusher. Then, the crushed ore is discharged via the primary crusher discharge feeder.

The transportation of the crushed ore from the primary crusher discharge to the covered conical coarse ore stockpile is implemented using a belt conveyor system which is comprised of the primary crusher sacrificial conveyor and the primary crusher discharge conveyor. An overband magnet at the discharge of the sacrificial conveyor removes scrap metal before entering the ore stockpile.

An automatic dust suppression system will be installed at the crusher feeding point and at the crusher discharge feeder to prevent dust emission from the ore unloading and crushing operations.

The belt conveyors are equipped with blockage detectors for discharge chutes (alarm and trip), zero speed switch in case of belt slippage (trip) and belt alignment switches for belt position (alarm and trip).

The ore stockpile is covered and has a live storage capacity of 24,000 t, equivalent to one day production. The stockpile total storage capacity will be approximately 80,000 t, equivalent to slightly over three days of production.

The ore is extracted from the ore stockpile using three variable speed apron feeders, which are of 1,215 t/h total capacity and each one is driven by a 75 kW motor. The feeders are located beneath the stockpile in a tunnel.

The feeders discharge onto a belt conveyor system which is comprised of the ore reclaim sacrificial conveyor and the SAG mill feed conveyor which reports to the SAG mill feed chute.

The feeders have been designed and located to maximize live storage capacity of the stockpile. The feeding rate to the mill is controlled by the control system in the plant.

An automatic dust suppression system will be installed at the apron feeders to prevent dust emission. The spray water is collected in the inclined reclaim tunnel floor.

The belt conveyors are equipped with blockage detector for discharge chutes (alarm and trip), zero speed switch in case of belt slippage (trip) and four belt alignment switches for belt position (alarm and trip) and will be covered and guarded.

GRINDING CIRCUIT AND PEBBLE CRUSHER
The primary grinding circuit is designed to reduce the feed ore with a particle size of 80% passing 150 mm to a product with a particle size 80% passing 120 µm. The size reduction is achieved by a two-stage wet grinding circuit comprising a SAG mill driven by variable speed motors, a ball mill (with a fixed speed motor) and a pebble crushing circuit. A gravity concentration circuit is integrated in the primary grinding circuit.

The transfer size of the material between the SAG mill and the ball mill is in the range of 80% passing between 2.0 and 3.6 mm.

The SAG mill has diameter of 9.75 m, effective grinding length 4.57 m and is driven by two 4.8 MW motors with VFD including the auxiliary lubrication circuits. The grinding media is balls of diameter 125 mm made from high quality forged steel. The SAG mill liners are Cr-Mo alloy cast steel.

Two shorthead type cone crushers will be installed for the crushing of the oversize pebbles (+12 mm) produced by the SAG mill. Total crushing capacity is 254 t/h and each one will be driven by a 450 kW motor.

The primary ball mill has diameter 7.01 m, effective grinding length 9.75 m and is driven by two 4.8 MW motors. The grinding media is balls of diameter 60 mm. The ball mill will be lined with rubber liners and lifters.

Regrinding Circuit
The concentrate regrinding is carried out by the regrind ball mill. The regrind ball mill has diameter 4.60 m and effective grinding length 7.00 m. The mill is driven by a 2.25 MW motor and is equipped with lubrication, drive protection and cooling systems. The regrind mill has rubber lining and the regrinding media is forged steel ball typically with 25 mm diameter.

The regrind ball mill operates in closed circuit with a cluster of 14 hydrocyclones, 10 duty and 4 standby. The operation of the regrind circuit will be controlled by the central DCS control system of the plant.

The regrinding circuit is fed with the concentrates of rougher and cleaner scavenger flotation as well as the regrind gravity concentration tailings and additional cleaner concentrate. These are directed to the hydrocyclones cluster. The regrind cyclone cluster overflow will flow to the 1st cleaning circuit.

The regrind cyclone underflow feeds the regrind ball mill. A portion of cyclone underflow will be fed to a gravity concentration circuit, which is different from the one in the primary grinding circuit. The regrind gold gravity concentration tailings flow back to the regrind ball mill.

For the first ten years of operation, the ore will be extracted from the open pit mine as well as from the underground mine for a total mill feed tonnage of 8.0 Mtpa, from the eleventh year of operation until the depletion of reserves the plant will process ore extracted from the underground mine at a reduced tonnage of 6.2 Mtpa. During years 11 to 15 oxide ore is being rehandled to maintain mill feed at 8.0 Mtpa during this period.

The plant will process the copper/gold ore with a LOM average grade of 0.49% copper and 0.74 g/t gold. Expected LOM average recoveries are 87.9% for copper and 82.4% for gold, respectively. The mill will produce a doré which contains approximately 80% gold and 10% silver and 10% copper and a concentrate that contains an average of 26% copper and 27 g/t gold. Metallurgical tests have shown that the ore contains a small amount of palladium, which will be collected into the copper/gold concentrate during flotation.

The process plant design provides for a nominal 8.0 Mtpa of ore throughput.

The unit operations comprise of:
• Primary crushing and ore stockpile.
• Grinding and pebble crusher.
• Flotation and regrind.
• Gravity concentration and recovery.
• Concentrate and tailings thickening.
• Concentrate filtering storage and loadout.
• Tailings filtration.
• Reagents and services.

The flotation is carried out in six stages
• Rougher.
• 1st cleaning.
• Cleaner scavenger.
• 2nd cleaning.
• 3rd cleaning.
• Additional cleaning, in order to produce a clean copper/gold concentrate.

The process equipment is of the latest proven technology and completely automated. The flow of concentrate from one area to the other is implemented using gravity flow, where possible, in order to minimize the pumping and consequently the energy consumption.

The concentrate regrinding is carried out by the regrind ball mill. The regrinding circuit is fed with the concentrates of rougher and cleaner scavenger flotation as well as the regrind gravity concentration tailings and additional cleaner concentrate. These are directed to the hydrocyclones cluster. The regrind cyclone cluster overflow will flow to the 1st cleaning circuit. The regrind cyclone underflow feeds the regrind ball mill. A portion of cyclone underflow will be fed to a gravity concentration circuit, which is different from the one in the primary grinding circuit. The regrind gold gravity concentration tailings flow back to the regrind ball mill.

The gold gravity concentration circuit has been designed in a vertical arrangement to maximize the gravity flow among successive processing stages.

The gold gravity concentration circuit is comprised of centrifugal Knelson gravity concentrators. The principle of operation of the gravity concentrators is based on the difference between specific gravities of gold and other accompanying minerals, through which separation is accomplished. This circuit is comprised of three stages.

A portion of slurry from cyclone underflow in the primary grinding circuit passes through two screens (with 2 mm opening) at the front end of the circuit and screen undersize slurry is equally distributed to two centrifugal gravity concentrators. The gravity tailings are recycled back to the SAG and ball mill sump. The produced concentrate from each of these two concentrators flows by gravity into a secondary gravity concentration circuit for upgrade.

The primary gold gravity concentrate slurry is collected in a settling tank for the coarse gold and a second settling tank for the fine gold, where sufficient time for solids settling is provided. Then the concentrate is fed to a shaking table where the final upgrade is carried out. A belt magnetic separator is installed above the shaking table for removal of any tramp iron and other magnetic particles.

The tailings from the above shaking table are collected and directed to a small semi automatic centrifugal gravity concentrator. This will act as a final gold scavenging stage prior to the tailings being discharged into the combined secondary gravity circuit tailings sump. From this sump, the tailings are pumped to the SAG and ball mill sump. The concentrate produced from this process is directed to a settling tank and then to a second shaking table for final upgrade. Solids settle out while the overflow ends up to the secondary gravity separation tailings sump.

The final gravity concentrate from both shaking tables is collected into a sealed concentrate storage box.

The fine material produced from the regrinding circuit passes through the screen (with 2 mm opening) installed at the beginning of the circuit and feeds three centrifugal gravity concentrators in the first step and two gravity centrifugal concentrators in the second step. The regrind gravity tailings are directed back to the regrinding circuit. The concentrate is directed to the secondary gold gravity separation circuit for upgrade.

The gold storage cabinet is transferred by an electric overhead beam hoist into the adjacent gold room once a day.

The gold room is constructed of reinforced concrete and is provided with dedicated flue gas exhaust and ventilation systems. In this room, the doré gold bars will be produced, stored and loaded in specially designed vehicles.

The final gravity concentrate delivered to the gold room is manually distributed in stainless steel trays, which are loaded into a drying/calcining oven. The dried and calcined material is temporarily stored on cooling racks within the gold room. The flue gases generated from the oven are collected and directed through a wet scrubber before being vented into atmosphere.

Approximately once per week the cooled dry product is blended with fluxes and transferred into an induction smelting furnace. The resultant smelt is poured over a series of 10 kg cascade doré gold molds which produces semi pure alloy of approximately 80% gold. The smelting slag is crushed manually within the Gold Room and tipped onto the SAG mill feed conveyor for re-processing within the concentrator.