Search Documents

By Sergio Romo, Gerardo Fuentes, Adriana Garcés

Refractory ores with gold entrapped in a sulfide matrix, must be processed in a first stage with chemical or physical treatment in order to liberate the gold for a subsequent step with cyanide. The goal of this research was to compare the refractory behavior of a gold concentrate treated using two pretreatment methods: pressure oxidation and fine grinding. Gold concentrate from Viacha Mining in Bolivia, assaying 50.11 gpt, with a P80 of 131 ?m was used. The concentrate contains arsenopyrite and pyrite as main sulphides in the matrix. A preliminary test with cyanide resulted only 13.5% of gold extraction with 4 hours of stirred leaching. Using an autoclave oxidation with sulphuric acid and 12 bar oxygen pressure at 220°C pretreatment, the cyanide gold extraction increased to 69.0% in 4 hours. When material was ground in a pulverizer, producing a P80 of 20 ?m and the product directly cyanided, stirred leaching gave a gold extraction of 65% in 4 hours. For both techniques, the effect of temperature, sulfur produced in autoclave oxidation, optimization of the cyanide and the nature of residues were also studied.

Jan 1, 2014

By J. Sugai

Formability of extruded round AZ31 magnesium alloy tubes with different manganese concentrations was evaluated by performing bending tests and measuring strain along the length of the tubes during the bending process. It was found that an increase in the manganese content from 0.3 % to 0.7% did not exert any significant effect on the values of yield stress, elongation, or compression to fracture. However the hardness tended to increase and the UTS decreased. The compressive ductility was significantly lower than tensile ductility for all manganese concentrations. Consequently, the formability of tubes decreased, causing cracks in areas where compression strain occurred during bending. It was also found that a larger strain occurred in the latter half of the region along the length of tube between the clamp-end and the tube-positioning block. Within this region, alloy tubes with higher manganese concentrations, 0.5 and 0.7 %, there was a clear strain peak within the strain profile, while for the 0.3 % Mn alloy, a more uniform strain profile was exhibited, indicating a more uniform deformation. This was confirmed by microstructural analysis of the specimens.

Jan 1, 2006

By K. Kawagishi, J. Ang, A. Sato, H. Harada

"We have recently proposed a new ""EQ Coating"" bond coat system to solve the problem of secondary reaction zone (SRZ) formation caused by bond coats applied on the hot section components of modem jet engines. A viable EQ coating alloy was previously found for a representative znd generation superalloy, TMS-82+. It would be ideal if the same could be done for a representative 5th generation superalloy, TMS-173. In this paper, we will show that the coating with the best oxidation resistance in the EQ system of TMS-173, although possessing considerably better resistance than TMS-173, did not have sufficient oxidation resistance. In order to improve the oxidation resistance of the EQ coating for TMS-173, we have ventured into ""Pseudo-EQ coatings"". A diffusion couple between the Pseudo-EQ coating and TMS-173, after 300hrs at 1100°C, had a 0-15µm IDZ, but no SRZ. A diffusion couple between a conventional coating and TMS-173 formed a 80-150 µm SRZ. After a 1 hr oxidation test at 1100°C, the PseudoEQ coating formed a thinner oxide layer than the EQ coating, but still contained NiO and spinel. Aerospace"

Jan 1, 2006

By K. G. Markkola

Since 2005, Compafifa Contractual Minera Candelaria (CCMC) has defined various strategies which aim to optimize the usage of the available energy in each of the SAG grinding circuit modules (SABC-A configuration). For this purpose, the use of larger size balls along with operating with an optimum balls/load ratio has been implemented. Additionally, in the ball mills, the amount of balls in relation to the equipment's usable diameter has been optimized. In the pebble crushing circuit, the percentage of material passing 12 mm (112 in) product has been increased by 15% by: (1) modifying the crusher mantle profile, (2) implementing a strategy for maximizing the power in each crusher, (3) using continuous improvement tools (Lean - Six Sigma), (4) redesigning SAG mills discharge grates, and (5) changing discharge grates in order to control the crusher feed size distribution profile. All these activities have improved plant throughput by 1,000 t/d with no significant capital investment. As part of the energy consumption optimization study, we have considered determining the optimum operational rate of each SAG mill, in relation to the mill liner life and also the improving the control strategy in conventional grinding.

Jan 1, 2007

By M. K. Gangal, A. E. Hall, S. B. V. Stewart

"An investigation into the problem of atmospheric fog in Canadian mines is described. Fog occurs in several mines and the reduced visibility can result in the closure of workings for safety reasons. Information on the mechanisms of fog formation are discussed and the methods for improving conditions are reviewed with their relative costs.INTRODUCTIONAtmospheric fog occurs in several Canadian mines during spring and summer. The fog which reduces visibility, contributes to loss of production and increased accident risk . Closure of ramps in mechanized mines causes mining to be stopped or rescheduled, and stoping and grade control may be affected . Increased corrosion of equipment may result from acidic fog .Fog is commonly caused by saturated air being compressed and heated by an inlet fan and then being circulated in an ascensional system with heat transfer to the walls of wet airways. This causes supersaturation and water droplets condense to create fog, with liquid water contents of 0.5 g/kg to 3 g/kg of air. These concentrations reduce visibility to less than 10m and prevent mining. Less severe fog is usually tolerated as a nuisance, about which little can be done."

Jan 1, 1989

By L. Fernandes

A factorial study was performed to investigate the relative effects of temperature, copper concentration and meso-2,3-dimercaptosuccinic acid (DMSA) concentration on the performance of thiosulfate leaching of gold. A total of 24 leaching experiments were performed representing eight unique combinations of high-low levels of the three factors performed in triplicate. All investigations were performed in an electrochemical cell with an electrolyte composed of 0.1 M sodium thiosulfate with a combination of copper (either 0 mM or 1 mM) and DMSA (either 0???or 1µM). All leaching investigations were conducted under open circuit while the electrolyte, which was open to the atmosphere and mechanically stirred, was maintained at either 25°C or 50°C. After 1.5h of leaching the amount of gold in solution was quantified using ICP-MS analysis from an aliquot of the electrolyte. Results indicate that DMSA addition improved gold leaching performance, most notably at 50°C and with addition of copper. Under this unique condition, the quantity of gold in solution was reproducibly and significantly above the average response of all other factors studied.

Jan 1, 2014

By W. R. Dodge

THE growth and development of the assay department of the Hollinger Consolidated Gold Mines, Limited, is a similar story to that of the mine itself. As the mine developed and enlarged, so also increased the demands upon this department, necessitating an increase of capacity for fire assaying and a widening of the scope of work to include chemical analysis, metallurgical tests, research, and screen analysis. The yearly total of determinations, taking an average for the past six years, is approximately 350,000. The major portion of this figure represents fire assaying for gold and silver, being approximately 950 fire assays per day. The assay department occupies the ground floor of the administration building and covers approximately 4,000 square feet of floor space. Ready access is had by motor truck and electric railway for the delivery of samples and the removal of crushed rock and rubbish. The combined weight of all sample rejects amounts to about 5,000 lb. per day, which is sent to the mill bins. There are six departments, each occupying a separate room-grinder room, fluxing room, furnace room, parting room, weighing room, and flux-mixing and storage room.

Jan 1, 1931

By S. Markstrom

The Boliden AITIK copper mine is located 100 kilometres north of the Arctic Circle in northern Sweden. A pre-feasibility study was conducted in 2005 with the aim of increasing the capacity of the existing processing plant. After analysis, options to upgrade the existing plant were abandoned in favour of building a new processing plant, with higher capacity and therefore greater profitability. A new plant could be designed with greater flexibility to make potential future expansions easier. A feasibility study for a new 36 Mt/y processing plant including fully AG grinding circuits and flotation was completed in 2006. The project was approved by the Boliden board shortly thereafter. Construction started in the beginning of 2007 and was completed during 2009. Commissioning of the new plant was undertaken in the first half of 2010. The grinding circuit consists of two lines with each line consisting of a 22.5 MW AG mill and a 10 MW Pebble mill in closed circuit with a spiral classifier. The nominal production level is 2 200 t/h for each grinding line. This paper presents the start-up and operation of the grinding circuit.

Jan 1, 2012

By K. V. Konigsmann

"The mill at Mattagami Lake Mines treats 3,850 tons per day. The mill feed is a heavy sulphide ore. It is composed ofPyrite & Pyrrhotite - 40%Marmatite - 17%Chalcopyrite - 2%Magnetite - 10%Talc and other Silicates - 31%The sulphides are intimately associated with each other and often finely disseminated throughout the siliceous gangue. An abundance of middling particles exist in the circuit and a delicate balance has to be maintained between concentrate grades and metal recoveries.The metal content of the mill feed fluctuates between 0.4 to 0.7% for copper and 6 to 13% for zinc.The flow sheet of the concentrator is shown in Figure 1.Eight sample streams have been selected for on-stream analysis, namely1. Copper Concentrate 2. Copper Cln Tailings 3. Copper Tailings 4. Zinc Concentrate5. Zinc Rgh Concentrate6. Zinc Cln Tailings7. Zinc Rgh Tailings8. Flotation TailingsThe mill feed is not being analysed. Only a very coarse sample could be obtained which would be difficult to pump, and the fluctuations in the copper content are too small in absolute terms to have much of a bearing on circuit operations."

Jan 1, 1969

By J S. Kermeen

This paper deals with drilling and blasting techniques at the Phoenix mine, near Grand F1nks, B.C., which is now operated by The Granby Mining Company Ltd. The difficulties of open-pit mining through former underground workings are unique to this operation. Rock types are related to drilling conditions, and the methods used, or contemplated, to overcome the several difficulties are outlined. The equipment used is also mentioned. Finally, blasting techniques and costs are reviewed. Introduction T HE Granby Mining Company operates on open-pit copper mine and concentrator at the site of the original Phoenix mine, near Grand Forks in south-central British Columbia. The pit is a side-hill cut measuring approximately 1,000 by 2,000 ft., and covering a vertical range of from 4,550 to 4,950 feet above sea level. The mill is currently operating at 1,900 tons per day. A 4.5 to 1.0 waste to ore ratio necessitates the removal of some 15,000 tons of ore and waste per working day. Mining conditions at Phoenix are extraordinary in that all the ore zones currently being

Jan 1, 1965

By A. D. McKay

Uranium exploration expenditure in Australia reached a record level of A$220.5 million in 2008, which has subsequently declined to A$185.2 million for the year ended 30 June 2009. Recent discoveries include: Ranger 3 Deeps east of Ranger open pit (Northern Territory), Pepegoona and Yadglin sandstone hosted deposits (South Australia), and Thunderball project (Northern Territory). Australia?s Reasonably Assured Resources recoverable at costs of <US$80/kg U at June 2009 are estimated to be 1,210,000 t U, 4% higher than the estimate for December 2008. The main additions were in the south eastern portion of the Olympic Dam orebody. The Olympic Dam Expansion is based on a large open pit and will increase output to an annual capacity of 19,000 t U3O8; 750,000 t copper; 800,000 ounces gold. The Four Mile in situ recovery project was approved in July 2009. An ion exchange facility is to be constructed at Four Mile and uranium-bearing resins will be transported to the Beverley plant to recover the uranium. Recent developments at Ranger mine, and Honeymoon, Oban and Yeelirrie projects are discussed.

Aug 1, 2010

By I. A. Menshikh, S. A. Kotov, N. L. Shinkarkin

All the cycle for conversion of natural and reprocessed uranium of ROSATOM’s “TVEL” Fuel Company is concentrated at the Siberian Chemical Plant. Conversion capacities are located on two sites, i.e. the Radiochemical Plant, which refines feedstock, and the Conversion Plant, which produces uranium oxides and uranium hexafluoride. Liquid radioactive waste arising in the process, after pretreatment, is transferred to a dedicated national operator for underground disposal. Environmental strategies pursued by the Fuel Company involve the introduction of process solutions intended to eliminate liquid radioactive waste and convert it into a solid phase. In the 5–7 years coming, work will be performed to refurbish and upgrade conversion facilities of the Siberian Chemical Plant that will principally change its radioactive waste management scheme. This paper describes the current waste management scheme and sets out planned schemes for management of natural and reprocessed uranium waste, which will be implemented in the refurbishment of conversion facilities of the Siberian Chemical Plant. The new technologies for the management of liquid radioactive waste from uranium conversion described in the present paper have an undeniable advantage over the current technology mainly due to significantly less environmental harm.

Jan 1, 2020

By M. Doucet

The introduction of high-nickel anodes to CCR in 2006 resulted in an increase of nickel in anodic slimes. Removal of the nickel in these slimes to below 3.0% is necessary in order to prevent slag viscosity problems in the subsequent pyrometallurgical step. A lower specification of 0.5% was chosen in order to minimize Ni losses to the slag and to recover this nickel as NiSO4. A 2-stage autoclave process, which was patented by CCR in 1992, was chosen to treat the high nickel slimes. Laboratory and plant?scale tests were performed in 2004 in order to validate the results of this process with the present slimes. These results along with corrosion tests were used to design two brick-lined autoclaves. These autoclaves were installed in October 2006 and successfully commissioned within 8 weeks. A "Design of Experiment" test program was performed in 2006 on an alternative autoclave process that reversed the cycles in the patented process. The effects of temperature and acid content were modeled using the results. The alternative process will be validated with the new autoclaves.

Jan 1, 2007

By B. Zhu, W. E. Lee, N. Li

Al powder is used extensively as an antioxidant in carbon-containing bricks, but it cannot be used directly in water-based carbon-containing castables due to its hydration tendency. To overcome this problem, in this work, SiO2-PVA (polyvinyl alcohol) inorganic-organic hybrid coatings have been prepared via sol-gel processing, and used to coat Al powder. The hydration behaviour of uncoated and coated Al powders has been investigated in water at 60 or 100°C for 36h by examining phase and microstructural changes. While significant hydration is detected in uncoated Al powder, no obvious hydration is detected in coated Al powder, indicating that its hydration resistance has been greatly improved. Oxidation resistance of graphite coexisting with uncoated or coated Al powder was also investigated in air at 1250°C for 1.5h. The oxidation ratio of graphite coexisting with coated Al powder was similar to that of graphite coexisting with uncoated Al powder, indicating that both powders showed similar antioxidising behaviour. This, along with the much improved hydration resistance, makes the coated Al powder a potential antioxidant in carbon-containing castables systems.

Jan 1, 2004

By M. Askari, M. H. Parsa, S. Kashani-Nejad

Sarcheshmeh refractory consumption rate is about 7.5 - 8 kg/ton of the produced copper, which is almost three times higher than the other smelters practice. High destruction rate of tuyere zone refractory controls vessel productivity and operational lifetime. In this research with the aid of cold model and computational fluid dynamic (CFD), melt recirculation pattern and gas - liquid plume properties in the wide range of gas blowing conditions were studied. Results revealed that in addition to the well-known mechanical refractory destruction mechanism, fast tuyere zone refractory destruction could also be related to the cold bubble cluster formation near tuyere tips and introducing severe thermal shock to the refractory surface. Removing accretion and its strong bonds to the refractories around tuyere tip, causes refractory mechanical destruction and increases tuyere diameter. Increased tuyere diameter, in turn assists gas discharge in the bubbling regime. Air blowing in the bubbling regime result in the formation of big bubble clusters, bubble breakup and back attack occurrence which all found to have major impact on the refractory lifetime. Outcomes of this research, contributed to the plant converting management, blowing hardware and instrumentation improvement.

Jan 1, 2004

By K. Swedburg, M. Samuels

"The expansion of the concentrator at New Gold’s New Afton copper-gold mine in British Columbia, Canada, completed in 2015, included an increase in the required retention time of the cleaning circuit. On site pilot testing of a Woodgrove Staged Flotation Reactor (SFR) had shown that a high grade copper concentrate would be produced with excellent recovery. The SFR was selected rather than conventional tank cell technology based on the much smaller footprint (which rendered a separate and expensive building extension unnecessary), high concentrate grades demonstrated in testing, and the lower energy requirement. Three SFRs, each capable of taking a nominal feed of 700 cubic meters per hour were installed and commissioned 10 months after the receipt of order. The SFRs operate in series and all of the cells produce a final concentrate typically grading 32% Cu, while the overall SFR stage recovery averages 55%. The tailings from the SFRs are treated in the original conventional three-stage cleaner circuit where the remaining values are recovered in a concentrate grading 28% Cu. Design production was achieved within days of the initial start-up of the SFR circuit.INTRODUCTIONThe New Afton concentrator, located near Kamloops, British Columbia, Canada processes a copper-gold ore with a designed head grade of 0.96% Cu, 0.72 g/tonne Au, and 2.22 g/t Ag.The New Afton Mine is a block cave mining operation. Ore is undercut from the bottom up and removed at the extraction level from a grid of draw points. Ore is trammed to the ore passes by underground loader and dropped to a haulage level for transport by trucks to the gyratory crusher. From the crusher, the ore is conveyed to the mill via a 4.5 kilometre long conveyor system. Block caving was chosen for the New Afton deposit because the underground infrastructure could be built below the historical open pit of the Afton Mine (a previous operation of Teck Resources) and this method is favorable to large-scale low-cost mining.The New Afton deposit is a copper-gold, alkalic porphyry system. Hypogene mineralization is characterized by discontinuous copper sulphide veinlets and disseminations, principally chalcopyrite and minor bornite, along with pyrite as the second most dominant sulphide. Gold particles vary in size from 5 µm to 40 µm and occur as native gold and electrum, associated mostly with the chalcopyrite grains and never with the pyrite. In some instances, a second alteration phase has resulted in the copper sulphides being replaced by tennantite-tetrahedrite locally and along faults that transect the mineralized body. The design head grade of arsenic is 244 PPM, although the presence of tennantite is inconsistent and can vary from trace levels to 15% of the total copper. No arsenopyrite is present in the hypogene ores. Supergene mineralization, consisting of native copper with accessory chalcocite occurs in minor amounts along faults and highly oxidized near-surface portions of the deposit."

Jan 1, 2016

By John F. Kearney

Labrador Iron Mines (?LIM?) Schefferville Projects consist of 20 major DSO deposits lying along a 120 km strike length in both the provinces of Newfoundland and Labrador and in Quebec. The town of Schefferville, Quebec lies close to the middle of this 120 km strike. LIM?s various properties comprise twenty different iron ore deposits which were part of the original Iron Ore Company of Canada (?IOC?) direct shipping Schefferville operations conducted from 1954 to 1982 and formed part of the 250 million tons of historical reserves and resources previously identified by IOC. LIM intends to develop all these resources in five separate stages. The first four stages will see the various deposits treated in three beneficiation plants, to be built adjacent to several of the deposits that comprise the Central, South Central and North Central Zones. The final stage is not yet planned but will likely involve a fourth plant. A Feasibility Study has not been completed on any of these Projects. LIM plans the development and mining of the various deposits in separate Stages. It should be noted that only the James and Redmond deposits have been permitted. All other deposits require environmental assessment and permitting. Only the James and Redmond deposits have been subject to detailed mine design. The Houston deposit has had preliminary mine design carried out. All other deposits will require detailed mine design when further geological data is collected and compiled. Detailed metallurgical testwork has been carried out on blue ore from James, Houston, Redmond and Knob deposits and to a more limited extent on yellow and red ores from these deposits. Only limited review work has been carried out on the high-silica blue ores. All of this work will be needed to be complete to confirm the final process routes. Additionally some metallurgical work will be required to be conducted on each of the other deposits to characterize their metallurgical response and to confirm expected recovery levels. Detailed mine design and metallurgical testwork has been carried out on those deposits that will form the basis of production for the initial years but further exploration, testwork and design is required for all the remaining resources. The first plant at Silver Yards and mine construction at James are now complete and initial production is expected in the spring of 2011. Metallurgical testing of the first stage deposits has established that the ore can be successfully upgraded to marketable standards. Based on the current testwork, a program of plant enhancement on the Silver Yards plant is now planned with the objective to increase both the plant throughput and recovery. Production of direct shipping lump ore and sinter fines is forecast at around 1.5 million tonnes in 2011 and, based on current and historical resources, growing to approximately 5 million tonnes in 2015 and thereafter. Production from the first four stages is expected to continue at about this rate through to about 2028.

May 1, 2011

By Lastram Rolando, Larry J. Hulbert, Louis J. Cabri, J. H. Gille Laflamme, John L. Campbell, Soey H. Sib, Chris G. Ryan

"The mineralogy of samples from the Wellgreen Cu-Ni-Pt-Pd deposit, Yukon Territory, is complex and the metals of interest are contained within fine-grained minerals which vary in size and mineral associations from sample to sample.The major Ni carriers are pentlandite and violarite; the major Cu carrier is chalcopyrite. About 700/0 of the major Ni carriers and 78% of the major Cu qurier will be liberated at a grind of 80% -U,.7 ,an (-500 mesh) for typical low-grade gabbro (West zone, sample No.1). For typical ""clinopyroxenite"" (West zone, sample No.2) and typical ""clinopyroxene-rich peridotite"" (East zone, sample No.3) the liberation would be 71% and 65% for Ni, 74% and 60% for Cu, respectively. Fine grinding, however, cannot liberate the Ni contained in silicates and pyrrhotite, estimated to total 24.5% for typical low-grade gabbro (West zone, sample No. I), 15% for typical ""clinopyroxenite"" (West zone, sample No.2), and 12% for typical ""clinopyroxene-rich peridotite"" (East zone, sample No.3).The principal Pt carrier (sperrylite) varies considerably in grain-size between samples and fine grinding, required to liberate the major Ni and Cu carriers, may be counterproductive in samples where sperrylite is relatively coarse-grained, as some would then be unrecovered (e.g. sample No. 1). In addition, sperrylite in samples No.2 and No.3 is closely associated with magnetite. Palladium occurs in diverse ways. It occurs in the form of several Pd minerals (e.g. sudburyite, testibiopalladite, merenskyite), as a replacement for Ni in several tellurides, antimonides and tellurantimonides (e.g. melonite), or as very low concentrations in pentlandite (up to 34 ppm), chalcopyrite (up to 9 ppm), and pyrrhotite (up to 5.6 ppm). Palladium minerals and Pd-bearing minerals are strongly associated with chalcopyrite in sample No. 1 and will thus report to a Cu concentrate.However, 22% of the total Pd is in solid solution in pentlandite in this sample and will thus report to the Ni concentrate. In samples No.2 and No.3, less Pd occurs in solid solution in pentlandite, and the PGM are distributed mainly among pentlandite, chalcopyrite, and magnetite in sample No. 2, and among pentlandite, pyrrhotite, chalcopyrite, and magnetite in sample No.3. Therefore, the feasibility of magnetic concentration should be investigated for samples like No. 2 and No.3, in order to improve Pt and Pd recoveries. Rhodium was found to occur as rare hollingworthite, in sparse cobaltite (up to 2.7%) and, as low concentrations in chalcopyrite (to 10 ppm) and pentlandite (to 12 ppm) by proton microprobe. Iridium occurs in hollingworthite (up to 7.8%) and very rarely as the native metal or in an undefined Re-Ir-Qs..Ru alloy. Ruthenium was determined by proton microprobe analyses to occur in pentlandite (up to 13 ppm) and chalcopyrite (up to 10 ppm), in addition to rare laurite and Re alloy."

Jan 1, 1993

By P. J. Deveaux, M. J. Kreczmer

The Tartan Lake deposit 12 km northeast of Flin Flon, Manitoba, was discovered in March of 1984 during the investigation of a VLF anomaly. The discovery hole intersected 16.83 metres averaging 21 gm Au/tonne. Since then, 99 holes, over 21,000 metres, have been drilled outlining 419,491 tonnes grading 11.79 gm Au/tonne in a di stinct group of mineralized lenses referred to as the Main Zone. Three other mineralized zones are known on the property but no reserves have been allowed for these, the South , East and Base Line zones. The Main Zone consists of seven parallel, east-west striking, quartz lenses. The lenses dip steeply to the north and plunge steeply west to northwest. Lens number l strikes for 160m and extends vertically to the 90m level. Within th is lens there is a 60m interval of low grade materi al. Lens number 2 is located about 5-l0m within the foo twall of Lens number I. Near surface this lens stri kes for 40m, at 200m level the strike length increases to 200 m. To date the deepest mineralization on Lens 2 has been intersected at a vert ical depth of 527m. Five smaller lenses are located in the foo twall of Lens 2 , containing gold va lues similar to the two main lenses but erratically distributed. The South Zone, located 100m south of the Main Zone, strikes east-west fo r l00m above the 100m level. The East Zone, found 230m east of the Main Zone, strikes fo r 80m and may possibly be an ex tension of the South Zone. The Base Line Zone is located 275m south of the Main Zone. Sporadic, but very significant gold intersections have been obta ined1n all three zones. Gold-bearing quartz-ca rbonate tourmaline veins are generally confined to a silicous crystal tu ff hori zon. The crystal tuff exhi bits fi ve distinct types of alteration: silicification, sericitization, ca rbonati zation, chlorit ization and pyritization. Silicification has resulted in a perasive addition of silica and the generation of multiple quartz veins. Sericiti zation is restricted to the crystal tuff. All rock units are pervasively carbonatized . Ankerite is abundant in the quartz veins and within the shear zones, particul arly those located along the chilled margin of a nearby gabbro-diorite intrusion. Chloritization is widely spread th roughout the rock units, while pyritization fo rms a broad alteration halo associated with the orebody. Sulphide accesso ry minerals associated with the gold , whi ch fo r the most part appears to be free, are chalcopyrite and pyrite. Generally, the higher grade gold values are related to an increased content of sulphides in the qua rtz veins. The average sulphide content in the deposit is less than 5 %.

Jan 1, 1986

In the leaching and recovery of gold from copper containing gold ores using cyanidation, some copper minerals also dissolve which can have a detrimental effect on the process efficiency and economics. There are also environmental concerns with current operations discharging cyanide and copper cyanide species into tailings dams, and strict regulations prohibit this for greenfield plants. Cyanide destruction however is not an economic option for new plants treating copper-gold ores. This paper evaluates a potential recovery process to overcome the copper cyanide problem. Despite gold dissolution occurring in copper cyanide solutions having no free cyanide, gold recovery from ores is lower compared to free cyanide. Thus, cyanide addition has to be optimised to maximise the gold recovery whilst endeavouring to eliminate free cyanide at the completion of leaching (CN:Cu less than 3:1). Subsequently, copper cyanide can be recovered using activated carbon. The loaded copper cyanide is easily recovered from the carbon with a cyanide soak and distilled water wash such that the eluted CN :Cu ratio is less than 4:1. The copper could be recovered efficiently by electrowinning, by decreasing the pH of the eluant to 5. The remaining cyanide solution, after electrowinning, is then neutralised to form NaCN and recycled.

Jan 1, 2005