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By R. Amelunxen

"The molybdenite associated with copper porphyry ores is generally treated in a flotation circuit consisting of a separation and a cleaning section. In the last few years column flotation cells have replaced the numerous counter-current stages that were previously used in the cleaning section. Despite this, little has been published regarding the number of column cleaning stages required. Therefore a Moly flotation database is assembled with the objective of verifying the existence of a common upgrading profile. This can then be used as a guide in the prediction of the number of stages required to achieve a given concentrate grade.INTRODUCTIONAlthough the first use of Column Flotation was on copper flotation products, the popularity and understanding of their operation was not really achieved until they were widely implemented in Moly circuits in early 1983. The use of columns in Moly plants has since left a legacy of information, both published and unpublished. The data collected since that time has been used primarily used to establish column operating parameters and to determine upgrading potential within the circuit.Byproduct Moly generally occurs with porphyritic copper ores in concentrations less than 0.1% MoS2. As such, piloting a new moly ore to produce, say 5 Kg/Hr MoS2, in a slurry approximately 1 Lt/min (at 10% solids), requires a moly plant feed of about 570 Kg/Hr, which translates to a mill feed rate of 200 to 800 Tonnes per day (depending whether you deal with Chilean or Canadian ores). Piloting Moly for any lower tonnage would reduce the accuracy of the testwork.Thus a need exists to collect operating data in order to establish the existence of process relationships that could be used for feasibility and design purposes.The object of this paper is to expose a preliminary relationship that could be used for the purposes described above and to apply the findings to a preliminary design of a moly cleaning circuit flowsheet."

Jan 1, 1990

By J. A. Finch, J. Vergouw

"Flotation selectivity becomes progressively more difficult to achieve with decreasing particle size. One of the possible problems is agglomeration, i.e. particles clustering together. One mechanism derives from the surface charge all particles develop in water. Agglomeration driven by surface charge is known as coagulation. One form is coagulation between particles of two different minerals of opposite charge. Another is coagulation between particles of the same mineral when conditions are such that the charge approaches zero.The paper describes the method of measuring charge and a new automated settling technique for measuring agglomeration. The phenomena are illustrated with galena and pyrite. Methods of assessing the importance of agglomeration in the plant environment are still to be developed.IntroductionFine particles can agglomerate (i.e. adhere to each other) under certain conditions. This is counter-productive for any physical separation process, including flotation. One cause of agglomeration derives from the surface charge that all particles develop in water. Agglomeration due to surface charge effects is referred to as coagulation. There are two forms of coagulation. If the surface charge is brought to near zero then the force of repulsion is lost and particles coagulate. This usually applies to particles of one mineral type and the technical tenn is homocoagulation. On the other hand, if two particles have opposite charge there is an attractive force. Since the particles in this case are usually different minerals this is known as heterocoagulation.Both types of coagulation can cause minerals to be misplaced. In the case of heterocoagulation this is fairly obvious. One example that may be familiar is slime coating (fines of one mineral coating larger particles of another) But even homocoagulation can be a problem an agglomerate of the same mineral particles can entrap particles of another mineral."

Jan 1, 1997

By R. Simkus, J. K. Gowans

Polaris Operations, a Cominco Ltd. lead-zinc mine, is located on Little Cornwallis Island, Northwest Territories, at Latitude 75°23' N and Longitude 96°75' w. When the Polaris Mine went into operation in late 1981, it became the world's most northerly base metal mine.The Polaris Concentrator, housed in a unique process barge, is a basic lead-zinc concentrator where flotation is used to produce the lead and zinc concentrates.The paper describes the ""coarse lead"" flotation feature incorporated in the concentrator flowsheet and describes the start-up problems and continuing operation of this circuit.

Jan 1, 1984

By Francis Chachula, Niel Erasmus

"In 2002, Titanium Corporation Inc. focused its attention on the development of a process for the recovery of valuable heavy minerals, including zircon and titanium bearing minerals from oil sands froth treatment tailings. The development of a process flowsheet to recover the titanium and zircon minerals from the Athabasca oil sand tailings has evolved through many laboratory, bench and pilot scale iterations.Initial development work focused on so-called ‘beach sand’ material, which is tailings material deposited over a period of time close to the tailings pipe discharge. The derived process flowsheet primarily consisted of gravity concentration stages, which is typical of all mineral sands operations. Titanium Corporation’s business plan calls for the utilization of tailings directly from the tailings pipeline – the process developed around the beach material is therefore risky as it differs from the live tailings in a number of aspects, with the effects on the process unknown. Titanium Corporation successfully obtained an intermediate product through a ‘bulk sampling plant’ coupled directly to an Oil Sands operator’s froth treatment tailings pipeline, during the summer of 2005. Subsequent further processing of this material resulted in poor separation or concentration of the heavy minerals due to a number of reasons. A more effective process flowsheet had to be developed.The newly derived process is basically a marriage between mineral sands technology and oil sands technology that can economically concentrate and recover the heavy minerals and entails the incorporation of a flotation process that takes advantage of the entrained bitumen in the feed to separate the heavy mineral fraction from the quartz gangue This paper will describe the newly derived process as well as the successful results of a subsequent on-site concentrator pilot campaign completed during the 3rd Quarter of 2006."

Jan 1, 2008

"The measurement of error in assays collected as part of a mineral exploration program or mining operation historically has been undertaken in a variety of ways. Different parameters have been used to describe the magnitude of relative error, and each of these parameters is related to the standard measure of relative error, the coefficient of variation. Calculation of the coefficient of variation can be undertaken in a variety of ways; however, only one produces unbiased estimates of measurement error: the root mean square coefficient of variation calculated from the individual coefficients of variation.Thompson and Howarth’s error analysis approach has also been used to describe measurement error. However, because this approach utilizes a regression line to describe error, it provides a substantially different measure of error than the root mean square coefficient of variation. Furthermore, because regression is used, Thompson and Howarth’s results should only be used for estimating error in individual samples, and not for describing the average error in a data set. As a result, Thompson and Howarth’s results should not be used to determine the magnitudes of component errors introduced during geochemical sampling, preparation, and analysis.Finally, the standard error on the coefficient of variation is derived, and it is shown that very poor estimates of relative error are obtained from duplicate data. As a result, geoscientists seeking to determine the average relative error in a data set should use a very large number of duplicate samples to make this estimate, particularly if the average relative error is large."

Jan 1, 2007

By S. R. Gilbert, C. O. Bounds, R. R. Ice

"Many ores are upgraded to produce auriferous sulphide concentrates (principally pyrite) which are not amenable to direct cyanidation. Often these refractory gold-bearing concentrates contain undesirable elements such as arsenic, antimony, etc., which present environmental and/or processing problems. The pretreatment techniques most often considered for processing these concentrates prior to cyanidation include roasting, pressure oxidation, chemical oxidation, pre-aeration, and recently, bacterial oxidation. In this paper, the technical performance and economics of bacterial pre-oxidation vs the classical roasting approach for treating a specific auriferous pyrite concentrate are compared. The results provide flow sheets, preliminary operating and capital cost estimates, and rates of return for the two processes and demonstrate the potential advantages of bacterial pre-oxidation. IntroductionThe increased attention that gold has received in the last ten years has prompted mining companies to look for lower grade orebodies. Many of these low-grade ores do not respond well to conventional direct cyanidation recovery techniques. As free-milling ores decline, gold is sought in refractory environments which, ultimately, drive up the treatment costs.There are a number of reasons for a gold ore to be termed refractory, most of which can be identified by thorough mineralogical examination. The mineralogical occurrence of gold can, to a great extent, impact the method by which the ore is treated and, consequently, the capital and operating costs. Gold in a refractory ore can often occur as discrete particles locked in a host of silicate or carbonate gangue, or associated with sulphides or sulphosalts. Depending on its particle size, gold can be liberated and recovered by mechanical size reduction such as regrinding. This approach is most effective for gold in the 10 - 50 micron range. Gold can also be associated with clay, iron hydroxides, or other minerals which tend to coat the gold surface and inhibit the penetration of the cyanide solution, resulting in poor plant recovery. Because of their resistance to cyanide dissolution, gold tellurides and other refractory gold minerals are also responsible for losses to the tailings. For these reasons, it is paramount that the mineralogical occurrence of the gold be thoroughly and accurately determined before selecting a final process flowsheet."

Jan 1, 1988

By Claude Bazin, François Robichaud

"The control of the operation of semi autogeneous grinding (SAG) mills is mainly achieved through adjustments of feed rate, water addition and speed of rotation. With the introduction of image based sensors it is possible to have on-line measurements of the size distribution of the ore feeding a SAG mill. This information could be used for the development of a control strategy to regulate a SAG mill feed size distribution to a set-point. The control of the size distribution is carried out by adjustment of the speed of the ore extractors underneath the ore stockpile. The potential and difficulty associated with the implementation of such control are assessed using a multivariable statistical analysis of data collected from the operation of the grinding circuit of the Agnico-Eagle Goldex mine in Val d’Or, QC. A control strategy for the ore feed size distribution is assessed using a simulator of the ore stockpile operation.INTRODUCTIONThe popularity of autogeneous (FAG) and semi autogeneous (SAG) grinding mills has been increasing over the last 20 years (Napier Munn et al., 1996). These mills use the ore or a combination of ore and steel balls as the grinding media. Since the ore is used as a media for the size reduction the mill ore feed size distribution impacts the performance of the mill. Several papers were recently published on this subject (Dance et al.,2011; Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011) showing the advantages of preparing the feed size distribution to maximize the circuit throughput or mill efficiency (Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011). With the availability of particle size sensor based on image analysis of the ore on a conveyor (Partanen et al., 2001; Bobo et al. 2004; Edwards et al., 2002; Wei & Craig, 2009) it is possible to control the size distribution of the ore feeding autogeneous or semi autogeneous mills (Partanen et al., 2001).This paper describes a method to control a size distribution index to a set point via the manipulation of the speed of the extractors that withdraw the ore from the stockpile feeding the grinding circuit. The paper consists of 4 sections. The first section describes the considered grinding circuit. A statistical analysis of the collected data on the grinding circuit is presented in the second section. This analysis demonstrates the necessity of using real-time adaptation mechanisms to develop a control strategy for the size distribution of a grinding circuit. The third section describes the method used to simulate the ore stockpile operation and the last section describes the preliminary development and assessment of a control strategy of size distribution of the ore feeding a SAG mill."

Jan 1, 2013

By C. Bazin, D. Hodouin

"A flexible model involving perfect mixers, stagnant zones, plug flow, material bypassing and recycling zones has been developed for interpreting mixing phenomena in mineral processes. The mixing properties are described by the residence time distributions which can be subsequently interpreted by the flexible model. User-friendly computer programs (for 1MB-PC or compatible) have been developed for simulation and modelling of RTDs using the proposed model. Seven mineral processing case studies are discussed in terms of experimental procedure, model calibration results and application of the estimated mixing model to process modelling, diagnosis, and performance evaluation. IntroductionIn any mineral processing plant there are continuous-flow machines that transform or separate the ore components. Grinding mills, flotation cells, conditioning and leaching tanks are examples of such machines. The performances of these vessels or reactors are controlled by the rate at which the reaction proceeds and the time spent in the environment where the reaction occurs. This paper deals with the latter aspect of mineral processes operation.Since most of the reactions occurring in mineral processing units are time dependent, knowledge of the residence time of the material in process vessels is important for modelling and diagnosis purposes. Due to the random mixing phenomena involved in the process equipment, there is a distribution of the residence time of the ore particles and the fluid in the vessel. The probability density function E(t) that a micro-element of the material flowing through a processing unit stays a time t in the vessel, is the residence time distribution (RTD) of the corresponding component."

Jan 1, 1988

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 R. D. Lindsay

"Pamour Porcupine Mines, Limited is involved in the mining, milling and primary refining of various gold, gold - copper and refractory gold ores. Pamour operates five mines in the Porcupine Mining District of Ontario, and two 2,550 t/d milling facilities.Both metallurgical facilities use conventional cyanidation technology including zinc cementation for precipitation of gold and silver from solution. The precipitation process, as applied to clarified leach solutions (less than 1 ppm suspended solids) at Pamour is called the Merrill - Crowe process. Precipitates from both milling operations are treated at Pamour's Area Refinery. Schumacher precipitates, unlike Pamour No. 1, require an ''Acid Treatment Stage"" prior to refining to remove the high contained copper. This is accomplished by leaching the precipitates initiallv with sulphuric acid to remove zinc and then with sulphuric acid and manganese dioxide to dissolve copper. Refining procedures after ''Acid Treatment"" are essentially the same for both precipitates.The following considerations highlight the basic principals applied to refining cyanidation precipitates at Pamour, including:"

Jan 1, 1986

By Pathies Nawej Muteb, Tom Strombotne, Karl Leetmaa, Irma Gabric, Keith Darby

"Continuous control of the final grinding product size is important to optimize gold liberation for Gold Milling operations. At Meadowbank Mine an on-line Particle Size Monitor has been employed successfully to achieve the targeted final grind which leads to optimizing the gold recovery at maximum grinding circuit throughput. The live monitoring of the final product particle size leads to a maximum grinding circuit throughput every minute during the process. This paper discusses the optimal utilization of the Particle Size Monitor to assist the Mill operator to maximize the grinding circuit throughput as the milling conditions change due to variable ore processed at Meadowbank Mine. Also the paper includes the usage of the Particle Size Monitor to guarantee the final grind target at each instant of the process.IntroductionAgnico-Eagle is a Canadian-based company that Trades on the TSX & NYSE (AEM). It owns eight mines (8) operating in Nunavut, Quebec, Finland and Mexico and has more than 7,500 employees (Figure 1). 1.6 M ounces of gold were produced in 2015 and the annual average annual production from 2016 to 2018 is forecast to be approximately 1.53 million ounces of gold.The Meadowbank Mine is located in Nunavut, Canada (Figure 2). The mine opened in 2010 and from 2010 to 2014 a total of 1.8M oz of gold was produced. The milling rate is 4.2 Mt per year with an average mill throughput of 11,900 tpd or 496 tph. Average annual gold production over life of mine is expected to be 450,000 ounces. Meadowbank has 775 employees, of which 265 are Inuit or 36% of the total workforce. Estimated life of mine is 2010 to 2018. Gold production in 2015 was 381,804 oz. Meadowbank was Agnico-Eagle’s largest Gold producer in 2015."

Jan 1, 2017

By D. Bois, P. Keating, J. Lemieux, M. A. Vallée, M. Allard, R. S. Smith, L. Z. Cheng, M. Chouteau, K. Fountain

"A comparison of data from INPUT, GEOTEM, and MEGATEM airborne electromagnetic (AEM) systems over the Aldermac deposit, Québec, Canada, was undertaken to evaluate the relative ef¬ficiency of these systems for the detection of small, deep conductors. The Aldermac orebody was mined between 1933 and 1934 to a depth of 200 m. A deeper mineralized zone was discovered in 1987. The MEGATEMII system was used to acquire 90- and 30-Hz data over the deposit in 2003 as part of a research project. These MEGATEMII data were compared with data acquired by other AEM systems (GEOTEM and INPUT), and the results show that the MEGATEMII response is ten times larger than the other sys¬tems. For the dB/dt response, 30 Hz gives a better signal-to-noise ratio, whereas 90 Hz has a better signal-to-noise ratio on the B-field.The data were evaluated by numerical modeling. The deposit was best represented by a complex mod¬el, comprising two weakly conductive bodies to approximate the alteration halo, two deeper prisms repre¬senting known orebodies, and three bodies ranging in thickness from 3 to 6 m and in resistivity from 12 to 20 ohm·m, representing the chargeable overburden. This implies that the MEGATEMII system is able to see a small conductive body at more than 200 m depth in an area where the response is contaminated by conductive overburden."

Jan 1, 2009

By R. H. Colquhoun

"INTRODUCTION The Holt-McDermott Mine is wholly-owned gold mining operation of American Barrick Resources Corp. that begun operations in June 1988.The property is located 0.5 km. south of Highway 101 East in the Harker-Holloway Townships of Northeastern Ontario, 7 miles south of Lake Abitibi and 15 miles west of the Quebec-Ontario border.Employees of the operation travel daily from various communities in the area, predominantly from Kirkland Lake, Ont., which is located 65 km. to the southwest.The significance of the operation is that it is the first full scale operation associated to the Oestor-Porcupine Fault to go into full production in recent history as well as the first Carbon-in-Leach (CIL) operation in the province owing to the natural occurring carbonaceous ores that are associated to the ore body and the geological fault systems.GEOLOGYThe Holt-McDermott deposit is located in the Abitibi Greenstone Belt, a short distance west of the Ontario-Quebec border. Ore grade material follows the footwall contact of a fault zone known locally as the McKenna Fault, which is thought to be a splay off of the Oestor-Porcupine Fault, itself a major feature in the Abitibi belt.The McKenna Fault strikes approximately east-west (as does the Oestor-Porcupine) and dips to the south at sixty-five to seventy degrees. The fault lies within a package of mafic volcanic rocks (basalts). These rocks have been strongly altered and mineralized adjacent to the McKenna Fault, and the resulting lens-like ore bodies constitute the mining targets."

Jan 1, 1989

By Panagiota Tsolis-Katagas, Basilios Tsikouras, Tassos A. Grammatikopoulos, C. Katagas

"Mineralogical techniques including microscopic methods, reflected and transmitted light microscopy, Cathodoluminescence (CL) and Scanning Electron Microscopy equipped with an Energy Dispersive Spectrometer (SEM-EDS), are useful in determining the quality of industrial minerals. They are essential for studying ores and metallurgical products including the evaluation of physical characteristics of the mineral, paragenetic relationships, associations with possible deleterious contaminants (e.g., sulphides), spatial distribution within the ore zones and liberation characteristics. The methods are inexpensive, and provide quick identification and distinction among the industrial minerals and the gangue minerals in a section. Additionally, SEM is necessary to reveal complex textural sequences and 3-D measurements of minerals and quick semi-quantitative analysis.Common analytical methods include Electron Microprobe Analyses (EMPA) and Whole Rock Analysis (WRA) (commonly with XRF and ICP) for major and trace elements. The EMPA analyses of industrial minerals provide information on the purity of the analyzed grains as opposed to the bulk chemical analyses that may be misleading due to contamination by other associated common “gangue” minerals. These techniques have been implemented for the case studies of wollastonite, V-magnetite and muscovite. Therefore, these methods can provide extremely useful information and lead to the implementation of a successful pilot-testing programme.MINERALOGICAL EVALUATION OF DEPOSITSDespite their considerable and increasing economic importance, industrial mineral deposits have traditionally been neglected by economic geologists, and are only rarely the subject of the detailed analysis widely lavished on metallic mineralization. Moreover, in that they generally represent concentrations of “normal” rock-forming minerals, albeit commonly with particular textural and/or geochemical features, they have tended to fall between the mainstream of petrogenetic studies and mineral deposit research (Grammatikopoulos 1999). For many industrial minerals, little attention has been paid to the factors responsible for their local concentration as large and high-grade bodies. This comparative neglect has particularly been the case for minerals such as talc and wollastonite, widespread in metamorphic rocks, in which they are reasonably interpreted as components of assemblages derived, respectively, from pelitic sediments, ultramafic igneous rocks and siliceous limestones."

Jan 1, 2003

By Casey Erickson

Goldstrike Property ??Located in Elko, NV USA?Carlin Trend ??Mining Since 1985 ??27 Million Ounces of Gold Mined ??Production of 400,000 tons/day ??1,600 Employees

May 1, 2011

By Bryn Harris, Liming Huang, Andre Laplante

"This paper reviews flowsheets generally used in gold rooms to upgrade primary gravity concentrates particularly those of Knelsons. It is concluded that even for small operations, which treat between 250 and lo00 kg of concentrates per day, there is a strong economic incentive to include in the flowsheet magnetic removal of tramp iron, at least two shaking tables, screening of either table feed or table tails at 150-212 um, and scavenging of table tails with a small centrifuge. Headroom, rather than floor space, is often inadequate in most gold room designs, and limits gravity flow and flowsheet flexibility. Experimental data are presented to demonstrate the effectiveness of a small Knelson Concentrator to scavenge significant amounts of gold (up to 91 %) from fine table tails.RésuméCet article examine les schémas d’évolution généralement utilisés pour traiter les concentres gravimétriques d‘or, en particulier ceux produits par concentrateur Knelson. Nous en venons à la conclusion que même pour un petit circuit d’une capacité de 250 A 1000 kg/jour pourrait bénéficier des opérations unitaires suivantes : concentrateur magnétique pour extraire la feraille, au moins deux tables A secousses, criblage de l’alimentation ou du rejet d’une ou des tables i une maille de 150-212 um, et épuisages des rejets fins de table avec un centrifuge de petite capacité. C’est l’espace vertical du lieu où se fait cette concentration qui représente la contrainte la plus importante, plutôt que sa surface, puisqu’il faut maximiser l’écoulement par gravite des pulpes et s’offrir une certaine flexibilité quant à l’agencement final des opérations unitaires. Nous présentons également des données expérimentales pour démontrer l’efficacit6 d‘un Knelson de laboratoire pour récupérer jusqu’ à 91 % de l’or contenu dans les rejets fins de table."

Jan 1, 1999

By S. Banisi, M. Poorkani

The Sarcheshmeh copper deposit contains 1 billion tones of ore averaging 0.7% copper and 0.025% molybdenum. It is located southeast of Iran and currently processes 41,000t/d. Molybdenite as a by-product is recovered in two separate stages. First, a bulk coppermolybdenum concentrate is produced and in the second stage these two are separated. The recovery of molybdenite in the first and second stages is 65% and 85-90%, respectively. The yearly production of molybdenite concentrate varies between 3500 and 4000 t. Sodium sulphide is used as the depressant of the copper and iron bearing minerals which due to a rather high consumption (17.7 kg/t) constitutes one of the major cost items of the plant (i.e., 58% of total reagents cost). A detailed laboratory scale study showed that replacing air with nitrogen could reduce the consumption of sodium sulphide (i.e., H2S or mineral origin). Nitrogen for the plant was supplied from an oxygen plant used for the smelter through a 1400 m long pipeline. The existence of the oxygen plant at the Sarcheshmeh complex was a key factor in the implementation and success of this project. After using nitrogen in the plant for a period of twenty one months, the consumption of sodium sulphide was reduced from 17.7 to 14.2 kg/t at constant metallurgy. The reduction was lower than expected which was mainly due to the type of sodium sulphide used and the lack of any automatic instrument to control the depressant (sodium sulphide) dosage. The latter in turn caused the operators to overuse sodium sulphide in order to reach and maintain the target metallurgy (i.e., higher molybdenite recovery along with lower copper content of the concentrate).

Jan 1, 2005

By Ed Beswick

Resource Update Drilling Program ? Pre-feasibility ? Exploration Metallurgy Program ? Grindability ? Metallurgical Optimization ? Hydrometallurgical Options Preliminary Mine Development Design Environmental Permitting Infrastructure Development

May 1, 2008

By Jean Cayouette, Paul Blatter, Paul Cousin

"The Laronde mill was commissioned in 1988 and has been treating the Laronde Penna ore since 2000. The Laronde plant, situated in the Abitibi region of Quebec, has experienced mill expansions to accommodate 7500 mtpd and to recover gold and silver as well as important amounts of base metals (Cu, Zn, Pb). Ten sequential metallurgical and environmental processes are responsible for metals recovery and process viability. Over 200,000 ounces of gold, 4,5 million ounces of silver, 7000 tonnes of copper, and 78,000 tonnes of zinc are produced annually.In 2008, the Goldex plant began trucking high grade (35-45 g/t Au) pyrite sulphide concentrate (TCGO) for gold extraction at a newly constructed treatment plant, increasing gold output at Laronde by 40,000 ounces. In 2009, the high gold grade Lapa ore processing facilities were annexed to the Laronde plant increasing output by 120,000 ounces of gold. TCGO and Lapa productions are forecasted until 2018 and 2015 respectively, with known reserves of each mine.Today, up to 140 permanent employees run and optimise the Laronde plant operation. As a result, Laronde has intricacies intrinsic to a large and varied operation. The paper will outline the metallurgical highlights, challenges, and future outlook related to the treatment of each ore type (Laronde, TCGO, and Lapa).INTRODUCTIONFollowing the 2000 to 2003 expansions, the Laronde plant evolved into a mature milling complex with developed processing, man power, and water treatment capabilities able to accommodate the complex Laronde ore. The ore requires a series of grinding, copper/lead flotation and separation, zinc flotation, zinc tails precious metals leaching, followed by a counter-current decantation circuit and Merrill Crowe precipitation. A paste backfill and SO2-air cyanide destruction operate intermittently. The tailings area has a dedicated cyanide destruction and metals precipitation plant prior to water recycling to the mill. A biological water treatment plant was put into service in 2004 to address the build-up of thiocyanate in the tailings ponds at Laronde. This was the result of the high sulphide content of the Laronde ore and 90% recirculation of the process water. The plant uses bacteria to oxidize and destroy thiocyanate, and removes phosphate from the water before it is released to the environment."

Jan 1, 2011

By Q. Zhang, J. A. Finch

"Sphalerite flotation in the absence of copper activation at moderately alkaline pH has been observed in laboratory studies. An in-plant test of this observation was conducted at Kidd Creek Mine LTD. using a portable minicell flotation circuit. The results confirm that raising the pH causes sphalerite flotation.The testwork with the minicell circuit is described. The circuit, built around 2, 51 cells donated by the Noranda Technology Centre, was funded by Air Liquide and NSERC. It includes a data acquisition system to monitor continuously dissolved oxygen, pH and Pulp potential.INTRODUCTIONFlotation of sphalerite is normally practised by adding CuSO4 as an activator under alkaline pH. Recently there have been a number of laboratory and plant examples of sphalerite flotation without deliberate addition of cu++ ions (Girczys, 1972; Mukherjee, 1976; McTavish, 1980; Marouf et al., 1986; Mielczarski, 1986; Leroux et al., 1987; Palsson and Forssberg, 1989; Sandoval-Caballero et al., 1990; Zhang et al., 1991).In batch tests, with a sample of Pb/Zn ore from Polaris, Leroux et al. (1987) found that sphalerite became floatable after adjusting pH to 9. 5-11: they ruled out natural floatability or the formation of elemental sulphur on the surface but suspected iron hydroxy compounds were responsible. Sandoval-Caballero (1990) conducted a continuous minicell test at Brunswick Mining and obtained a 44 % Zn recovery and 40.5 % Zn grade concentrate without deliberately adding CuS04 at pH= 10.5. With single mineral studies, Zhang et al. (1991) revealed good sphalerite floatability (over 80% recovery) in pH range 8 to 11 and found that this floatability was related to pH, dissolved oxygen (DO) level and Fe++ species present in the system.To explore further the conditions under which sphalerite can float without addition of CuSO4 , a continuous pilot test using minicells was performed at Falconbridge Ltd., Kidd Creek division. The set-up is briefly described (a more detailed description is given elsewhere in these proceedings) with emphasis on the data acquisition system - dissolved oxygen, pulp potential and pH were monitored continuously. Results of testwork investigating pH, CuS04 addition and gas composition are presented."

Jan 1, 1992