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By B. Roy

The Somita plant processes low sulphide gold ore in Burkina Faso with a CIL process. After experiencing recurrent low gold recoveries, Somita undertook an investigation on the source of the problem. COREM conducted a mineralogical study on different pit samples. A series of tests was also conducted on a composite sample from the plant which shows a content of 3.4 g/t in gold and 0.48 % in graphite. A reference test in laboratory resulted in a recovery of gold of 75%, similar to plant recoveries for the evaluated ore. In these tests, even if cyanide is not deficient, it was diverted mainly into iron cyanide complex, copper cyanide complexes and by graphite adsorption. Causes of low recovery were also identified and quantified. Solutions to increase gold recovery were suggested such as increasing gold dissolution kinetics by O2 or by lead nitrate addition, change in activated carbon addition point location to the 1st cyanidation tank and finer grinding to increase liberation. Increase in the pH could also contribute to reduce cyanidation consumption. Following these recommendations, Somita changed carbon addition from tank 2 to tank 1 in the CIL circuit. This change led to an increase of more than 5% in gold recovery. A system for the addition of lead nitrate is also under implementation at the plant and benefit will be evaluated in 2013. Finally, pit management for cyanidation will be done based on mineralogy instead of gold content in the future.

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

By A. Himsley

"There is an ever-increasing demand these days for efficient methods of removing metals from solutions especially for selectively recovering a particular metal from a mixture of metals in solution.Up until the mid l 930's the only methods widely practiced for such separations involved precipitation techniques or electrolysis. Both these methods have limitations and indeed can only be used for certain metals.It is only within the last 40 years or so when synthetic ion exchange materials became available that chemists in industry and universities have been able to apply ion exchange techniques to a wide variety of separations that not only would be impossible by precipitation methods but which are now achieved with much greater efficiency.Ion Exchange TechniquesIon exchange is a unit process that can be applied using either liquid ion exchangers or solid resin particles. Liquid ion exchange - or solvent extraction as it is often referred to - is used in a number of special areas dictated mainly by the nature of the separation.On the other hand, ion exchange resins are much more widely used and consequently a great deal more is known about their application to a variety of metal separations.In both systems there is a loss of the ion exchange material which in the case of solvent extraction results in hydrocarbons going to waste. The ion exchange resin systems lose resin in the form of small particles, referred to as ""resin fines"", and these are insoluble.The loss of liquid ion exchanger is due to the fact that it is slightly soluble in water and consequently the loss is a function of the volume of liquid with which it comes into contact and is not related to the amount of metal processed."

Jan 1, 1975

By H. R. Bullis

"Abstract- This paper discusses what appears to be a flaw in the current belief, widely held within the gold mining industry and investment community, that the creation of ever-larger gold mining companies is desirable for mining companies and investors alike. The basis of the flaw is in the underlying assumption that very large gold producers can discover or acquire new gold deposits of the size necessary to replace extracted reserves on a year-on-year basis over the intermediate to long term (i.e., five to ten years and longer). As it will be noted in this paper, deposits of the size required to replace annual production of very large gold producers (VLGPs) are relatively few in number. Although discoveries of new gold deposits continue to be made, they are predominantly in the 0.5 to 2.0 million ounce range. Data represented here suggest that, although gold production from individual VLGPs continues to increase, the change is due to merging or acquiring other companies or projects with similar reserve life profiles. Therefore, the reserve life profiles of VLGPs have remained flat or have decreased.The paper will note recent changes in the industry and the remarkable increase in gold production reported by individual mining companies over the past several years. It will provide an overview of the geographical and geological distribution of known gold deposits and the range of deposit sizes. Finally, the paper will discuss the challenges that face large gold producers in replenishing their reserves that are being depleted at an ever-increasing rate. Unless otherwise stated, all production and reserve data have been taken from company documents in the public domain or publicly available information."

Jan 1, 2001

By Jim Wickens, Andy Cole, John Pekrul

"The roaster facility at Barrick Goldstrike Mine treats double-refractory sulfidic/carbonaceous gold ore. Prior to oxidation in the roasters, the ore is ground in two parallel dry grinding circuits utilizing 19’ Ø x 75’ double rotator mills. The design plant throughput was 12,000 dstpd. However, process improvements have significantly increased throughput rates by 25-50%. The current daily throughput record is 20,078 dstpd. Grinding circuit improvements contributing to this success include feed size reduction, ball size optimization, automatic ball charging optimization, and operating strategy changes. In addition to the changes already made, this paper discusses future optimization initiatives including on-stream particle size analysis and “real-time” ore hardness determination techniques.INTRODUCTIONBarrick Goldstrike Mines is located 29 miles Northeast of Carlin, Nevada, USA. Through 2001, the property has produced over 23 million ounces of gold making it one of the premier gold operations in the world.Prior to Barrick’s ownership, the property began producing gold from a small heap leach operation in 1975. The small-scale heap leach production continued through 1986, and in January 1987, Barrick acquired the property. Increased exploration activity by Barrick significantly increased gold reserves. In August 1988, a 4,500 dstpd mill facility was commissioned to treat high grade oxide ore. Also, heap leach operations had significantly expanded. In 1990, the mill was expanded to 6,000 dstpd, and the first autoclave was commissioned to treat refractory sulfide ore. By the end of 1996, the mill and pressure oxidation facility had expanded to 16,000 dstpd and heap leaching had ceased as oxide reserves were exhausted. In 2000, a roaster facility was commissioned to treat double-refractory sulfidic/carbonaceous ore."

Jan 1, 2003

By F. Lindroos

A cornerstone in the LKAB Green Pellet concept is to produce pellets with a small footprint in terms of energy con-sumption, CO2 emission and emission of other elements harmful to the environment, like dust, sulphur dioxides, nitrogen oxides, chloride and fluorine. The concept also encompasses our determination to make pellet products that contribute to low environmental impact and low energy consumption in our customer?s processes. The presentation will focus on the pelletizing process. It will describe both how we have attempted to decrease the generation of unwanted components in the process gas, and how we have tested or evaluated equipment for external gas cleaning. We will discuss and give examples from recent and on-going efforts and outline some of the future challenges we see, to meet our own environmental standards and also the tough demands imposed by EU and national legislation, while striving to boost production.

May 1, 2011

By Joe Holmes, Peter Langlois

"Falconbridge’s Raglan Concentrator, located at the northern limit of Quebec’s Nunavik Region, has been in operation since December 1997. Originally designed to process 800,000 tonnes per year of high grade nickel ore, plant capacity is currently approaching 1,000,000 tonnes per year. Optimization of the ABC grinding circuit has proven to be the most challenging component of this capacity increase. Concurrent flotation circuit modifications have allowed concentrate grades to increase from 16 to 18 percent nickel, while maintaining recoveries. Concentrate drying problems encountered during the first 15 months of operation have been addressed through equipment upgrades. The challenges associated with surface deposition of filtered tailings in an arctic environment are now better understood, and detailed design of the progressive reclamation cover is underway.GRINDINGCircuit OverviewOre from open pit and underground mining operations is introduced to the underground ore handling system via surface and underground truck dumps. Two underground bins with a combined capacity of 4000 tonnes provide roughly a day of live capacity for ore storage. Ore from the two bins is conveyed to a 150 tonne surge bin immediately ahead of the grinding circuit in the concentrator.The Raglan grinding circuit has a conventional ABC configuration (Figure 1) with a design capacity of 100 mtph (800,000 tonnes per year). The 24’ AG mill is powered by a 3000 hp motor with variable speed drive. AG discharge is sized on a 8’x16’ single deck vibrating screen, with screen oversize recirculating to a 5 %’ short head cone crusher (400 hp) before returning to the AG mill. Screen undersize is combined with ball mill discharge to feed up to 5 15” Krebs hydrocyclones. The ball mill is a 14’x 21’ overflow mill driven by a 3000 hp motor at 73% critical speed. Target Pgo for the grinding circuit product is 65 microns"

Jan 1, 2001

By T. Yalcin

"This paper describes the development of a mathematical model for determining the liberation of an ore at various grinding sizes, based on the combined results of microscopic counting, grinding tests, and chemical analysis.The method is illustrated with an application to the magnetite ore of Sherman Mine in Temagomi, Ontario. For this ore, a degree of liberation of 98.7% was determined at a grinding size of 90% passing 44 microns, which closely corresponds to the final grinding size employed at the Sherman Mine Concentrator.Introduction Liberation characteristic of ores have direct and significant bearing on overall economics of concentration processes.Target grades and recoveries can be achieved only it minerals are adequately liberated prior to separation. Degree of liberation depends on the extent of size reduction. However, due to the high costs associated with size reduction processes, it is important to keep size reductions within limits of satisfactory liberation. Prevention of excessive size reduction is also necessary because of inefficient mineral separations in very fine sizes. Hence, a study of reduction size — liberation relationship is essential for optimum profit from milling operations.A number of procedures have been developed for liberation studies (1-12), but there is lack of method for universal adoption.A new approach is presented in this paper that might provide a realistic estimate of liberation and give a practical guidance as to the degree of size reduction required for satisfactory liberation."

Jan 1, 1985

By J. K. Gowans

"INTRODUCTIONTo bring a mining property into production in the Canadian High Arctic a company must deal with the harsh environment, the short shipping season and the high cost of labour. In trying to minimize the effect of these factors, the concept of a barge-mounted process plant was investigated as an alternative to the conventionally built concentrator. Cominco's previous experience gained at the Black Angel Mine on Greenland's west coast suggested that this would be a practical alternative.The normal shipping season at Little Cornwallis Island is approximately six weeks long, from mid-August to the end of September. All equipment and materials must be delivered during this small ''shipping window."" The logistics problems are enormous; if the conventional approach had been used the major process equipment would have been delivered to Little Cornwallis Island at the same time as the completed barge actually was! Construction would also be starting at the start of the worst climatic conditions, the Arctic winter. Nine months production was gained by opting for the ""Barge"" approach.What the Arctic winters do to the shipping season, so do they affect the productivity and the cost of labour. Construction labour must come from southern Canada. Thus higher transportation costs and higher wage rates, with more overtime and incentives, increase the labour costs to approximately 50% over normal southern rates. Couple these rates with the lower productivity involved with men working in the extreme cold temperatures and the net effort is worth avoiding. A process plant being constructed in Southern Canada can take advantage of more accessible labour at lower costs, increased productivity and a faster construction schedule. The barge concept must still deliver the equipment and materials during the ""shipping window"" but it can in the completed form."

Jan 1, 1983

By R. Gallinger, B. Kelley

"In 1997, a group of Canadian mining companies met to review the Mine Environment Neutral Drainage (MEND) program and to consider future activities. They concluded that (1) that Canadian industry and Natural Resources Canada (NRCan) should fund a small secretariat in Ottawa to support technology transfer in Canada and, (2) that it was important for industry to continue the momentum of MEND on an international scale.The need for the International Network for Acid Prevention (INAP) was strongly endorsed by consultants, company, university, and government experts from Australia, Canada, and the United States who convened in Vancouver in February 1998. The initial focus was the development of a solid knowledge base covering the collective experience of INAP’s members to promote the dissemination of successful practices, the identification of knowledge gaps, and the formulation of effective research proposals. The sharing of information on past successes and failures alone was viewed assufficient justification for the formation of INAP.INAP was officially launched in October 1998, and is now an effective, industry-based initiative to globally coordinate research and development in the management of sulphide mine waste. Significant improvements in the management of sulphidic mine materials through knowledge sharing, research, and development of technology by INAP will reduce liability associated with acid drainage.This paper will explore the international efforts and activities of INAP."

Jan 1, 2005

By Scott Martin

This paper will discuss the challenges currently facing the mining industry. By sharing experiences gained in the CIM Leading in Mining Course, I will highlight what’s working, what’s not, what are the challenges, what are the concerns, what are the opportunities, what needs to happen. Specifically, I will discuss what is currently happening with leadership and leadership development in members’ own organizations, what skill sets are important to develop in different hierarchical groups and how comprehensive leadership courses like the leading in Mining course can help improve leadership within individuals and the industry. Leadership development is a process not an event.

Jan 1, 2014

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 D. Hatfield

"Flotation plants typically aim to maximise recovery at or above a target grade. As the ore from the mine varies, operators increase or decrease the mass pull in different flotation banks by adjusting air and level set points in order to achieve target concentrate grade. A methodology has been developed to maximise overall recovery at target grade by determining the optimal mass pull for each bank. However, the approach of targeting equal grade for each ore type in the deposit leads to sub-optimal overall recovery. This is because some ore types sacrifice a large amount of recovery to reach target grade while others easily achieve target grade and can be processed at higher than target grade for little or no loss in recovery. Processing different ore types at different grades and blending the concentrates therefore gives a higher overall recovery.This paper presents a case study where an 80 year production forecast was made for a large Chilean copper concentrator in feasibility stage. A circuit model was created using SGS’s Integrated Geometallurgical Simulator (IGS) for the different ore types to be processed. The model was calibrated from standard rougher flotation tests (MFT), cleaner tests and from the results of plant surveys of similar operations. The pull rates for each bank in the circuit were determined to achieve an optimal graderecovery curve for each ore type. These curves were combined at different concentrate grades for each ore type to obtain the mathematically optimal overall recovery at target grade. The paper shows a methodology for obtaining an optimal combination of operating conditions for each ore type to maximise overall recovery at a set combined concentrate grade."

Jan 1, 2013

By A. R. Chambers

Malagash received its name from the Indians many years ago on acco.unt of its turbid (milky) waters, although,' the red men did not appreciate the significance of this turbidity, and were, of course, entirely ignorant of the crystal beds of pure salt that lay beneath their feet. The salt lake of the Mississippian, or Lower Carboniferous period, in which these beds occur, was apparently a large one, for there is evidence of its existence at intervals all the way from Sydney, Cape Breton, to St. John, New Brunswick, 300 miles as the crow flies, and there is no knowing what its other dimensions may have been, as much of this horizon is now buried on the one side beneath the waters of the Gulf of St. Lawrence, and on the other underneath the Atlantic ocean. About midway of the area and underlying the counties of Colchester and Cumberland, there occurs a synclinal basin of which the Malagash salt outcrop forms only a small portion of its northern limit. This basin has been described very fully by A. C. Hayes. (1) The operations carried on at Malagash since Hayes wrote his report have brought to light very little in the way of new geological data, except to confirm his pre-dictions as to the probable faulting of the salt strata, and the effect which these faults would have on mining operations.

Jan 1, 1924

By Pierre Pelletier, Julie Fournier, Gianni Bartolacci, Yves Breau, Dominique Girard, Ahmed Bouajila, Carl Duchesne

"The development of on-stream analyzers (OSA), such as those based on X-ray fluorescence (XRF), has been key to improve metallurgical performance of flotation processes. However, since these analyzers are shared between several streams, they typically provide grade measurements every about 10-15 minutes. In general, this frequency is too long compared to process dynamics and so, implementing an efficient feedback control strategy based on OSA can be difficult. In addition, due to the limited capacity of OSA systems, several important intermediate streams are not analysed, but could provide relevant information about the state of a flotation circuit.To overcome this problem, vision systems have been developed in the past to predict concentrate grade based on froth color – this could be very useful to control flotation performance between OSA measurements. However, applications reported in the literature typically use univariate summary statistics of the full red, blue, and green color distribution of froth images, while RGB images are truly multivariate. Although, in some cases, grade prediction was achieved with reasonable accuracy, one may, by using only summary statistics, miss important information about the state of the flotation process.This paper investigates the application of Multivariate Image Analysis (MIA) concepts to flotation froth images. These techniques are based on multivariate statistical methods such as PCA and PLS, which have the ability to analyze large databases (such as digital images) and summarize the relevant information content of the data into a small number of latent variables, capturing the most important features of the images for concentrate grade prediction. The potential of application of these methods in the mineral processing industry is illustrated using a few industrial case studies."

Jan 1, 2003

By K. H. B. Wenzl

"Since the autumn of 1973, the world's attention was captured by the startling revelation that conventional energy supplies were not inexhaustible but were in fact quickly dwindling with future shortages imminent until lagging exploration and technology yielded additional energy sources to catch up with the spiralling increase in demand. This situation was precipitated by the Arab Oil Embargo and the resulting soaring cost of world crude oil.During the past few years there has been tremendous activity in the energy field which has had a direct impact on the Inca Metals Company Ontario Division operations. The company's reaction has been to establish an energy management program which will be described later in the paper. First, a brief description of the Inca Metals Company and the primary energy resources which are utilized will assist in putting the program into perspective.The Inca Metals Company, a unit of Inca Limited, is responsible for the production and sale of primary metals products worldwide. The company's operations in the Sudbury District include mines, ore concentrating plants, a Smelter, Nickel and Copper Refineries and an Iron Ore Recovery Plant. The Ontario Division also has a mine and concentrating plant in Shebandowan near Thunder Bay, and a metals refinery at Port Colborne. Inca is the largest producer of nickel in the non-communist world and is also a major producer of copper and platnium group metals."

Jan 1, 1978

By Peter Wells, Adrian Sabau, Turgut Yalcin

"The nickel ores of Sudbury contain substantial amounts of the gangue mineral pyrrhotite (Fe7S&) which causes certain diflculties in the milling of these ores. This paper proposes a process called “magnetoflotation” as an effective means of dealing with pyrrhotite. The method reduces the pyrrhotite content of concentrates by applying a magneticjield on the path product as it leaves the flotation cell.Successful performance of magnetoflotation requires the optimization of several variables. In this study, three of the variables (air flowrate, collector concentration, pother concentration) were investigated by Box- Wilson experimental design, and others (magnet speed, grind size, pulp density, flotation time) were examined individually. The process was tested on a Falconbridge nickel ore, assaying 1.92% Ni, and under the best operating conditions, a concentrate was obtained with 13.56% Ni grade and 74.32% recovery in a single step. Further improvements to grade and recovery were possible by cleaning and scavenging. The same ore was also tested with conventional methods for comparison, and results indicated that magnetoflotation performed better.INTRODUCTION“Magnetoflotation” refers to the process of carrying out magnetic separation and flotation simultaneously within the same device, and is intended to prevent magnetic minerals from reporting to the froth products while non-magnetic minerals are being floated. Typical cases where this method might be beneficial include the processing of magnetite and nickel ores. Its use on magnetite ores has been reported elsewhere [Yalcin, 1992; Yalcin, 19931. This paper describes its application to nickel ores, more specifically to the nickel ore of Falconbridge Limited in Sudbury."

Jan 1, 1999

By Samantha Beeson, Collin Poole

"This paper examines the development of two new selective flocculation techniques for recovering KCI values from low grade carnallite potash ore.Direct recovery of KCI was achieved using one of two mid-range molecular weight polymers, one non-ionic and the other strongly cationic, in conjunction with the same selective anionic dispersant. From initial response studies on pure minerals and selective flocculation tests on synthetic mixtures, investigations were extended to recover KCI values from a natural carnallite potash ore. A single-stage process achieved products grading up to 61.3% KCI at recoveries up to 85.4% for the cationic flocculant suite and up to 64.8% KCI with recoveries as high as 90.3% for the non-ionic flocculant suite, with dispersant and flocculant dosages of 150-250gr1 and 10-75gr1, respectively. These results clearly demonstrate not only the feasibility of treating such problematic ores at low reagent dosages, but also the potential elimination of desliming prior to froth flotation.The experimental results presented are considered in terms of the reagent adsorption phenomena present in the system and various aspects of floe formation and cleaning.IntroductionAs high-grade sylvinite potash ores become scarce, lower grade sylvinite and carnallite potash ores are becoming more attractive. However, these ores often contain large amounts (up to 15% by weight) of water-insoluble slimes which must be removed prior to KC! flotation, owing to their high adsorptive capacity for the amine collectors used. The slimes are generally removed by scrubbing followed by hydroclassification and any residual material not eliminated is blinded with reagents such as guar gum or starch to inhibit interference in subsequent KC1 flotation.An alternative approach used commercially in Canada by Agrium Inc. involves water-insoluble slimes flotation prior to KC! flotation. In this process, the slimes are selectively flocculated and then separated by froth flotation. Losses of KC! to the slimes product and associated brine requirements are both reported to be less than those encountered in conventional desliming, but the technique is somewhat limited in its application at present."

Jan 1, 1997

By Henri F. Letient

At the 25th Annual B.C. Mine Reclamation Symposium in September 2001, Huckleberry Mines Limited received the Jake MacDonald Reclamation Award for its exemplary mine waste management plan, conceived to prevent the onset of acid mine drainage and facilitate reclamation of the site. The mine is an excellent example of a total mine waste management plan, with all materials scheduled in detail for the life of the mine. The filling plan for the waste management facility takes into consideration the requirement for permanently flooding the potentially acid generating rock. From an acid generation point of view, the tailings fall in the zone of potentially acid generating material. To avoid long-term ARD concerns, it is planned to float pyrite out of the tailings towards the end of the mine life, thereby generating a non-acid generating tailings stream which will be used to cap the deposited mine waste. The pyrite-rich tailings will be discharged into the deeper portion of the impoundment. On closure, a permanent water cover will be maintained over the mine waste. Only beaches of non-acid generating tailings will remain exposed. Mining will be phased to also allow one of the pits to provide storage for potentially acid generating rock and to be flooded at the end of mine life.

May 1, 2002

By L. C. St. Arnaud

"Open pit mines can produce large disturbances within regional hydrogeologic systems. The degree of disturbance and the consequences on the local environment depend on the size of the pit and on the physical setting of the mine. Comparisons and contrasts are presented for three Noranda base metal open pit mine sites.The Brunswick Mining and Smelting No 6 open pit has a surface area of 11 ha and a depth of 185 m. It is located within complexly deformed schists and meta-volcanic rocks. The massive sulphide and iron formations visible along the north and southwest walls of the pit generate acid drainage. Mining ceased in 1983, and the pit is presently being filled with waste rock and neutralized drainage water. Water balance calculations estimated a total fill up time of sixteen years. As the water level rises, leakage to groundwater could take place across a fault zone intersecting the pit.The Mattabi Mines F-Group pit has a surface area of approximately 3 ha and a depth of 50 m and is also set in meta-volcanic rocks and acid generating iron formations. Production at this site ceased in 1982, and the low pH pit water is neutralized and pumped into a nearby lake at least every five years. Major water inflows are from precipitation and groundwater. Groundwater mounds within adjacent high land surrounding the pit provide a natural hydraulic barrier to groundwater flow, except for the southeast side. As a consequence, the pit water elevation is being kept below the elevation of the nearby creek in which some of the pit water could drain.The Bell Copper open pit covers an area of 40 ha and is 270 m deep. With present ore reserves, the mine might be in operation until 1992. The geology around the pit consists of andesitic tufts, siltstones and shales, and the acid generation potential of the rock is relatively low. A recent hydrogeological study estimated fill up time by precipitation and groundwater to be in the order of 260 years.With increasingly more stringent requirements for mine closure, it is anticipated that more detailed hydrological and hydrogeological studies of open pits will be required to determine post-closure environmental impacts and the suitability of open pits for mine waste disposal."

Jan 1, 1990