Gold-copper skarns have developed mainly along the contact between intrusives and carbonate units. Two different types of skarn have been recognized at El Valle-Boinás. The first is a calcic skarn related to limestone units and the second is a magnesian skarn, called “black skarn”, that is related to dolomite units.

The gold-copper-bearing skarns at Carlés are generally calcic skarns. Better grade goldcopper mineralization is associated with high magnetite and bornite content that is localized in generally continuous, relatively thin (four metres thick) layers of retrograde skarn.

At the El Valle-Boinás deposit, reactivation of fracture zones (along northeast-southwest, east-west and northwest-southeast orientations) produced widespread brecciation and favoured the emplacement of porphyritic dykes. A low-temperature alteration and mineralization event is spatially and genetically associated with the subvolcanic porphyry dykes, which overprint all previous lithologies. Depending on the host rock, there are different styles of hydrothermal alteration and mineralization, such as: sericite-adularia-carbonates (+py-aspy) in granites and skarns; leaching, enrichment, and silicification in skarns (+ native copper and chalcocite), and silicification (+py) in dolomites

Highest gold grades occur where the low-temperature mineralization overprints previously mineralized gold-copper skarn, forming jasperoid or semi-jasperoids with native copper and minor chalcocite and cuprite. The associated geochemistry is characterized by an increase in As, Sb, and Hg. This low-temperature event is the principal gold-mineralizing episode at El Valle.

Gold, and in some cases, base-metal mineralization, has been found in association with late tectonic breccias related to low-angle thrust faults at El Valle-Boinás. The origin of the gold mineralization in these structures is thought to be due to remobilization of previous skarn or jasperoid related gold mineralization. Gold associated with low-angle structures is important at El Valle-Boinás where a significant percentage of the open pit minable gold mineralization extracted from the Boinás East Zone came from this type of structure.

Mineralization at the El Valle-Boinás copper-gold deposit can be grouped into several significant deposits related to the Boinás granitic intrusive and carbonate rocks of the Láncara Formation (Cambrian age).

The gold mineralization system has a strike length of two kilometres and a width of at least 0.5 km. The intrusive is elongated trending N35°E with a length of 500 m and an average thickness of 300 m. A copper-gold mesothermal skarn was developed mainly along the contact between the igneous rock and the carbonate unit.

Late reactivation of the main northeast-trending fracture system was accompanied by two or more phases of epithermal mineralization as well as the intrusion of porphyry dykes. These events produced hypogene oxidation with further enrichment of gold, arsenic, antimony and mercury (Martin-Izard, et al., 1999). Rhyodacite dykes, which are always sericitized, were emplaced along fractures and breccia zones trending north-northeast. The intense silicification along fractures and breccia zones resulted in the formation of hematitic jasperoid that is characterized by enrichment in gold, arsenic, antimony and mercury.

The Carlés deposit is a gold and copper bearing skarn developed predominantly in the Devonian limestones of the lower portion of the Rañeces Formation along the north margin of the Carlés granodiorite. The Carlés intrusion is approximately circular in plan with a diameter of about 750 m. The intrusion is located at the intersection of major faults (east-west, northeast-southwest and southeast-northwest) and it is bisected from west to east by the Río Narcea River. The northern part of the granodiorite is in contact with the lower part of Rañeces Formation and the southern part of the intrusion is in contact with the siliciclastic Furada Formation. Several barren Permian porphyritic and diabasic dykes crosscut the existing lithologies.

Mineralization is continuous for over 1,000 m. It ranges in thickness from 1.5 m to over 25 m, dipping 50º to 90º away from the granitic intrusion. The skarn is known over a vertical continuity of 400 m and remains open at depth.

The Carlés skarn is of calcic composition and is an exoskarn although irregular endoskarn has developed locally. It consists of layers of garnet (grossularite-andradite composition) intercalated with layers of pyroxene skarn, mostly of hedenbergite composition. Retrograde phases of the skarn results in the formation of irregular magnetite layers associated with amphibole. Inside these bands is where most of the copper sulphides and gold mineralization occur. The skarn mineralization transitions into coarse-grained marbles then non-altered limestones away from the intrusive. The latter may show narrow intercalations of distal garnet-pyroxene incipient skarn.

Gold mineralization at Carlés is closely associated with copper sulphides, which consist of disseminated and patchy chalcopyrite and bornite that precipitated mainly in the magnetite zone. Other minerals common in the skarn are arsenopyrite, löellingite, pyrrhotite, and latestage pyrite.

The current mining methods used at Boinás Mine are overhand drift and fill, and transverse longhole stoping. Due to decreasing thickness of the remaining Boinás skarns, RPA changed the design of the longhole mining from transverse to longitudinal where appropriate. Drift and fill mining will continue to be used in the oxide areas of the mine. Ore is hauled to surface via ramp and/or skipped via shaft, depending on location and ore type. Backfill is placed by truck and scoop, consisting of cemented rock fill or waste fill, as appropriate to the mining sequence and geotechnical demands.

Carlés Mine uses longitudinal longhole stoping methods. Ore is hauled to a surface stockpile via underground truck, and transferred to highway-rated surface trucks for transport to Boinás. Backfill is via waste fill, as stopes are separated by rib pillars.

Ore extracted from EVBC is stockpiled and classified according to mineralogy, gold, copper, and/or deleterious element grades. When possible, a suitable ore blend is prepared to feed the mill with material containing 2 g/t Au to 5 g/t Au and < 1% Cu. A front end loader collects the ore from the various stockpiles and delivers it to the crusher feed bin equipped with a 600 mm aperture slotted grizzly screen. The crusher feed bin has an apron feeder that feeds the primary single toggle jaw crusher and any apron feeder spillage is collected on a conveyor belt below. The jaw crusher set to 153 mm. The spillage and the crushed ore which is less than 153 mm in size are transferred to the next conveyor belt equipped with load cell type belt weightometer and this belt discharges into the 100 ton mill feed bin. A level indicator and controller in the mill feed bin controls the apron feeder feed rate to the jaw crusher to maintain the set level in the bin.

Dry crushed ore at P80 < 153 mm is fed to the SAG mill (5.5 m internal diameter by 2.3 m effective grinding length) by means of a conveyor belt equipped with a load cell type belt weightometer. The load signal controls the feed rate from the mill feed bin discharge conveyor onto the conveyor belt feeding the SAG mill. The pulp density in the mill is controlled by adding water to the SAG mill inlet. Oversize material is removed from the mill discharge pulp with the discharge end trommel screen. The oversize fraction is delivered to pebble crushing circuit, while the undersize fraction flows to the common pump box for both the SAG mill discharge and the ball mill discharge. The combined mill discharge is pumped to a cluster of seven 250 mm diameter Weir Cavex hydrocyclones for classification. The cyclone overflow reports to the flotation section and the cyclone underflow is split with part reporting to the gravity concentration section (approximately 100 tph) and the balance reporting to the ball mill feed. The ball mill has a 3.85 m internal diameter by 5.65 m effective grinding length. The cyclone overflow at 40% solids at has a size of P80 < 74 µm. If flotation is not necessary when processing low grade copper oxide ore then this flow can be diverted directly to cyanidation.

Milling reduces the size of the ore to less than one mm and liberates metallic gold, copper and silver particles (in sizes between 100 µm to 500 µm) that is recovered by gravity concentration due to their density differences in comparison to the host rock. Two types of products are obtained from this circuit:
1. High grade gold in copper concentrate – treated in the ILIX plant.
2. Low grade gold in copper concentrate – final product for packaging.

The cyclone underflow that is split off to the gravity circuit passes over a vibrating screen with two mm polyurethane panels. The screen oversize material returns to the ball mill feed. The screen undersize material is split between a bank of 12 (12LM3/2) rougher double spira located near the screen and a Knelson-30 centrifugal concentrator (K30). The K30 concentrate is discharged to the Knelson concentrate tank, while the K30 tails feeds two other banks of 12 (12LM3/2) rougher double spirals. The concentrate from the three banks of rougher double spirals feed the Knelson-20 (K20) concentrator, while the tails flow to the gravity tails box.

After grinding and gravity concentration, significant quantities of fine native gold and native copper and variable amounts of copper sulphide (mostly chalcopyrite) remain. The native metals and sulphides can be recovered in flotation. Copper recovery is extremely important as it has a negative effect on cyanidation and gold recovery.

The mill cyclone overflow flows to the mechanically stirred rougher conditioning tank where flotation reagents are added and the material is sampled. The slurry is conditioned by the addition of the following reagents for 2.5 minutes and is reduced from 40% to 25% solids:
• lime to achieve a neutral pH, if necessary
• Danafloat 123 promoter
• Dowfroth 250 frothing agent

The conditioned slurry flows by gravity to the rougher cells and scavenger cells. The flotation tails are pumped to the CIL feed thickener, while the rougher and scavenger concentrates are directed to a cleaning stage and then a re-cleaner stage. The re-cleaner concentrate gravitates to the concentrate thickener where it is sampled just before the thickener.

The re-cleaner copper concentrate flows to a six metre diameter concentrate thickener. Flocculant is added to the feed box as necessary for optimum solids settling and compaction. The thickener overflow containing suspended solids in the froth flows to a lamella clarifier for clarification where additional flocculant is introduced. The clear lamella clarifier overflow can is used as process water.

After flotation, the tails still contain a significant amount of fine gold and copper. Residual copper will cause problems both in the cyanidation process (high consumption of cyanide) and carbon adsorption (slow kinetics and reduction in the gold adsorption capacity), and also in the subsequent desorption, electro-deposition and smelting processes. Cyanide leaching and the CIL process will recover gold and undesired copper by dissolution and adsorption in the same tanks.

The Carbon-in-Leach tails flow to two tanks for cyanide destruction using the INCO SO2/Air process, one tank of 528 m3 live volume and one of 355 m3 live volume, where residual cyanide is destroyed and heavy metals, e.g. copper, zinc, iron, etc. are precipitated. The detox tails flow to the final tails safety screen that recovers any activated carbon should a carbon retention screen fail. The screen underflow is pumped to the tailings storage facility.

Loaded carbon from the first CIL tank is transferred via a vertical centrifuge pump to a cleaning screen and to an elution column, while the slurry returns to the first CIL tank. Elution follows the Anglo-American AARL process, using a nine cubic metre cylindrical column with a capacity for four tonnes of carbon.

Stripped carbon is transferred to a self-draining chute in the regeneration kiln by means of a water pressure injector. The regeneration kiln is a horizontal rotary kiln with a capacity of 250 kg/h and is heated by propane to a temperature between 650°C and 750°C to thermally reactivate the carbon. The reactivated carbon is screened to remove carbon fines before being returned to the last CIL adsorption tank. Regeneration of the carbon takes approximately 15 hours to complete.

The eluate from desorption and elution is stored in the electrolyte tank of the EW process. There are four electrolyte cells equipped with rectifiers and each cell has nine stainless steel wool cathodes. Electrolyte solution circulates through the cells and contains less than 10 g/t Au. Gold and silver are electrodeposited on the cathodes and later washed off as EW sludge. EW takes less than 16 hours to complete.

Gold deposited as cathodic sludge is removed on a weekly basis and calcined in a kiln at 750°C before direct smelting. After calcination, the gold and silver residues are mixed with fluxes (silica, borax, and nitrate) before smelting, which is carried out in a 0.04 m3 smelting furnace capable of reaching 1,200°C. The refined gold is then cast into molds as doré bars.