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By Jean-Claude LeRoy, Herman Keller

"INTRODUCTIONEROCON's work is the reclamation of mine wastelands. It has had considerable experience in particular in the reclamation of gold mine tailings in the Canadian Shield since the latter part of the sixties when most gold mines were closing down.A typical example is the famous Hollinger Mine in Timmins which produced $566 million worth of metal from 1910 to 1968 and had left behind some 56 million tons of tailings covering about 1,000 acres which were reclaimed on very short order and are now lush meadows of agricultural quality.To date, practically all gold mine tailings have been reclaimed once and for all by establishing a permanent, self-sustaining and maintenance free vegetation cover.RECLAMATION & THE MINING TOWNSThe first appreciation of people in a mining community where the tailings have been revegetated will be noticing one day the elimination of dusty conditions that used to arise formerly on every dry and windy day. In that respect, EROCON encourages not only vegetating the tailings but also the immediate surroundings of the mills and plants themselves at active mines. The elimination of windborne dust will result in an appreciable decrease in maintenance costs of plant and machinery and more pleasant working conditions for all."

Jan 1, 1975

By H. O. Ghaffari

A complex network approach on a rough fracture is developed. In this manner, some hidden metric spaces (similarity measurements) between apertures profiles are set up and a general evolutionary network in two directions (in parallel and perpendicular to the shear direction) is constructed. Also, an algorithm (COmplex Networks on Apertures: CONA) is proposed in which evolving of a network is accomplished using preferential detachments and attachments of edges (based on a competition and game manner) while the number of nodes is fixed. Also, evolving of clustering coefficients and number of edges display similar patterns as well as are appeared in shear stress, hydraulic conductivity and dilation changes, which can be engaged to estimate shear strength distribution of asperities.

May 1, 2009

By D. G. MacBeth

Most belt scales on the market are factory calibrated to minimum accuracy figures of 0.1% - 0.2% at several load points. However, if the same belt scale is not properly installed and maintained on the customers conveyor then the scale will not measure up to the claims of the manufacturer or to the requirements of the operator.

Jan 1, 1975

By G. Halverson, K. Blower

"Giant Yellowknife Mines Limited (Giant) has been producing gold since 1948. Located at Yellowknife, Northwest Territories, Canada, see Figure 1, the mine processes approximately 1,200 short tons per day of 0.23 ounce gold per ton of ore.The majority of the gold in the ore is sub-microscopic and associated within the mineral arsenopyrite. This association complicates the extraction method since straight cyanidation will only recover 30% of the gold. The operations at Giant involve four essential steps, with an overall recovery of 87%. The steps shown in Figure 2 are crushing and grinding, flotation, roasting and cyanidation. These processes result in a number of effluent waste streams characterized in Table 1. These streams are combined and sent to a tailings storage pond for subsequent treatment prior to discharge to the environment.Giant is located just north of the 62nd parallel where winter temperatures drop to -40°C for extended periods. This entails storing effluent in dams over the seven month winter period for treatment during the relatively short summer season.The treatment method currently employed at Giant is an ""alkaline chlorination-arsenic precipitation-flocculation"" process. The pond water is subjected to three phases of treatment prior to being discharged to the environment as shown in Figure 3."

Jan 1, 1987

By Jose R. Hernandez-Aguilar

"Three models to predict the effect of bubble size (Db) on metallurgical response were tested in a pilot-scale flotation column at Teck Highland Valley Copper mine. The column was fed with slurry collected from the Cu/Mo separation circuit. The first model was the one developed by Gorain et al. (1997), which assumes that the collection zone rate constant (kc) is inversely proportional to Db. The second model was the one proposed by Hernandez-Aguilar et al. (2010), which considers kc as an inverse square function of Db. The third model was the one developed by Yoon and Luttrell (1989) which provides expressions to describe the efficiency of particle collection (E) based on first principles. The results showed a significant bias between experimental and predicted data using the model of Gorain et al. (1997), possibly reflecting the model’s inability to explain the impact of bubble size, bubble rise velocity and flow regime on E. The model of Hernandez-Aguilar et al. (2010) well described the experimental evidence. The model of Yoon and Luttrell gave accurate and unbiased predictions. However, Yoon and Luttrell’s approach calls for a criterion to determine whether the flow regime is potential or intermediate, a subject that requires more fundamental research.INTRODUCTIONParticle collection, i.e., bubble-particle collision followed by attachment, is the basis of flotation. In a quiescent environment (such as the one observed in a column, or away from the impeller in a mechanical cell), a collision event can take place as long as the “degree of turbulence” caused by the rising bubble does not prevent particle and bubble to flow in close proximity. Upon collision, the particle slides on the bubble surface for a period of time. If the particle is hydrophobic, attachment occurs when the “sliding time” is longer than the “induction time”, which is the time required to thin and rupture the intervening liquid (usually water) film forming a three-phase contact angle. While the induction time is largely influenced by the degree of hydrophobicity of the particle, the degree of turbulence is governed by the bubble size. Generally, direct control of bubble size is not exercised by most mineral processing practitioners, in spite of its obvious importance. Studies to model the effect of bubble size on particle collection using fundamental approaches have been conducted (e.g., Anfruns and Kitchener, 1977; Webber and Paddock, 1983; Yoon and Luttrell, 1989; Dai et al., 1999). However, these models have been validated under highly idealized laboratory conditions (micro-scale experiments using single minerals and uni-sized bubbles), far from resembling those existing in practical industrial applications. Fundamental studies to interpret the effect of bubble size on particle collection in industrial environments are largely absent in the literature. Partly because of this lack of practical knowledge, many mineral processing practitioners perceive laboratory “bubble size” studies to be valuable academic exercises, but not something that can be directly applied to daily production issues."

Jan 1, 2011

By John Lindsay

Outline ??Recovery plant requirements and concepts ??Common recovery plant unit processes ??Recovery plant design considerations ??Peripheral issues Process Design ??Liberation ??Separation ??Recovery Processing ??Density 3.5 g/cm3 ??Luminescent under x-ray ??Hydrophobic (non wettable) surface ??Non-magnetic ??Inert to strong acids / bases

Apr 1, 2005

By Daniel Hodouin, Claude Bazin, René del Villar

"A traditional linkage has long been established at Laval University's Department of Mining & Metallurgy between the Mineral Process (MP) and the Mining Exploitation (Min) on one hand, and between the Extractive Metallurgy (EM) and the Materials Engineering (Mat) on the other. This academic separation of Mineral Processing from Extractive Metallurgy at the educational level has become increasingly at odds with today's industrial practice in which both disciplines are not only complementary but also often intertwined. For instance, in gold ore processing, comminution, flotation, and gravity separation units are interfaced with atmospheric or pressure leaching modules and with carbon adsorption and electrowinning units. The new copper plants in South America combine comminution and flotation circuits with heap leaching, solvent extraction and electrowinning processes.In order to reflect this new industrial reality and to harmonise the options offered in both its engineering programs, the Department of Mining and Metallurgy of Laval University has introduced major changes in its curriculum in 1996. These changes reflect also the recognition that the somewhat secondary status given so far to Mineral Processing and to Extractive Metallurgy within their ""mother"" programs, Mining/Mineral Engineering and Materials metallurgical Engineering respectively, was in sharp contrast with their actual significance for the Canadian economy. Leading industrial voices have raised similar concerns and the Canadian Engineering Accreditation Board has also addressed this problem in its latest program evaluation.The changes implemented at Laval aimed at a) better representing MP and EM within the two undergraduate programs; b) training top-level versatile and multidisciplinary engineers, fluent in all aspects of MP and EM, irrespectively of which of the two ((mother)) programs they had chosen and capable of meeting today's industrial requirements; c) training students in process analysis, optimisation and automatic control using state-of-the-art software, and d) taking into account the co-operative nature of both undergraduate programs.Results are very encouraging in terms of student enrolment and industrial response. However, some concern is expressed on the sustainability of the renewed curriculum considering the drastic budget restrictions in effect at Laval as in most other Canadian universities."

Jan 1, 1999

By Rudy Tenbergen, Dick Farmery, Jim Naylor

"Paste backfill has replaced cemented rock fill as one of the components to improve ground · control and mine safety at the Macassa Operation of Kinross Gold Corporation in Kirkland Lake, Ontario. It is the first operation to utilize "" flash thickening ""of very fine gold tailings to a paste consistency in the ""Paste Production and Storage Mechanism"", a joint technical development of Inco and GL&V. The tailings paste is mixed with alluvial sand to produce an engineered quality paste backfill. The development leading up to the decision to proceed with this novel technology is discussed, as well as commissioning of the backfill operation.IntroductionThe occurrence of a major rockburst on November 23, 1993 prompted a review of mining and backfill procedures at the Macassa No 3 shaft . This shaft is some 7238 ft deep and lies in a mining belt that is prone to seismic events. The recommendations were to change the mining from underhand cut and fill to mechanized long hole stopping and to look at alternatives to the rockfill backfill system in place. The Macassa cemented rock backfill operation, while producing the desired fill stiffness had drawbacks as compared to paste fill. With paste fill, placement rates are faster, more openings can be filled, it is less expensive and less prone to the occurrences of compensable incidents.Macassa mining personnel visited the Sudbury operations of Inco in 1993, where development on paste backfill was in progress. As some of the R&D was taking place with funding from the Premier's Council Technology Fund of the Ministry of Trade and Technology of the Province of Ontario , Inco personnel at that time did some feasibility and conceptual consulting on the use of paste fill for the industry. Inco had developed the use of alluvial sand paste fill, an engineered blend of alluvial sand and tailings fines and was doing development work to eliminate the ""conventional"" way of producing paste from mill tailing by thickening and filtration. This latter development work was taking place with GL&V Ontario. That work led to the filing of a joint Inca/ GL&V patent application for the PPSM, the Paste Production and Storage Mechanism with all marketing rights being held by GL&V."

Jan 1, 1997

By Euler De Souza

The authors conducted a study of backfill practices in underground mines across Canada to assess the current state of backfill usage and to overview potential hazards associated with backfill practice, from preparation to transportation to placement. The primary aim of this study was to cover each component of a typical mine backfill system, and to identify potential incidents leading to the disruption of backfilling operations and any associated corrective actions. This paper, Part I of the study, deals with surface operations consisting of backfill material supply and backfill plant preparation. Part II of the study, which deals with underground operations consisting of backfill transportation and backfill placement, is also presented in these proceedings. A review of safety measures associated with current backfill operations and recommendations for procedures of best practice that should be implemented to reduce fill-related hazards in all mine backfill workplace environments are presented. In general it has been found that system failures normally lead only to minor losses to the operations.

May 1, 2009

By M. Kalin

"Ecological Engineering, a biotechnological approach to the decommissioning of base metal mining wastes, is being tested on two pits (Gloryholes) in Newfoundland, Canada. Pit #2 has an average pH of 3.5 and flows into Pit #1 where, through additional ground water contribution, the pH is circumneutral, ranging between 5.5 to 7.The AMD in Pit #2 originates from underground workings and exposed massive sulphide on the pit walls. The overflow from Pit #1 averages about 10 L/s with less than 0.2 mg/L of Cu and 40 mg/L of Zn.Ecological Engineering utilizes organic substrate as a source of carbon and nutrients for microbial populations which can generate alkalinity and reduce metals. In 1989, limnocorrals (enclosures of 4 m diameter and approx 3 m depth) were placed in both pits to test the most suitable organic amendment (peat and sawdust). Microbial alkalinitygeneration was evident 95 days after placement of the amendment on July 4, 1989. The pH had increased after 120 days from 3.5 to 4.5 and by day 480 the pH was 7 in limnocorrals in Pit #2. Zinc concentrations in the limnocorrals in Pit #2, dropped from an average of 35 mg/L to about 2 mg/L or less, by day 480.A scaled-up experiment has been under way since August 25, 1990, where 390 m3 are treated in pit #2 and 750 m3 in Pit #1. One hundred eighty days after placement of the amendment, zinc was reduced in the acid Pit #2 from 40 mg/L to 15 mg/L. Research is underway to determine those factors which limit microbial activity."

Jan 1, 1992

By E. LaRocque

"Long term contracts requiring delivery of 6,000,000 pounds of U308 per year to its customers have made it necessary for Denison Mines to expand its milling capacity from 7,100 tons per day to 15,000 tons per day.Previous papers (1) and (2) described the old leaching and product recovery practice. The flowsheet briefly includes a crushing and grinding circuit, sulphuric acid leaching in a Pachuca plant, liquid-solid separation with wash thickeners and two-stage drum filtration, concentration and purification by moving bed ion exchange followed by two-stage precipitation to produce an ammonia yellowcake.No. 1 Shaft ore, about 4,000 tons per day, will be processed through the truck dump into the old crushing and grinding plant and through the old leaching plant. Ore from Stanrock (No. 3 shaft) will be hauled by truck and also processed through the truck dump as No. 1 shaft ore. Beginning in 1982, ore production from Stanrock will gradually increase until 1985 when some 3,000 tons per day will be milled from the source.A new 8,000 ton per day circuit was designed and constructed to increase the milling capacity to 15,000 tons per day. The new circuit is separate from the old one except that it coins together for the precipitation and tailings neutralization sections of the plant. The feed to the new plant will be from No. 2 shaft.The flowsheet is much the same as the old plant with the exception of the grinding circuit and the ion exchange plant. A semi-autogenous grinding (SAG) plant, unique in the Elliot Lake area, has been installed rather than the conventional crushing and grinding plant; the ion exchange plant is the stationery bed type.This paper describes the major equipment installed for the expanded grinding and hydrometallurgical plant."

Jan 1, 1982

By Dan Alexander, Ivan Crnkovic, Sarah Schwarz

"Several plant studies have been conducted at the Zinifex Century Mine over the past 5 years. For each of the plant studies, surveys of the zinc primary circuit were conducted, characterisation measurements for each operating cell taken, and the floatability components of the ore mapped to develop a model of the circuit. The floatability components and machine parameters were then used to simulate changes in circuit configuration and operating conditions, enabling further potential improvements in circuit performance to be identified.This paper reviews the current flotation practice at the mine, as well as various circuit changes and developments implemented by site. The application of the methodology and the evolution of the standard procedure are also discussed.INTRODUCTIONBase metals were first discovered in the Lawn Hill region of Queensland’s North West in 1869 and the first mining lease in this area was pegged in 1887 over what is now known as the Silver King vein deposit. The area encompassing these veins, which includes the Century deposit, was proclaimed the Burketown Mining Field in 1899. Exploration titles were granted to CRA in 1987, precisely 100 years after the initial discovery at Silver King, hence the name ‘Century’.Century was commissioned by Pasminco in late 1999, with the mine becoming fully operational on 1 March 2000 and the official opening taking place in April. Pasminco had bought the Century and Dugald River deposits from CRA in 1997 for A$345 million, and invested some A$710 million in bringing Century into production. Today, Century is the world’s second-largest zinc mine in terms of its output, having reached full capacity in 2003.The Zinifex Century Limited (“ZCL”) lead-zinc-silver deposit is located approximately 250 kms north-north west of Mt Isa in north-western Queensland. The deposit is a large flat-lying ore body covering an area of some 1.2 by 1.4 km, the 45 m-thick stratabound sulphide orebody is hosted within an 850 m-thick sequence of Proterozoic siltstone, shales and sandstone, amenable to open-cut mining methods. The mine has ramped up to full production delivering five and a half million tonnes of ore per annum, yielding approximately 900,000 tonnes of zinc concentrate and 90,000 tonnes of lead concentrate. Mine life is expected to be about 8 years"

Jan 1, 2008

By H. Wyslouzil

"The Kelsey Centrifugal Jig is a gravity separation device which utilizes the principles of the conventional mineral jigs but with the additional feature of being able to vary the apparent gravitational field. This is accomplished by utilizing the specific induced movements of the conventional jig within a centrifuge.Pilot plant trials have been underway since July 1988 to evaluate the potential applicability of this technology to various process streams at Rio Kemptville Tin.The grade/recovery relationships for samples treated in the jig were consistently better than results obtained from tables. Recoveries of more than 90% were attained at concentrating factors ranging up to 20. When compared with electrostatic separators, the jig achieved equivalent metallurgical results in a two stage process, without the expensive requirement of filtration and drying. Recoveries exceeding 96% were achieved with final concentrate grades averaging higher than 60% tin.INTRODUCTIONMineral jigs have been used for hundreds of years. Many different types have been developed, and each one has numerous variations. In recent years, the idea of applying centrifugal forces to conventional jigs has been put to practical tests and today a number of units are in commercial production. In principle, by enhancing the apparent gravitational field, it should be possible to recover finer particles by overcoming viscosity and surface effects which tend to limit existing gravity devices. In addition, it may be possible to use gravity to separate minerals with much smaller specific gravity differences."

Jan 1, 1990

By Ning Shi

This paper suggests an alternative design approach for an innovative shovel dipper, for use in ground conditions where cutting of virgin ground rather than scooping of blasted material is required. Not only does the new design have an optimized dimension ratio (width:height:depth) and attachment arrangement (angle, position), but also a non-linear pro¯le in 2D and 3D. The approach is based on the simulation of a shovel's duty cycle to verify a pilot dipper model. In order to achieve a realistic simulation that could be used to evaluate a dipper design, the model is e®ectively a descriptor of actual shovel activity under soft ground conditions. The shovel geometry, motion and operation, with an emphasis on digging and dumping kinematics and dynamics are the principal elements of the simulation. The simulation considers a fresh face location, repeating tucking, digging and dumping swing motions within the geometric range of the machine, ending when no more face material remaining within reach without re- positioning. The simulation yields digging force, friction force, trajectory, ¯ll time, ¯ll factor, etc, which may be integrated into two basic considerations: power draw (output) and volume delivered (production), as indicators of dipper performance. The simulation has been used as a tool to modify an existing dipper design, prior to looping further simulation cycles. Ultimately, an optimized or totally-new-concept shape is realized. A prototype dipper as a fractional scale model will be built to verity the performance under actual mining conditions.

May 1, 2004

By M. Ounpuu

"Gold, because of low head grade and recovery, is usually of minor importance in the economics of base metal flotation mills. Because of the above little metallurgical attention is usually provided to gold in these mills. But collectively, the gold losses can be significant. For instance, Noranda presently has interest in 13 base metal flotation mills with a collective gold recovery of only 50%. Gold losses are in the order of 150.000 oz per year!The Noranda Technology Centre, in cooperation with Noranda's milling operations, has been investigating the potential of gravity concentration of gold from base metal flotation mills. This paper outlines the pilot and plant testing conducted at three operations with distinctly different ore types; porphyry copper, complex massive sulphide and polymetallic semi-massive sulphide. Testing has been with the Knelson Concentrator which has proven to successfully recover fine gold in a simple, trouble free circuit. Discussion of issues which have risen from the testwork is included. INTRODUCTIONGold, because of low head grade and recovery, is usually of minor importance in the economics of poly-metallic base metal ores. TABLE I presents gold heads and recovery for some of Noranda's base metal mills. It should be noted that the stated recoveries are to payable concentrates, but does not reflect the actual payable gold recovery which in cases is significantly lower. Noranda's average gold recovery into a gold payable concentrates is about 45%. This equates to a collective loss in the order of $57 MM per year of gold"

Jan 1, 1992

By Fred Delabbio

There are presently several issues that the mining industry is facing, such as underground air quality, deep mining and green house gas emissions, which could be addressed by the application of alternate energies. Several aspects must be considered by the industry in any drive to introduce new technologies within the context of a producing underground mine. This article presents several operational and support aspects that will evaluate the relevance of alternate energies, with emphasis on hydrogen power systems: operational drivers, commercial supply of technology, site supply system requirements, health and safety, capital and operating costs, mine production, corporate philosophy, and role of government. It is concluded that introduction of alternate energy technology requires a step-wise progression: Applicability to the industry, proof of concept testing, development of generic infrastructure, power systems and regulations, whole system studies and detailed studies of cost and risk versus benefits. The current initiative to apply fuel cells is at the proof of concept stage, entering the generic infrastructure stage.

May 1, 2007

By M. Rebolledo, N. Mora, G. Zarate

For decades, the mining industry has been focussed to find an efficient and economically attractive process for the leaching of chalcopyrite. Jetti Resources has developed a novel hydrometallurgical alternative to extract metals of value from mineral ores that the mining industry has struggled to process. Jetti’s catalytic technology, whilst being especially advantaged for extracting copper from low-grade and highly refractory minerals, such as chalcopyrite, seamlessly integrates with existing heap leaching methods and downstream processing operations. The technology uses a catalyst which weakens the metal-sulfide bond in the crystalline structure of the chalcopyrite mineral and alters the structure of the passivation layer allowing the leaching process to continue. Metallurgical test work was conducted to examine the effect of ore variability, in terms of grade and copper deportment. The results presented in this study highlight the amenability of the technology and demonstrate that it is possible to enhance leaching performance of copper primary sulfides over a wide range of copper grades, from 0.1 to 2.2%. Further, it is concluded that despite the copper grade, the tests with increasing chalcopyrite content render copper recoveries up to 4 times higher compared to the tests under non-catalyzed conditions. Jetti’s catalytic technology is currently being evaluated at larger scales, namely a commercial dump leach operation and 50-tonne crib tests. The results obtained thus far are compared to the results from conventional sulfide heap leaching tests. INTRODUCTION For decades, the mining industry has been focussed to find an efficient and economically attractive process for the leaching of chalcopyrite. Chalcopyrite (CuFeS2) is the most abundant and recalcitrant of the copper sulfides (Singer, 2017). Owing to its abundance and economic importance, it is the most demandable natural source for copper worldwide. In fact, the projected demand for copper is further expected to rise much higher beyond the current demand by the year 2020 (Panda et al., 2015). Therefore, it becomes imperative to find low intensity hydrometallurgical technologies for the extraction of copper from chalcopyrite in ores of such low-grades that are un-economic to concentrate and to augment the supply by extracting copper from ores with complex mineralogy that cannot be concentrated (Watling, 2013).

Jan 1, 2019

By Dwight Newman

This paper discusses the author?s work identifying international law norms of consultation with indigenous communities and poses a potentially counterintuitive argument that corporate stakeholders? interactions with indigenous communities may have effects on the content of those norms. The claim is counterintuitive because the international law of pertinence is customary international law, which arises from consistent state practice and opinio juris (a belief by states that they are acting in a way they are legally bound to do so). Customary international law traditionally has no reference to corporate conduct per se. The paper first surveys several sources on developing international law on consultation obligations. Then it references the impact these norms can have on corporate stakeholders in industries like mining. Then, it poses an argument as to an economic ratcheting effect that can occur through various agreements between corporate stakeholders and indigenous communities that then feeds through into international law via several mechanisms. The consequences for mining companies are diverse and include the possibility that conduct by one company can have longer-term implications for the obligations of other companies and, indeed, of states. The effects will differ in the context of different competitive environments and may actually have effects on the economics of different competitive models.

Aug 1, 2013

By K. Yamaguchi, Sekimoto. H.

"Platinum Group Metals (PGMs) used as catalyst which reduces toxic substances in emission gas exhausted from an automobile should extensively be recycled because they are rare metals produced only at the limited regions. The recycling process is to smelt an alumina honeycomb, in which PGM is built, in molten iron with flux such as lime and silica to separate them. PGMs are concentrated in molten iron while alumina in slag.The objective of this study is to understand the factor to concentrate PGMs in liquid iron as much as possible. Fundamental research of the distribution of the PGMs between the slag and liquid iron was carried out to recover the value metals of platinum, palladium and rhodium from the three-way auto catalyst. The distribution ratios of platinum, palladium and rhodium between the Al2O3-CaO-SiO2 slag and liquid iron are from 10-6 to 10-5, and these are concentrated in comparison with slag phase by about 1 million times to the liquid iron at 1873 K. The distribution ratios of platinum between the FexO-CaO-SiO2 slag and liquid iron decrease from 0.1 to 2?10-4 with increasing the CaO content in the slag. The CaO addition to the FexO unary slag is effective in concentrating platinum in molten iron.Based on the distribution ratios obtained in this work, recovery of PGMs was determined. The recovery of platinum, palladium and rhodium increase with increasing the mass ratio, Mmetal/Mslag. In the case of the conditions of Mmetal/Mslag = 0.04, the recoveries of PGMs showed more than 99.97%. The use of liquid iron and carbon saturated liquid iron as the collector metal is advantageous for the PGM recycling of the three-way auto catalyst. We found that PGMs are enriched to the molten iron or the carbon saturated iron and the slag loss of PGMs is negligible small."

Jan 1, 2014

By G. Deschênes

The Eleonore property (James Bay district of Northern Québec), owned by Goldcorp, hosts gold within stockworks of quartz-tourmaline-arsenopyrite veins and veinlets contained within microcline (potassic alteration) and brown tourmaline replacement zones. The preliminary investigation to recover gold from drill core samples showed that flotation followed ultrafine grinding and cyanidation is the preferred option. Ultrafine grinding is used to liberate gold finely disseminated in sulphides. The flotation concentrate used in the investigation had a P80 of 11 microns and contained 75.5 g/t Au, 5.0 g/t Ag, 10.3% S, 0.07% Cu and traces of antimony (0.03%). The mineralogical characterization showed the presence of 65% gangue minerals, 23% pyrrhotite (hexagonal), 9.6% arsenopyrite, 2.2% pyrite and 0.08% chalcopyrite with trace amounts of galena (0.05%) and sphalerite (0.03%). Gold was in the form of native gold and electrum. A straight cyanidation conducted at pH 11.0, DO of 3-5 ppm, 2000 ppm NaCN, 20 C 72 hours extracted 97.0% Au and consumed 31.5 kg/t NaCN. The equilibrium was reached within 24 hours. Increasing the dissolved oxygen to 16 ppm reduced the consumption of cyanide to 22.0 kg/t with 96.7% Au extraction. A 16-h pretreatment with 2.0 kg/t lead nitrate, followed by a 24-h cyanidation reduced the cyanide consumption to 6.0 kg/t with a gold extraction of 97.5%. Increasing the lead nitrate addition to 6.0 kg/t reduced the cyanide consumption to 1.2 kg/t by passivating the surface of pyrrhotite. The gold extraction remained at 97.0%. It was possible to decrease the concentration of cyanide to 800 ppm NaCN without compromising gold extraction. The associated cyanide consumption was 1.2 kg/t NaCN. The passivation of gold grains was an issue when the cyanide concentration was decreased to 550 ppm NaCN. The gold extraction decreased to 92.2%. The average lime addition for all the tests was 7.4 kg/t and the average silver extraction was 80%. Efficient cyanidation of an ultrafine sulphide concentrate with only 800 ppm NaCN represents an advance in leaching of ultrafine sulphide concentrates.

Jan 1, 2013