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"Automatic flotation control in varying degrees of sophistication is practiced at a number of concentrators throughout the world. Practically, the variable conditions in the flotation circuit and the difficulty in maintaining accurate instrumentation information all add to the complexity of automatic control.Further development at the Kidd Creek concentrator has resulted in a 3075 mtpd copper flotation circuit capable of running twenty-four hours a day under the direct control of an optimizing supervisory computer.Three main streams, continuously monitored by an XRF analyser, supply circuit information to the optimizing computer. Circuit efficiency calculations (from smelter contract terms) and target setpoints allow the computer to determine the economic optimum direction to move the circuit based on the grade/recovery relationship. Circuit manipulation is accomplished by precise pumping of collector and frother reaqents together with accurate control of flotation bank pulp levels. The former uses computer controlled diaphragm pumps and the latter is achieved by control loops utilizing a float type level transmitter, variable position dart plugs and a process computer working to setpoints determined by the optimizing computer.The development of a fully automated flotation circuit is by no means complete, but increased knowledge and major hardware improvements have allowed its progression to a higher level of sophistication."

Jan 1, 1983

By Gerald Whittle

What is Enterprise Optimization? Releasing value through better planning decisions: Philosophy, Methodology, Software The key to 5% -35% increase in enterprise value (NPV) The culmination of 25 years development in optimization for the mining industry ?Jeff Whittle Simultaneously optimizing: ?all steps in the value chain ?all assets in the enterprise portfolio ?all periods together

May 1, 2010

By M. N. Brodie

"The ability to adjust the speed of a SAG mill offers benefits to the process and to equipment maintenance. It can be provided with some slight effect on availability ranging from somewhat enhanced to somewhat reduced, depending on the particular equipment that is used. Several techniques have been developed for monitoring the magnitude of the contained charge, and several algorithms have been used to control mill feed rate and mill speed. The various benefits and the means by which they can be realized are examined and compared. Reasons for Variable SpeedProcess RequirementsThe recovery of minerals from ore is dependent on particle size and residence time in process. Separation in a given plant is most efficient at a particular rate and a particular size (e.g., 80% minus 200 mesh). If the particle size increases, less mineral will be liberated. If the particle size is significantly reduced, the process of choice may not be able to effect separation, and power, media, and maintenance costs would be increased.If a SAG mill-ball mill combination delivers the desired throughput at the correct size for efficient separation when processing ore of the design hardness, some adjustment must be made to maintain this condition when the ore changes. The processing of ore softer than design will result in overgrinding, unless:• the throughput is increased inversely to the change in hardness which lowers recovery by overloading the separation facilities;• or the SAG mill feed rate remains constant, but the speed is reduced to feed coarser material to the ball mill for finish grinding to the desired size.There is also the effect of changing milldiameter to consider. A nominal 36 ft diameter SAG mill will have an effective diameter of 35 ft with new liners and 35 ft 6 in. with worn liners. A mill designed to operate at 0.76 p.u. (per unit) of critical speed with new liners (9.84 rpm) will be changed to 0.765 p.u. with worn liners. If 0.76 p.u. is the desired value, the mill speed should be reduced to 9.77 rpm."

Jan 1, 2003

By D. E. Calisaya, É. Poulin, A. Desbiens, R. del Villar, A. Riquelme

"A 15 cm diameter by 732 cm height, fully automated pilot column was installed in the Agnico-Eagle’s Laronde concentrator (Québec), for evaluating the performance of newly developed sensors and advanced process control algorithms. For eventual comparison purposes, the pilot column is continuously fed with the same slurry feeding the copper-circuit third cleaner column. It is equipped with electrical conductivity-based sensors for estimating the three variables used in the multi-variable control algorithm under evaluation: froth depth, fraction of wash water underneath the interface and gas hold-up. Froth depth is regulated using a PI controller while a predictive controller is formulated as the minimization of the tracking errors of gas hold-up and bias rate, keeping several operating constraints between their practical upper and lower limits. In particular, an unreachable high set point of gas hold-up can be selected in order to maximize the bubble surface area available for particle collection, while maintaining bias rate above a minimum value required for froth cleaning. The proposed strategy may indirectly help optimizing flotation column operation by using these secondary variables, without the implementation of a real-time optimization layer.INTRODUCTION Flotation columns have already been used for more than 30 years, mainly in cleaning stages, as a result of their froth washing capabilities. Despite their widespread usage, the measurement of some pertaining variables and their use for automatic control is rather scarce. Among the most important variables directly related to the process metallurgy (concentrate grade and recovery), froth depth (hf) is the only one normally monitored and controlled. Even though commercial sensors are available, gas hold-up (eg), a variable directly related to concentrate recovery, is seldom measured and never controlled. Another important variable, bias rate (Jb) responsible for concentrate cleaning, is simply not measured at all, as there is no commercial sensor available. Therefore, besides froth depth, continuous monitoring of process variables is limited to flow rates (gas, feed, tailings, wash water), i.e. to manipulated variables. Some new available sensors have been industrially tested, e.g. the Sonar-TracTM for eg (O'Keefe, Viega & Fernald, 2007) and the McGill hold-up sensor (Tavera & Escudero, 2002). Others, for bias rate and frother concentration, have been successfully validated at laboratory scale by Laval’s LOOP researchers (Maldonado, Desbiens, del Villar & Chirinos, 2008b; Maldonado, Desbiens, del Villar & Aguilera, 2010a). Some devices have been extensively used for monitoring bubble size and space distri-bution, e.g. McGill bubble viewer (Gomez & Finch, 2007) and others."

Jan 1, 2012

By Yves-Laurent Turcotte

Notre conférencier, Monsieur Yves L. Turcotte, Vice-président et Directeur général chez Promaintech Novaxa est diplômé en génie industriel de l?Université du Québec à Trois-Rivières et détient une maîtrise en gestion de projets de l?Université du Québec à Montréal. Il est une véritable sommité canadienne dans le domaine de l?amélioration de la performance des organisations. Provenant d'entreprises manufacturières et de services, il possède plus de 20 ans d?expérience en tant que développeur de technologies, conseiller et formateur auprès d?entreprises. Il a à son actif la réalisation de mandats dans plus de 400 entreprises. Il est reconnu pour sa grande capacité d?analyse et pour trouver des solutions novatrices et profitables aux problèmes des organisations les plus complexes. Aujourd'hui, il dirige Promaintech Novaxa, une division de DMR, qui a célébré son dixième anniversaire en 2006. Cette organisation ayant des places d?affaires dans la région de Montréal, Québec et Vancouver compte plus de 50 professionnels et a une carte de visite dans 16 pays. Les axes d?intervention sont le Management, la Productivité, la Fiabilité, l?Innovation et l?Amélioration Continue.

Feb 1, 2008

The construction, commissioning and operation of a continuous pilot-scale selective agglomeration (oil agglomeration) unit are described. This plant has been used to recover fine coal from an effluent stream that is produced by a jig plant which scalps the coal from a New South Wales (Australia) colliery. The reject slurry is extremely fine (66.5% minus 53 um) and has an ash content of approximately 40%. The pilot plant can treat up to 0.3 tonne/h of this slurry, producing average product ashes as low as 9.1% and tailings ashes greater than 80%. The average yield observed was62%, which corresponded to a recovery of coal material of 88%. The operation of this pilot plant yielded data that enabled an economic assessment of this type of processing to be carried out. This study showed that selective agglomeration is an economically viable process for eliminating the substantial losses of coal to tailings ponds and minimizing consequent disposal and environmental problems. A commercial plant has been constructed and commissioned on the basis of the data obtained from this pilot-scale operation.

Jan 1, 1979

By Andrew G. Patton

The Alimak mining method was introduced at the David Bell mine in 1998 to extract an otherwise uneconomic and inaccessible reserve. The success associated with the extraction of these reserves led to major expansion of the method. To date, results have been extremely positive in terms of grade control, reduced dilution, stope recoveries, productivities and mining costs. It is hoped that through further development of the method at David Bell mine we will continue to see positive results in these areas and be able to increase mineable reserves to include proven resources that do not currently meet the criteria of our current stoping plan.

Oct 1, 2001

By M. N. Bassim, W. W. Tennese

"Concepts of linear elastic fracture mechanics were applied to study two cases of failure, namely an equalizing beam from a tandem rear axle of a highway tractor unit and the main beam of a flat bed trailer. In the first case, the presence of a crack, initiated by stress corrosion cracking followed by significant fatigue crack propagation produced a state of stress which was sufficient to produce brittle failure. Calculation of the fatigue life agreed with the number of cycles that the structure experienced before final fracture. In the second case, it was demonstrated that a combination of low temperature and high loading rate was sufficient to cause brittle fracture.IntroductionFailure analysis of engineering components has generally been carried out without the benefit of fracture mechanics. It is generally understood that real structures inevitably contain internal or surface flaws which under an applied load, may propagate to fracture. Fracture mechanics relates the applied stresses to the crack size in specified geometries. Since its introduction approximately thirty years ago, fracture mechanics has evolved as a design philosophy which can be used to develop fracture-safe designs. It can be also applied in analysis of failures, to ascertain if sound design and fabrication methods have been used."

Jan 1, 1985

By M. F. Almeida

"Due to the randomness of all factors influencing the sintering operation, a two-stepped procedure is proposed to find the best way of using stocked fines at a plant producing Fe-Mn alloys for the steelmaking industry. The approach is based on the following: (1) predict the result of the sintering operation by establishing the best multilinear regression model using some of the characteristics of the loaded material; and (2) find the best solution under a minimizing cost criteria by solving a linear optimization problem using the regression equation plus constraints on the feed and product. IntroductionSome tests carried out using 20% to 30% of sinter in the material loaded into a 60 000 KVA Tysland-Hole electric furnace producing Fe-Mn alloys for the steelmaking industry indicated: (1) a meaningful decrease in the off-gas temperature and, consequently, a better energy use due to the material higher porosity; (2) more uniform heat distribution as a consequence of the electrode reaction zone enlargement; (3) an improved furnace control resulting in a smoother operation; and (4) a decrease in gas production facilitating the off-gas treatment. For some years, this plant used only the coarse fraction of the Mn ores, thus stocking all the received and own-generated fine particles. In view of the mentioned results, a strategy for increasing the use of stocked fines was formulated, including a step where it was necessary to decide which components and respective quantities should be used in the blends to be sintered (Almeida, 1984).It is well known that several factors affect productivity and energy consumption in sinter production, also influencing the product quality. Raw materials characteristics, namely chemical and size composition, affect permeability to gases and may or may not allow a deeper bed. In general, finer ores decrease permeability, increasing electricity consumption by fans, but due to their higher specific surface area, react more completely with other components in the mixture. For coke granulometry, a compromise must be obtained in order not to burn it too quickly nor to widen the flame front, therefore decreasing productivity. Decreasing coke in the blend also acts in sinter bonding, decreasing its consistency and increasing crushing and return fines. Moisture addition also influences components mixing, pre-agglomeration, bed permeability and, consequently, sintering productivity. Therefore, optimum conditions must be established for both a given sintering plant and the components in the feed (Dawson, 1993)."

Jan 1, 1998

By Fred W. Brackebusch

"The use of paste technology for surface tailings disposal may offer economic, technical, safety, and environmental advantages for both underground and surface mining operations. Paste is a non-Newtonian fluid having some shear strength during transport and placement. The strength is augmented after placement by consolidation, water loss, and by addition of chemicals. Economical dewatering systems are necessary to the success of surface paste tailings disposal. In an underground mine, the paste backfill plant can produce paste for surface disposal about 50% of the time because only about 1/2 of the tailings are needed for backfill. The distribution of paste to the surface stack involves expensive pumping equipment and thus a nearby location of the stack is desirable. The viscosity of paste can often be manipulated by modifying pH. Placement of the paste in a designed stack configuration depends on topographic and climatic factors as well as paste properties.INTRODUCTIONThe use of paste technology for mine backfill is becoming accepted for underground mines mainly because of ground control and economic reasons. Paste technology is just starting to be applied for tailings disposal with only a few hardrock metal mines worldwide using the technology; however, it has been used extensively for red mud disposal in the aluminum industry. The disposal of dry tailings is very attractive from environmental and safety points of view because of limited water which might potentially escape and pollute river systems. Dry stacking systems, however, are expensive because of filtration and truck transportation costs. Paste can be produced from most tailings streams by special types of thickeners. Furthermore, it can be transported in pipelines, therefore paste technology may be more economic than dry stacking but still have some of the major advantages. Understanding of paste technology has improved through experience with operating underground backfill systems in recent years and because of testwork and research programs by mining companies, consultants and equipment vendors. Thus we may be on the verge of taking a new look at tailings management using paste technology."

Jan 1, 2002

By M. G. Dellow, D. W. Hendriks

"In early 1988 the decision was made by Noranda Minerals Inc. to install pressure filters at the GECO mill to replace the vacuum disk filters and bunker-C fired rotary kiln dryers. Two filters were installed, one for copper concentrate and one for zinc concentrate. Completion of the project was scheduled for the third quarter of 1989. The filters were made operational in October 1989. Some work remains to be done yet on loadout conveyors for total project completion. Capital expenditures for the project have been approximately $2,700,000 to date. The paper describes the selection criteria for the filter, the installation, startup and associated problems.1. INTRODUCTIONThe filter model chosen for the project was the Verti-Press model VP-50-6(7) (Figures1 ,2 and 3) by Filtra Systems of Farmington Michigan, USA. The decision to purchase the Filtra Systems filter was based on gaining familiarity with a different, mechanically simpler pressure filter than the Larox design and to a lesser extent a necessity for repairs to the the disc filter, rotary kiln dryer system used at the GECO operation for its' copper and zinc concentrates. The fact that it is manufactured in North America was an important consideration. The same basic principles of pressure filtration as the Larox machine are used but the implementation is quite different and somewhat simplified with some interesting innovations. 2 filters were installed to allow direct and immediate comparison of the same machine on both fine and coarse materials. Copper concentrate is reground for cleaning on the flotation circuit while zinc is not. Screen analysis of composited samples of concentrates for 1988 indicated that expected copper concentrate would be approximately 66% minus 400 mesh and zinc 50% minus 400 mesh (Tables 1 & 2).The filter is composed of six horizontal chamber module layered vertically and held together in a frame by 1 O large bolts; torqued down to 4000 ft-lbs. 6 individual belts are used; each with its own drive system. The belts themselves are of a 3 piece design with a filter cloth portion, a support belt and a rubberized intermediate portion. They are held together with steel clipper lacings. A basic description (1) of the filter operation is as follows. ""The horizontal filler chamber is closed by hydraulically pressurizing seals across each end of the filter chamber to 200 psi. The feed pump then pumps slurry into the filter chamber. Suspended solids are retained on the filter belt and the filtered solution drains into a lower drip tray and flows from the filter by gravity. At the end of the filtration or feed cycle, a rubber diaphragm in the filter chamber is pressurized hydraulically to 100 psi pressure. This removes all free liquid from the filter chamber and the filter cake. After the diaphragms are retracted the filter cake is blown with compressed air to remove any residual cake moisture. The seals are then retracted and the belt is advanced to discharge the filter cakes. During the cake discharge the belts are washed with a spray header located at the drive end of the filter. The belt assemblies move one (1) complete revolution during the cake discharge cycle. The cycle automatically repeats until the filter is shut down by the operator."

Jan 1, 1990

By K. C. Armstrong, A. B. Cron, L. A. Melis

"The Johnny Mountain gold mine in northern British Columbia, owned by skyline Gold Corporation, has been in operation since August 1988. Mill operating difficulties at the start of the operation resulted in a depleted cash flow situation and quick improvements were required to generate a self-sustaining financial basis of operation. In October 1988, Melis Engineering Ltd. was contracted to identify metallurgical difficulties, implement necessary modifications, and undertake mill operations and personnel training in order to achieve production targets.This paper presents the operating difficulties which existed at start-up and describes the methodology used to overcome production problems. Metallurgical testing results are described along with the modifications made to the milling circuit to suit ore metallurgy. Plant changes led to the operation of a simple gravity/flotation circuit to recover gold as bullion and in a copper concentrate. Operation of the originally installed cyanidation circuit was discontinued. This contributed to substantial savings in operating costs and virtually eliminated environmental problems. PROPERRTY LOCATION ANR DESCRIPTIONThe Johnny Mountain mine site is located in northern British Columbia along the Iskuit River about 60 km east of Wrangell, Alaska and 275 km north of Terrace, B. C.. A location map is presented in Figure 1 and a site plan is shown in Figure 2. The complex is located on a mountain plateau about 3, 600 feet above sea level. The plateau is flanked to the north by Johnny Mountain and surrounded on the other sides by steep cliffs."

Jan 1, 1990

By Don Peel

Mineral aggregate, the most widely used mineral resource globally, has attracted little academic or public interest. Of all mining developments, the aggregate resource is seldom factored into resource management decisions. This lack of recognition to the strategic aggregate resource has significant economic, environmental, and social implications. Impact to the economics, the environment, and the community coincide with the key elements of the United Nations concept of ?sustainable development?. To build on the concept, the mining sector contracted the International Institute for Environment and Development to assess how mining activity contributes to sustainability. As a result, a useful assessment process was developed by the Mining, Minerals and Sustainable Development (MMSD) Task Group entitled ?Seven Questions to Sustainability: How to Assess the Contribution of Mining and Mineral Activities.? In Alberta, recent opposition to mineral resource development by community residents offers an opportunity to employ the MMSD?s assessment procedure. The public opposition to the mining activity was addressed through a stakeholder planning process initiated by the municipal council of Sturgeon County, Alberta. As a result the Calahoo/Villeneuve Area Structure Plan (CASP) was formulated and adopted. To measure the effectiveness of the CASP planning process in terms of sustainability, this paper applies the MMSD?s Seven Question assessment to the Calahoo/Villeneuve Area Structure Plan.

May 1, 2004

By J. Yue

"Mount Milligan mine is located in northern British Columbia. The concentrator employs conventional SABC grinding circuit with bulk sulphide flotation processes. Commissioning started in September of 2013 and nameplate production of 60,000 TPD was reached in the fourth quarter of 2015. As is typical with many mill ramp ups, bottlenecks developed as throughput increased. Since then, plant modifications and new control philosophies in the comminution and flotation circuits have been implemented and evaluated. The mine-to-mill collaboration played a key role in reaching target throughput rates by improving blast fragmentation and implementing temporary pre-crushing. Results of the projects as well as ongoing challenges/bottlenecks and solutions will be discussed.IntroductionMount Milligan Mine is a copper-gold producer located approximately 253 km northwest of Prince George in British Columbia. Production commenced in September 2013 and reached nameplate capacity of 60,000 TPD in December 2015. Average head grades for the mine life average 0.18% Cu and 0.39 g/t Au.Circuit DesignOriginal process design began with run-of-mine ore from the pit crushed through a 60-110 gyratory crusher, conveyed to a live ore stockpile and fed to the mill via four variable apron feeders. The concentrator circuit employs SAG mill - Ball mill - Pebble Crusher (SABC) grinding configuration to grind the ore down to a target P80 of ~200µm. It is then fed to two parallel trains of flotation tanks, each consisting of five 200 m3 tank cells with the first two being roughers and last three being scavengers. Concentrate streams from roughers and scavengers are separately reground with a closed circuit Tower mill and two open circuit ISA mills operating in parallel, respectively.Three stages of cleaners upgrade roughers/scavenger products to 25% Cu followed by a cleaner scavenger stage. The final concentrate is then thickened and filtered to approximately 9% moisture and shipped off site (Figure 1)."

Jan 1, 2017

By Jean Rosaire

?GWIOP Location ?Climate and Physiography ?History ?Geology ?Deposit Types ?Mineralization ?Historical Mineral Resources ?Conclusions and Work Program

May 1, 2007

By Ralph Harris, Qi Xia

"Gold cyanidation rates are known to be limited by the rate of oxygen dissolution in the cyanidation tanks. The effect of solids loading on the oxygen dissolution rates is not obvious and the results of previous studies are ambiguous. The present paper presents an analysis of oxygen dissolution rates using a ""spent"" commercial gold slurry diluted to various levels of solids loading. It was found, in an 170 litre stirred tank reactor using an inverted cone sparger, that under certain high loading conditions, there was a decrease of as much as 50 %, i.e., there was a halving of the rate of oxygen dissolution as compared to that in clean water. In particular, it was observed that, due to gas flooding of the impeller, there was a drop in kLa even with an increase of gas velocity. Gas holdup reached a maximum value of only 1.7 %. The total power consumption was between 0.06- 0 .10 kw/m3 within the range of gas velocity before flooding occurred. The decrease in oxygen dissolution rate with increased solids loadings is believed to be mainly due to the increase in the apparent viscosity which reduced the turbulence and shearing effect and resulted in larger bubbles and hence reduced the reaction area a. The interrelations between solids loading, gas holdup, flow pattern, power consumption, and liquid mass transfer coefficient are discussed.IntroductionThe reactors generally used in gold cyanidation are Stirred Tank Reactors (STR). Compared with other types of reactors like fluidized bed reactor and column reactor, STR is of particular importance due to its wide application in industry. It is reported that the stirred tank reactor is used for more than 50 % of the world's chemical production (Xu and McGrath, 1996)."

Jan 1, 1997

By P. Lefebvre, J. Sargent, K. Major, B. Darnton

"Stillwater Mining commenced mining operations at Nye, Montana in 1987 for the recovery of platinum and palladium. Exploration identified that the J-M Reef extended for 43 km. In 1998 Stillwater Mining made the decision to develop the East Boulder mine complex, located approximately 13 miles from the Stillwater mine complex at Nye, to increase production. This paper describes the design, construction, and start-up of the East Boulder process plant.INTRODUCTIONStillwater Mining Company have been producing Platinum and Palladium from their Nye, Montana operations since 1987. The mine production from Nye had been increased in stages with mine production being delivered through several portals and one vertical shaft.The deposit is narrow veined which prevents the application of bulk mining methods and thus limits the production capabilities.In order to increase the production of platinum (Pt) and palladium (Pd) Stillwater made the decision to access the ore from the J-M Reef from the East Boulder site. The East Boulder mine site is located about 21 km (13 miles) from Nye, however as shown in Figure 1, the travel distance is closer to 188 km (114 miles). Distance, on strike, and space availability at Nye precluded the expansion of the Nye concentrator to handle the increased mining rate and supported the decision to install a new concentrator at East Boulder."

Jan 1, 2003

By Paul Laframboise, Nigel MacDonald, Mike Hallewell, Robert Thorpe

"Galmoy Mine, owned by Arcon Mines Limited, has operated since 1997, and is one of three operating lead/zinc mines in Ireland. Changes in ore type starting in mid 2002 created challenges to metallurgical operations at Galmoy. This resulted in unacceptable lead and zinc flotation performance, which prompted initiation of a formal plant review and a subsequent stabilization programme. This involved a structured review of operational practices, development of a new Standard Operating Procedure (SOP) which in turn led to significant changes in process control equipment needs in conjunction with extension of zinc flotation cell residence time.This paper will describe the metallurgy of the Galmoy operation. Reasons for the poor metallurgy experienced in 2002 will be described, as well as the solutions identified to overcome these problems. By the second quarter of 2003, these solutions had been implemented and recoveries had risen to record levels.INTRODUCTIONGeneralGalmoy Concentrator is located in County Kilkenny (Refer to Figure 1 below) and was constructed during 1996 and commissioned in 1997.The concentrator at Galmoy had been designed initially to treat 650,000 tonnes per annum of primarily CW ore type. The main mineral of economic significance was sphalerite (Typically 10 to 11% Zn) with some minor galena (0.7%Pb) in a dolomite (Magnesium Carbonate) host rock.The only other competing sulphide mineral was pyrite (Circa 3 to 5% Fe in Mill Feed. This CW ore type predominated the first 6 years of production, but in August 2002, the ore mineralogy changed due to significantly higher levels of pyrite rich G ore body being blended in with CW ore."

Jan 1, 2005

By G. Papachristodoulou, O. Trass

"Performance data for both wet and dry grinding of materials have been obtained in this novel, planetary ring-roller mill over wide ranges or operating conditions. The Szego mill consists of a stationary cylindrical shell inside which grooved, radially mobile rollers rotate, driven by a central shaft. The mill has several internal design variables such as number, diameter and length of rollers as well as size, shape and pitch oi helical grooves. It is compact, has a high specific capacity, modest power consumption, and can grind many materials down to ultrafine powders. Illustrative data on product fineness and mill power consumption are given along with results from studies on grinding kinetics.INTRODUCTIONThe extensive use and importance of grinding as a unit operation, coupled with the rather low energy efficiency of conventional equipment provide a challenging incentive to investigate novel methods of comminution. Yet, most current development work is taking place in mill performance modelling and optimization, or with various detailed improvements rather than in the field of new equipment development.This paper is a report on the development of a ring-roller mill invented by the late L.L. Szego which embodies several novel concepts. The mill has a broad range of applications, may be operated wet or dry, and can grind materials down to ultrafine sizes with modest energy consumption. It has been developed by General Comminution Inc. in close cooperation with the University of Toronto."

Jan 1, 1986

By Dean Lindsay, Inna Dymov, Mariam Melashvili, Jim Smolik, Chris Fleming

"This paper presents results from the proposed cyanide detoxification circuit for Seabridge Gold’s KSM project. The proposed flowsheet uses a two-stage CCD (counter-current washing/decantation) circuit to transfer the flotation products CIL(carbon-in-leach) barren solution to a SART process (Sulphidization, Acidification, Recycling of precipitate, and Thickening of precipitate) to recover copper and free cyanide from the copper cyanide complex. The SART barren solution is then treated by an AVR process (Acidification, Volatilization of HCN gas, and Reneutralization) to concentrate the cyanide that was liberated in SART for recycling to leach. The washed CIL barren pulp (the CCD underflow) is treated with the SO2 /Air process to oxidize residual cyanide.The goal of this complex flowsheet is to recover and recycle as much cyanide as possible and produce a final effluent containing less than 0.5 mg/L copper and CNWAD at the lowest cost. The SO2/air process is capable of achieving low residual CNWAD (<1 mg/L) in the final product, but the persistence of aqueous copper above the target level has been detected in the treated solution in some instances. The expected high concentrations of copper and cyanide in the KSM CIL barren ( >200 mg/L each species) and the difficulties this might create for the SO2 /air process was a key factor in the decision to consider including the washing, SART and AVR steps in the project flowsheet.The target concentrations of CNWAD and Cu were readily achieved after aging the SO2 /air treated pulp, as would occur in a tailing or polishing pond. This suggests no additional treatment will be required to sustain low levels of these species in the final plant effluent."

Jan 1, 2015