The Kumtor Mine area is dominated by several major thrust sheets with an inverted age relationship (Figure 7-3). The dominant structural direction is northeast southwest (D 4), with moderate dips to the southeast. Each thrust sheet contains older rocks than the sheet it structurally overlies. Four major structural sheets (zones), stacked upon one another, and bounded by long-lived faults that were active repeatedly, have been identified, as follows:

* Zone 0 consists of Cambro-Ordovician limestone and phyllite, thrust over Tertiary sediments of possible continental derivation that in turn rest on Carboniferous clastic sediments with apparent profound unconformity;
* Zone 1 constitutes the KFZ, whose upper limit is the Upper Kumtor Fault. The KFZ is generally a dark-grey to black, graphitic gouge zone. The KFZ strikes northeasterly, dips to the southeast at moderate angles and has a width of up to several hundred metres. The adjacent rocks in its hanging wall are strongly affected by shearing and faulting for a distance of up to several hundred metres;
* Zone 2 includes the mineralization which is hosted by Vendian meta-sediments, grey carbonaceous quartz-sericite-chlorite schists or phyllites that are strongly folded and schistose, with a large proportion of faulted and sheared rocks. Zone 2 is delimited in the footwall by the Upper Kumtor Fault and in the hanging wall by the Lysii Fault. It appears that the mineralizing event, itself multi-phase, has healed some of the earlier brittle features within Zone 2;
* Zone 3 consists of phyllites, also of Vendian age, that show several phases of folding. Zone 3 is sub-divided into three units based on the orientation of the foliation. The dip of the schistosity is shallow to steep to the southeast in the Zone 3A and 3C rocks, and steep to the northwest in the Zone 3B rocks. The subsequent brittle deformation is less strongly developed compared to Zones 1 and 2. Zone 3 is sub-divided into three units based on the orientation of the foliation. Zone 3 is important for the pit slope stability discussed in Section 16 because of the development of the D3 and the D4 back-thrust faults within the zone.

It is important to note that most fault structures in the area are persistent, with thick gouge or tectonic breccia and are therefore potential failure surfaces. This is attributed to D 4 reactivation of pre-existing faults.

The main structures of the Central Deposit are also present in the Southwest and Sarytor Deposits. The main thrust faults at Sarytor strike E-W, and the faults and S 1 schistosity along with the mineralized zones have a shallow southerly dip. The structures are truncated in the west, with Zone 0 juxtaposed against a steep NNE SSW trending fault. This fault is strike-slip in nature indicating a D 3 origin, and is likely the over-steepened continuation of the Lower Kumtor Fault, which separates Zones 0 and 1 in the Central Pit area. The E-W orientation of thrust faults with a southerly dip is a function of D 3 deformation, with limited overprinting by the D 4 structural event.

Structures start to swing to a NE-SW trend as they continue eastward into the Southwest Deposit area, where there is a strong D 4 structural overprint. This D 4 overprint can be traced northward into the Central Pit and thus the structural make up of the Southwest Deposit is more like that of the Central Pit than the Sarytor Deposit. These observations indicate that the D 3 structures were rotated into a NE-SW trend by the D 4 structural event.

Gold mineralization of economic importance occurs where the Vendian sediments have been hydrothermally altered and mineralized, an event that may have taken place in Permian time as discussed above, based on structural considerations (Ivanov et al., 2000) and as reported by Mao et al. (2004). Gold mineralization is developed over a strike distance of more than twelve kilometres. The Central Deposit is the most important accumulation identified to date and has considerable dimensions with a strike length of 2.4 kilometres, a vertical extent of one kilometre, and a width of up to 300 metres. Other known occurrences along the mineralized trend are the Southwest Deposit and Sarytor Deposit as shown on Figure 7-2. Figure 7-8 provides a composite of the main ore accumulations of the Central, Southwest, and Sarytor, Deposits.

According to Ivanov et al., 2000, mineralization in the Kumtor Mine area took place in four main pulses. An initial pulse resulted primarily in pervasive quartz carbonate-albite-chlorite-sericitepyrite alteration, with little gold of economic consequence being deposited. However, this early alteration may have “stiffened” the host rocks sufficiently to make them susceptible to the intensive veining, stockwork and hydrothermal breccia development during the next two pulses that deposited all of the economically significant gold. A few typical examples of mineralized drill core are in Figure 7-9.

The temperature of formation of the second stage veins was 310 ± 15°C, according to Ivanov & Ansdell, 2002. The mineralogy during the main phases includes early K-feldspar followed by later albite and variable amounts of carbonate (calcite, dolomite, ankerite and siderite), quartz, pyrite, sericite and chlorite, in addition to small amounts of chalcopyrite, haematite, barite, strontianite and accessory magnetite, scheelite, ferberite, rutile, cassiterite, sphalerite, galena, native gold and tetrahedrite, as well as a number of silver-gold, lead and nickel tellurides. The feldspars combine to comprise nearly 20% of the ore, the carbonates collectively 25 to 30%, pyrite 15 to 20%, quartz 5 to 10%, and the remainder are host rock inclusions.

The mineralization is most intense, and the gold grade is the highest, where the metasomatic activity was continuous through mineralization phases two and three. This is the case for the Stockwork and SB Zones, and explains their higher-than average gold grades. The last pulse created planar carbonate-pyrite metasomatic rocks that are associated with zones of intense deformation of previously altered phyllites.

Native gold and the gold-silver tellurides are intimately associated with pyrite to the extent that gold grade and pyrite content are “positively correlated” (Ivanov et al., 2000). The gold and the gold-bearing minerals occur as very fine inclusions in the pyrite, with an average size of only 10 microns. This, together with the poor cyanide leach response of the gold tellurides, accounts for the partly refractory nature of the ore. The refractory characteristics are reflected in the relatively low historic and forecast gold recovery of around 80%, despite the very fine grind applied to the pyrite flotation concentrate from which most of the gold is recovered. However, the fine grain size of the gold also renders assaying of this mineralization relatively reliable, with only a small minor nugget effect.

Most of the mineralization takes the form of veins, veinlets, and breccia bodies in which the mineralization forms the matrix. In the more intensely mineralized areas, the surrounding host rock has also been altered. Post-ore faulting is generally parallel to, or at low angles with, the mineralized sequence. These faults often carry significant quantities of graphite, and other carbonaceous components which constitute the sources for the preg-robbing character of some of the mineralization.

Mining operations at the Kumtor Mine use conventional open pit mining methods. Mining in the Central Pit is done on 10 metre benches to allow more efficient use of the larger mining equipment purchased in recent years. Ore at the smaller Southwest and Sarytor pits will be mined on nominal 4 metre benches for better mining selectivity of the smaller ore zones.

Blast holes are drilled using six diesel powered Sandvik DR-460 rig and two Drilltech D55SP rotary-percussion drill rigs, with a hole diameter of 300 millimetres (mm). Charging the holes is undertaken by special bulk explosives trucks delivering either ammonium nitrate with fuel oil, or emulsion explosives for wet holes. The explosives consumption is about 0.26 kg per tonne of ore or waste.

The current plant flowsheet (Figure17-1) reflects the fine-grained nature of the gold and its intimate association with pyrite, and consists of crushing, grinding, pyrite flotation, and two-stage regrinding of the flotation concentrate. Two separate carbon in leach (CIL) circuits extract the gold from the re-ground concentrate and from the flotation tails, with final gold recovery accomplished by electro-winning and refining. The Mill was originally designed with a capacity to process 4.8 million tonnes of ore per year. The Mill throughput is currently 5.9 million tonnes per year or a nominal capacity of 15,900 tonnes per day.

The ore to be milled is managed through a number of stockpiles that receive ore of different metallurgical character and of different grade ranges as determined by grade control data and thus allow blending of the mill feed for optimum gold recovery. A gyratory crusher reduces ore to minus 200 millimetres. The ore is then fed to a coarse ore stockpile from which it is reclaimed for grinding. The stockpile has a nominal capacity of 100,000 tonnes and a live capacity of approximately 15,000 tonnes. Primary grinding is completed in a 9.14 metre diameter by 4.27 metres long 6,340 kW semi-autogenous (SAG) mill operating in closed circuit with a pebble crusher and a ball mill feed sizing screen. Secondary grinding is completed in a 5.49 metre diameter by 7.92 metres long 4,100 kW ball mill in closed circuit with 660 millimetre diameter hydrocyclones.

Hydrocyclone overflow at 80% passing 140 microns gravitates to the flotation circuit, comprised of two parallel banks of nine 42 cubic metre naturally aspirated flotation cells. A bulk sulphide flotation concentrate representing 7 to 11% of the original mill feed is produced with a grade of 30 to 50 g/t Au, about ten times the Mill head grade, and a flotation gold recovery of 87 to 92%. The flotation concentrate is thickened in a 15.2 metre diameter concentrate wash thickener.

Ultra-fine grinding of flotation concentrate is completed in two stages. Flotation concentrate from the wash thickener underflow is first re-ground to 90% passing 20 microns in a 5.49 metre diameter by 7.92 metre long 4,100 kW ball mill in closed circuit with 150 millimetre diameter hydrocyclones. After thickening in the 15.2 metre diameter CIL feed thickener to 50% solids, it is further ground to 95% to 98% passing 20 microns in a 2,600 kW IsaMill that was commissioned in October 2005. The IsaMill provides additional liberation of the fine gold (2-5 microns) enclosed in pyrite.

The concentrate is diluted to 45% solids, pre-aerated for 40 hours and leached for 80 hours in the CIL circuit consisting of six 14.6 metre diameter by 14.6 metre tall agitated tanks in series. Cyanide solution, slaked quicklime, and activated carbon to maintain a concentration of 14 grams per litre (g/L) carbon are added to the concentrate CIL circuit.

The flotation tailings are thickened to 50% solids in a 25.9 metre diameter flotation tailings thickener and leached in the tailings CIL circuit, which consists of three 14.6 metre diameter by 14.6 metre tall agitated tanks in series. Cyanide additions are lower in the tailings CIL circuit compared to the concentrate CIL circuit and carbon concentration is 8 g/L.

Overflow from all four thickeners is recycled through the process.

The carbon in both CIL circuits is moved counter-current to the slurry flow, and the loaded carbon from the first flotation tailings CIL tank is pumped to the third concentrate CIL tank to continue loading. Loaded carbon from the first concentrate CIL tank is pumped to the carbon elution circuit. Total carbon inventory is approximately 300 tonnes.

The loaded carbon is stripped in two parallel 7-tonne carbon stripping circuits. Gold is subsequently recovered by electro-winning. Gold flake is washed from the cathodes, dried and smelted in an induction furnace and cast into doré bars. Carbon is reactivated in a 2,400 kW electrically heated horizontal kiln for reuse in the CIL circuits.

Gold recovery
Historically, the overall metallurgical recovery of gold in the Kumtor processing plant has averaged 79.4%. With KGC current knowledge, the LOM plan annual recoveries are expected to range from 54% to 83%, averaging 78% depending on the head grade and metallurgical characteristics of the ore. Work continues at the Kumtor Mine on implementing strategies to improve gold recoveries.

Gold recovery within the CIL circuits is 30% of the residual gold in the flotation tailings and 90% of the gold contained in the sulphide concentrate. Overall, 90% of the recoverable gold is from the concentrate CIL circuit, with the remainder from the tailings CIL circuit.