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By K. Hoover, G. Labarre, S. Rantanen

"As part of the Namew Lake project a mill was constructed to treat 1900 tpd of ore grading approximately 0.90% CU and 2.45% Ni to produce copper and nickel concentrates.Nickel occurs mainly as pentlandite and violarite while copper occurs as chalcopyrite. Two main ore types are present1 lower grade disseminated sulphides in a talcose ultramafic gangue and a higher grade solid sulfide with a granitic breccia matrix.Primary crushed ore from underground is treated on surface in a conventional closed circuit crushing plant using 2-stage cone crushing. Crushed ore is conveyed to storage and then fed at a controlled rate to a grinding circuit employing a rod and ball mill with hydrocyclone classification. Rougher flotation separates and cleans nickel and copper sulfides from waste materials followed by a high lime separation circuit to separate the minerals into final concentrates. Dewatering is achieved through use of high capacity thickeners and pressure filters prior to delivery to a loadout area. The crushing and milling control is highly automated.Plant design was based on metallurgical samples obtained from drill cores. After the first year of operations, plant results compared favourably with design projections.INTRODUCTIONThe Namew Lake Project is a joint venture partnership between Outokumpu Mines Ltd. (Canada) and Hudson Bay Mining and Smelting Company Ltd. (60%). The project included construction of a mine and milling complex to treat approximately 500,000 tonnes per year of copper-nickel ore grading o.91% and 2.44% Ni. Known reserves are estimated at 2. 65 million tonnes. The deposit is located approximately 65 km south of Flin Flon (120 km by road) and about 5 km east of the Manitoba-Saskatchewan border. The decision to construct a concentrating facility was made June 15, 1987 and construction started a week later. The facility, designed for 1900 tonne per day throughput, was ready for commissioning in September 1988. Outokumpu Engineering of Finland was chosen to prepare metallurgical flowsheets, select and develop plant layouts, and provide construction management. Kilborn (Toronto) contributed detailed engineering for civil, structural, mechanical and electrical components and outokumpu Equipment of Canada, the design, the supply, and commissioning of the instrumentation package. Hudson Bay undertook the initial metallurgical test work and preliminary design work."

Jan 1, 1990

By F. Hassani

This Paper describes a sudden spill of rock waste used for backfilling from a waste pass at one of the underground mine in Canada thus causing a major failure/accident. The mine operator immediately launched a detailed investigation to understand the causes of this accident and develop procedures to prevent any recurrence of such events in future. The investigators from McGill University were invited to collaborate with the engineers at the mine to examine this accident. The events leading to the accident were studied and appropriate measurements were taken and samples of the waste material was collected for further test and analysis. The field visit and the interview of the witnesses allowed a detail documentation of the events. The sequences of events related to the incident together with the experimental results obtained in the course of the investigation is detailed. The waste material was characterised in the laboratory with regard to its behaviour in the presence of water. This information is then discussed and the failure mechanism involved in this accident is identified. The full recommendation as well as the immediate operational measures implemented by the mine to prevent future occurrence of such events are discussed and presented.

May 1, 2003

By M. J. Allan

"The paper describes the author's experiences in the startup of autogenous and semiautogenous grinding circuits gained over the last twenty years on both large and small grinding circuits. Typical phases in commissioning from construction to startup with ore are covered. The paper also describes experiences with material handling systems, slurry pumping, vibrating screens, trommel screens and pebbles crushing systems. INTRODUCTIONThe commissioning of any new mineral processing plant is an important and exciting milestone in any project. It is at this point in time that the efforts of developers, design engineers, equipment suppliers and operators culminate in a finished product. The successful completion of any project requires that the plant achieve design throughput as quickly as possible. The success of any startup can generally be anticipated by the amount of careful planning which takes place before equipment is operated. The first part of this paper describes the steps in going from construction through to introduction of ore. The organization and planning of startup activities is also covered.The second part of this paper describes some typical problem areas and past experiences in the startup of semi-autogenous grinding circuits. Semi-autogenous grinding circuits present some special challenges in the startup phase. The description of some of these problems and experiences will hopefully provide other operators with suggestions on how to avoid (or at least prepare for) these problems."

Jan 1, 1992

By Fabian Mohr

The methyl substituted binuclear cyclometallated gold(I) complexes [Au2(?-n-MeC6H3PPh2)2] [n = 4, 5 and 6] all undergo oxidative addition reactions with halogens and, in the case of the 4- and 5-methyl substituted analogues, with dibenzoyl peroxide to give metal-metal bonded digold(II) complexes [Au2X2(?-C6H3-n-MePPh2)2]. The 4- and 5-methyl substituted digold(II) compounds on warming or exposure to light undergo carbon-carbon coupling reactions to give gold(I) complexes containing the corresponding 2,2'-n,n'-biphenylyldiphenylphosphines. The 6- methyl substituted digold(II) complexes rapidly isomerise above ca. -20° C to binuclear heterovalent gold(I)/gold(III) compounds [XAu(?-2-Ph2PC6H3-6-Me)AuX{?2-(6-MeC6H3-2-PPh2)}], containing a fourmembered cyclometallated ring on a gold(III) centre. Surprisingly, no carbon-carbon coupling could be induced for any of the 6-methyl substituted compounds, even under forcing conditions. This difference in reactivity of the 6- methyl derivative, compared to that of the 5-methyl and unsubstituted analogues was studied by ab initio calculations using a series of model compounds. The results are in excellent agreement with the experimental findings and suggest that the key-step in the C-C coupling reaction is the partial opening of the Au(III)-P bond in one of the transition states. In the 6-methyl system this transition state is energetically inaccessible and hence no C-C coupling occurs.

Oct 1, 2003

By R. E. Jones

"Crusher sizing and application in mining operations is the selection of a crusher or a series of crushers to reduce run of mine ore to a size suitable for either autogenous or conventional grinding. It is important that the crusher or crushers be sized to operate within the range to which they have been designed. By the range to which they have been designed, we mean that the reduction ration must be held between specified limits. Reduction ration is the relationship between feed size to product size.In todays plants reduction is usually accomplished in three stages by primary, secondary and tertiary crushers.Primary crushers are either jaw or gyratory and are designed to handle run of mine feed at an open side setting of 5-1/2” to 7”. Reduction ration should not exceed 7 to 1.Jaw CrushersThe model size of a jaw crusher signifies the feed opening. A model 60 x48 jaw has a feed opening 60” wide by 48” long. The 48” is the dimension between the top of the swing jaw and the top of the stationary jaw. The 60” is the width between the two side frames. The crusher open side setting is the discharge opening at the bottom of the crushing chamber with the jaws in an open position. When corrugated jaw plates are used, the open side setting is measured from tip of one jaw plate to the valley of the opposite jaw plate. If a jaw crusher has a 6” open side setting with a 1” jaw motion, you could also say it has a 5” closed side setting."

Jan 1, 1974

By F. D. Ford, Z. Dai

"Vale’s Ontario Operations re-introduced Cu/Ni separation at Clarabelle Mill in 2006. The mill processes ores from a number of mines in the Sudbury basin. A saleable Cu concentrate is produced at 30-60% Cu recovery from mill feed, which has to meet the quality specification. In particular, Ni content in Cu concentrate must be less than 0.6%, otherwise, penalties will apply and diminish the value of Cu concentrate, in addition to the lost Ni value. Generally, a Ni grade of less than 0.4% is obtained in Cu concentrate. More recently, while processing Cu-rich ore, Cu concentrate contained more than 0.6% Ni which is abnormally high. Detailed mineralogical investigation shows that Cu-rich ore contains more millerite in comparison to typical Sudbury Ni sulphide ore. Millerite is found to be more floatable than pentlandite and is much more difficult to depress using high-lime aeration conditioning, which is the standard practice in Cu/Ni separation to float chalcopyrite while depressing pentlandite. The current operation uses sodium cyanide at 20 g/t to depress pentlandite and pyrrhotite. However, it is found that a high dosage is required to depress millerite. It is also shown that starch and dextrin are promising in millerite depression.INTRODUCTIONPrior to bulk smelting in 1991, Cu/Ni separation was carried out at Vale’s Copper Cliff Mill, Ontario Operation. The typical Cu concentrate grade was around 30% Cu and 1% Ni. This is very high in comparison to today’s value of 0.4% Ni and would not be acceptable for external processing. However, the Cu concentrate was processed internally at the Copper Cliff smelter/refinery, so the Ni contained in Cu concentrate was not lost and no actual penalties applied. During the SO2 abatement initiative, bulk smelting was chosen because of the perceived simplicity and consequently, Cu/Ni separation was eliminated (Agar, 1991) and the Copper Cliff mill was subsequently demolished. In 2005 and after more than one decade of bulk smelting, business changes required the removal of some Cu concentrate from Clarabelle Mill bulk concentrate to provide smelter capacity for Voisey’s Bay concentrate in the Copper Cliff smelter. In essence, this strategy calls for Cu removal at Clarabelle Mill (Lawson, 2006). In this case, Cu concentrate is sold externally and it must meet the quality requirement of less than 0.6% Ni."

Jan 1, 2011

By Jean-Pierre Côté, Alexandre Noiseux

"Hatch and ArcelorMittal Mines Canada have combined their expertise to develop a dynamic model of the latter’s Iron Ore Pellet Plant in Quebec, Canada. This model was used to predict the benefits of different expansion projects, as well as asses the best implementation sequence.Two modelling tools were used: a discrete-event simulation software used to develop the dynamic model of the Pellet Plant and an in-house process simulator. The latter was based on operational data and known behavior of the process, and was calibrated with historical data gathered under different operating conditions.The results of the process simulator were used as input values for the dynamic model, which included all of the operation steps of the pelletizing process, from concentrate beneficiation to baking of the pellets. The model has been validated by assessing its capacity to represent adequately the operation of the plant during a 20-month operating period. It was able to reproduce the plant capacity within a precision of 1.2%.Simulated “what-if” scenarios allowed quantifying the production increase associated with the addition of new equipment such as ball mills or balling discs, implemented individually or in combination. Results revealed a potential 10% increase in the annual plant production by adding additional equipment in the two most limiting sectors of the plant, grinding and agglomeration. The simulation results made it possible to quantify the synergistic effect of two projects on plant throughput; the model showed that the combination of two projects resulted in a greater production increase than the sum of the same projects evaluated individually.Now that it has proven reliable, the simulation tool can readily be used by ArcelorMittal Mines Canada process engineers to perform simulations for equipment optimization, and to evaluate the impact of various operating parameters such as equipment availability and tonnage, product mix or concentrate quality. This work showed that a reliable model can be developed based on known parameters in combination with the proper integration of the process knowledge."

Jan 1, 2010

By Mike Samuels, John Starkey

"This paper examines test results and grinding mill design for the recent Fortune Minerals’ NICO Project, located in the Northwest Territories. SAG milling and rod milling options are compared. Rod milling is clearly shown to be the best option because of the ore’s extreme resistance to breakage in a SAG mill environment.Surprisingly, the Bond Ball Mill Wi was 13 kWh/t for this ore. The Bond Rod Mill Wi was 20 kWh/t, whereas the SAG pinion energy to grind the ore to T80 1.7 mm was 30 kWh/t. The existence of ore with this ratio of SAG hardness to Bond BM Wi has not been previously observed. If not considered, this critical relationship may adversely impact the accuracy of mill design. This paper substantiates the significance of these findings to achieve successful grinding design on ore that is macro-competent during primary size reduction, but relatively normal during secondary grinding.INTRODUCTIONThe NICO Cobalt-Gold-Bismuth-Copper deposit is located 160 km northwest of the City of Yellowknife, Northwest Territories. It is the largest IOCG-type deposit (aka Olympic Dam-type) currently recognized in Canada with Mineral Reserves of 21.8 million tonnes contained within a larger 57 million tonne, lower-grade Mineral Resource. The project, owned 100% by Fortune Minerals Limited, has been assessed in a positive definitive feasibility study, subject to underground test mining and pilot plant testing, and is now in the permitting stage to develop a combined open pit and underground mine and process plant. An updated process flow sheet incorporates conventional crushing, grinding and bulk flotation, followed by filtration for shipment of a concentrate to a purpose-built refinery proposed to be built near Saskatoon, Saskatchewan. At the refinery, separate bismuth and cobalt concentrates will be produced following regrinding. The hydrometallurgical process methods include: pressure acid leach, precipitation, ion exchange, and electro-winning for recovery of cobalt as 99.8% cathode; solvent extraction and electro-winning of by-product copper cathode; ferric chloride leach and electro-winning for recovery of bismuth as 99.5% cathode; and cyanidation of the leach residues for recovery of gold as doré."

Jan 1, 2010

By E. S. Yan, R. Graley

"The StackCell was developed to overcome the engineering and design challenges that are commonly associated with existing flotation technology such as pneumatic, self-aspirating cells, mechanical cells, and flotation columns. These challenges include high energy consumption, large floorspace requirements, and size. They are most evident when designing systems to process streams containing a high volumetric flow rate coupled with a significant float yield. Unlike many metalliferous and industrial mineral applications, fine coal flotation circuits are typically characterized by feed streams containing a low solids content (<5%, by weight) where up to 85% of the feed can report to the product launder. With input from major engineering houses and end-users, the Eriez Flotation Division (EFD) developed the StackCell technology with the goal of providing an additional flotation option for end-users. This technology was designed specifically to overcome these challenges by providing an efficient method of separation in terms of metallurgy, but also with respect to equipment size, energy consumption, and installed costs. Besides being placed in tandem with existing cells to provide cost-effective supplemental capacity, this technology can be installed in a stand-alone arrangement as recently demonstrated at Patriot Coal’s Kanawha Eagle preparation plant. This manuscript provides an understanding of the underlying fundamentals of this new technology and presents data from Kanawha Eagle in addition to other industrialscale installations.INTRODUCTIONOver the past decade, column flotation has continued to gain acceptance as an alternative method for recovering fine coal due to superior metallurgical performance as compared to mechanical flotation cells. This point has been proven on numerous occasions by comparison of plant flotation data with the release analysis separation curve (Dell et al., 1972). Data from a release analysis test approximates the ultimate flotation response of a coal sample. In North America, a number of investigations have been published that document bottom-line improvements achieved using column cells (Luttrell et al., 1999; Kohmuench et al., 2004; Baumgarth et al., 2005). According to these reports, the benefits are derived from an overall increase in plant yield that can be achieved due to the improved product grade in the flotation circuit. In particular, the application of wash water to a deep froth utilized in column flotation minimizes the nonselective recovery of high-ash ultrafines that are hydraulically entrained in the froth by conventional flotation machines."

Jan 1, 2013

By T. Schmidt, J. Merza

"This study considers two lead/zinc/silver orebodies which have historically had a problem with pyrite, pyrrhotite and sphalerite contamination in the lead circuit. The problem was previously addressed with high dosages of sodium cyanide as the primary depressant. However, large amounts of cyanide led to environmental problems in the effluent, especially during winter months. The objective was to reduce cyanide consumption while maintaining or improving overall lead grade and recovery via the reduction of the undesirable sulphides.Extensive laboratory and mill testing was necessary to determine the optimal operating parameters for continuous plant application. Evaluation of data received from this testing has shown that selective addition and optimization of the dosage of Unimax SD100 to specific areas of the flotation and grinding circuits has resulted in improved Pb grade and recovery. Parameters such as pH control, regrind operation, the cyanide/Unimax ratio and sulphide mineral content may also affect the performance of the depressant. Each factor is discussed and optimal operating conditions were determined based on the testing.Results are presented for two case studies. Both indicate that Unimax SD100 improves the lead grade/recovery curve by increasing iron and zinc rejection while reducing cyanide consumption and the resulting environmental concerns. 1. Hart"

Jan 1, 1992

By M. Massinaei, S. Abasi, M. Oliazadeh

Reliable measurement and control of the pulp-froth interface (froth depth) in the flotation columns is critical to achieve the desired metallurgical performance. Shallow froths result in a reduction of grade because of particle entrainment and deep froths reduce recovery because of the drop back of entrained and weak hydrophobic particles to the pulp zone. The cleaner circuit at Miduk copper concentrator consists of three parallel MicrocelTM flotation columns 4 m in diameter and 12 m in height. The interface level of the flotation columns is measured by two pressure sensors located at 1.4 and 2.4 m from the cell lip (in the collection zone). In this study the performance of the level detection system at different operating conditions was evaluated. A calibrated float-ultrasonic location system was used for measuring the actual froth depth. The results indicated that the existing level control system was subject to considerable error. The absolute error was found to be 33.27 cm. This was mainly attributed to the constant set-point value assigned to the froth density (ρf=0.3 g/cm3), which actually varies with the operating conditions. The actual froth density at different conditions was calculated and its average value (ρf=0.12 g/cm3) was re-set in the control system. This resulted in substantial reduction in the level-detecting error. Recommendations for more accurate measurement of the interface level were presented. INTRODUCTION Froth flotation is a complicated process because of many involved variables. One of the most significant controlled variables in the flotation process is the pulp-froth interface (froth depth). The froth depth has a profound influence on the metallurgical performance parameters (i.e. grade and recovery). Shallow froths result in grade reduction because of particle entrainment and deep froths reduce recovery resulted from the drop back of entrained and weak hydrophobic particles to the pulp zone (Finch and Dobby, 1990). Various techniques have been developed for measuring the interface level of the flotation cells in the last few decades. These techniques can be classified as: Techniques based on hydrostatic pressure measurements (Finch & Dobby, 1990), Techniques based on the density gradient (float) (Moys & Finch, 1988), Techniques based on conductivity measurements (Uribe-Salas, 1991), Techniques based on temperature measurements (Moys & Finch, 1989).

Jan 1, 2019

By G. S. Dobby, J. A. Finch

"I. INTRODUCTIONA significant development in flotation over the past few years has been the increasing industrial use of flotation columns, primarily in Canada. The column differs dramatically from conventional mechanical flotation machines, both in design end operating philosophy, which has been a principal reason for its slow acceptance by the mineral industry. Boutin, Tremblay and Wheeler (1. 2) developed the concept in the early 1960's and since then plant and laboratory columns have been marketed commercially by Column Flotation Company of Canada Ltd. In 1980-81 at Mines Gaspe, Quebec, three columns replaced 13 conventional cleaners in a molybdenum upgrading circuit, with superior results (3). Since then many of the copper producers in British Columbia have installed columns for copper and molybdenum cleaning, including Gibraltar, Lornex, Highmont and Island Copper. There now appears to be widespread interest in flotation columns.The column is particularly attractive for applications involving multiple cleaning stages and can upgrade In a single stage compared with several stages of mechanical cells. This results in simpler, more controllable circuits. Importantly, the column itself is well suited to computer control. A significant problem, however, has been that of translating laboratory data to plant performance. The purposes of this communication are to address the problem of column scale-up and to describe some of the transport mechanisms of a flotation column, leading to a kinetics based model.A schematic of a flotation column is shown in Figure]. Industrial columns are up to 13 m high and 0.3 to 1.8 m in cross section (either square or circular). Three zones can be identified:a collection zone, with feed entering 2 to 3 m below the top of the column and descending against rising bubbles generated at a gas sparger; a washing zone, where a packed bubble bed is generated by a downward flow of wash water; and a conventional froth (2 - 5 cm thick), the sole purpose of which is particle transport to the launder. The tailings withdrawal at the bottom of the column is controlled at a rate slightly greater than the feed flowrate (called a positive bias).An important element is the washwater, added just below the froth zone. The component of the washwater that flows downward (to make up the bias) is sufficient to prevent coalescence of rising gas bubbles. The packed bubble beg so generated is very efficient at suppressing gangue entrainment to the concentrate."

Jan 1, 1985

By R. S. Simkus, J. R. Egan

"Changes have recently been applied to the Sullivan Concentrator lead and zinc circuits to treat fine-grained ores which make-up the majority of ore reserves in the mine. Pilot plant test work conducted in 1990-91 indicated that concentrate grades would be substantially improved by reducing the particle size (P,0) of the lead from 38 to 15 µm and the zinc from 45 to 24 µm.Circuit revisions involved installation of major new equipment including: two GUO kilowatt Vertimills for regrinding, a 4 metre diameter column flotation cell and a 4 metre diameter hyperbaric pressure filter for dewatering. Duties of the existing regrind mills were reassigned.IntroductionThe Sullivan ore body is a massive sulphide deposit consisting of fine grained mixtures of sulphides, sometimes inter-banded with country rock and varying widely in their content of lead, zinc, iron and silicates. Principal economic minerals are marmatite and galena. Iron is mainly present as pyrrhotite (85%), and to a lesser extent as pyrite (15%) .The orebody has been actively mined since 1923, with 1992 ore reserves estimated at 15,800,000 tonnes with an average grade of 4. 7% lead and 7. 9% zinc. Current plans call for treatment of 1,600,000 tonnes/year, (7,000 tonnes per operating day), with flotation feed grades of 5.9% lead, 10.1% zinc, and 22.0% iron."

Jan 1, 1994

By F. Flynn, P. Downey, G. Ritson

"Bell Creek Mill which has been in continuous operation since 1987 has undergone numerous operational changes in the past three years. The addition of over 2. 5 million tons of resource at over 13 git from Hoyle Pond Mine precipitated an expansion program for the Bell Creek Mill. This paper discusses the major changes that have occurred over the past two years to increase throughput and efficiencies. In addition, the refractory ore treatment options are also discussed.IntroductionThe Bell Creek Mill is located in the Timmins Goldcamp approximately 8km NE of porcupine Ontario. (Fig I.)The mill was originally constructed by Camamax in 1987 to treat the Bell Creek ore at a design throughput of 350tpd. The flowsheet (Fig 2.) consisted of two stage crushing, grinding, thickening, leach and drum filtration with gold recovery by Merrill Crowe.As ore depleted at Bell Creek, the Mill was purchased by Falconbridge to process ore from their Hoyle Pond property. Kinross purchased the Hoyle Pond property and Bell Creek Mill from Falconbridge in 1992.The successful exploration program at Hoyle Pond has led to several expansions and process design changes, with further changes planned over the next twelve months.In 1994, the mine produced 61,865 oz. of gold at an average cost of 197US$ per oz. In 1995 production was 91, 611 oz. at a cost of 187US$ per oz. Projecting the nine months ended September 30, production for 1996 will be 155,000 oz. at a cost of 165US$ per oz.In early 1997, the CIP expansion will be complete, with a target of 175,000 oz. at a cost of 176US$ per oz for 1997. The mill throughput will have increased from 400tpd to 1500tpd between 1994 and 1997.The changes to Bell Creek since 1994 have been numerous with very little of the existing equipment still in operation. The total cost of the expansion to date is 15 million C$. This paper discusses the major projects carried out during in this time. The focus will be to highlight the justification of each project, the design and engineering, some of the initial start-up problems and solutions.Figure 3 shows a plan of the existing mill facility together with the various additions that were carried out."

Jan 1, 1997

By David Estrella, Hector M. Lizama, Jhonny Carrión

"Microcel™ sparging systems were installed on one column of the molybdenum cleaner circuit and one column of the copper first cleaner circuit at Antamina. The Microcel™ system is known to generate finer bubbles than jetting-type spargers, and smaller bubbles have been associated with higher recoveries in flotation. Alternating tests were carried out on the molybdenum column, using either jetting-type spargers or the Microcel™ system. Experimental conditions of gas flow rate, wash water rate, and froth depth were tested. Use of Microcel™ decreased the bubble sizes from 2.6-2.9 mm to 1.5-2.0 mm. In addition, Microcel™ improved both unit molybdenum recovery and unit molybdenum concentrate grade. In the copper first cleaner circuit, side by side tests were carried out that compared the Microcel™ column with a parallel columns fitted with jetting-type spargers. Best results were obtained at manifold pressures of 260-290 kPa, where bubble sizes of 1.6-2.6 mm were observed. In either copper or molybdenum flotation, the Microcel™ system required only 50-60% of the gas flow used by jetting-type spargers.INTRODUCTIONAntamina is a copper-zinc mine located in the Peruvian Andes. Compañía Minera Antamina is owned by BHP Billiton (33.75%), Xstrata (33.75%), Teck Cominco (22.5%), and Mitsubishi (10%). The mine was commissioned in 2001 with a capital investment of US$2.3 billion, and has a nominal 100,000 t/d mill capacity. Mining, milling, flotation, and tailings impoundment take place at the mine site, which sits at an altitude of 4,300 m above sea level. The deposit is a polymetallic skarn ore body with intrusive phases of quartz-monzonite porphyry into Cretaceous aged limestone [Evans et al, 2005]. Ore body thickness is 100 to 150 m and has the shape of a flat-lying crescent. Copper is present mainly as chalcopyrite with a minor bornite component. Zinc is present as sphalerite. Molybdenum is present as molybdenite. Some lead, bismuth, and silver are associated with the sulphides. The deposit is currently classified into six main ore types [Dechert, 2006], grouped into two categories: copper and copper-zinc. The mill produces four concentrate types: copper, molybdenum, zinc, and lead-bismuth. Copper, zinc, and lead-bismuth concentrates are transported by pipeline down to the port of Huarmey. Molybdenum concentrates are bagged and trucked to port."

Jan 1, 2008

By Colin Hardie, Simon Pelletier, Rakan Rhabani, Guy Comeau

"In order to improve the nickel concentrate thickener operation at Raglan Mine, two pulp deaeration cyclones were installed. These units obtained from Outotec Pty Ltd, also called the FrothBuster®, have been in continuous operation since mid-October 2006. The primary objective of the de-aeration cyclones is to lower the solids content in the thickener overflow by de-aerating the concentrate pulp before it enters the concentrate thickener. This dearated slurry should help decrease the nickel particle losses that report to the thickener overflow because of the flotation effect within the thickener.Optimizing the de-aeration cyclones has been a slow and ongoing process, however, initial testing is encouraging. The latest test results indicate a 50 % reduction (0.070 to 0.035 % solids) in nickel losses to the concentrate thickener overflow resulting in an additional $350,000 in annual revenues.This paper will present an overview of installing the de-aeration cyclones within a limited space, as well as the preliminary commissioning and testing challenges. This paper will also discuss the most recent concentrate thickener operating results since the de-aeration units were installed.INTRODUCTIONThe Xstrata Nickel Raglan Concentrator, located at the northern limit of Quebec’s Nunavik Region has been in production since 1997. In early 2002, the concentrator implemented the Zero Process Water Discharge (ZPWD) project to recycle the process water to decrease fresh water consumption and reduce discharges to the environment. Due to this initiative, the chemistry of the process water changed significantly resulting in flotation concentrate streams, which were very frothy and difficult to manage. This phenomenon has resulted in operational challenges in many areas of the concentrator including slurry pumping, flow measurement and concentrate thickening.In October 2005, the Raglan Mine underwent a conversion from an AG mill to a SAG mill operation in order to increase throughput and to optimize key process areas. As part of this project, the first two North American pulp de-aeration (FrothBuster®) units from Outotec were installed to improve the operation of the concentrate thickener. The primary objective of the deaeration technology was to lower the solids content in the thickener overflow by de-aerating the concentrate thickener feed pulp. This in turn was to reduce the froth, which accumulates on the thickener surface, and more importantly reduce valuable nickel losses to the thickener overflow because of the flotation effect within the thickener."

Jan 1, 2008

By J. S. Stickling, P. J. Whittaker, N. O. Lotter

"Assessment of opportunities to improve mineral processing performance in existing concentrators and in describing undeveloped orebodies is key to improving the business case of a mining company. This assessment has traditionally been performed using the distinct and separate inputs from mineral processing and mineral science.More recently, process mineralogy has been used to integrate these two disciplines. This trend has been driven by the need to develop and process more challenging ore deposits, and has been assisted by the availability of modern automated instruments. Together with the acquisition of a QemSCAN (quantitative evaluation of materials by scanning electron microscope) from CSIRO in Australia, and with the use of a supporting SEM with EDX/WDX capabilities and an XRD, the development of process mineralogy at Falconbridge commenced in 1997.The sampling challenges in presenting a QemSCAN with representative sample material at the polished section level are significant, but have been effectively addressed. The full description of this aspect is beyond the scope of this paper.It is shown that, by use of this approach, the set nature of an orebody can be best exploited with selected flowsheet changes in a concentrator which address the generic characteristics of the orebody. In this way, the flowsheet provides a process that more effectively fits the mineral processing requirements of the ore. This leads to clear and tangible improvements in grade and recovery."

Jan 1, 2002

By Michel Robert

"In 1974, Copperfields Mining Corporation and Soquem formed a new company, Niobec Inc, to bring the niobium deposit of the St-Honore carbonatite into production.St-Honore is located in the Saguenay area, near the cities of Chicoutimi and Arvida, and about one hundred and forty (140) miles north of Quebec. The discovery of the carbonatite complex was made in the fall of 1967 as a result of an airborne radiometric survey in search of uranium. The detailed exploration of the carbonatite indicated two niobium-bearing formations besides the original discovery of a rare earths zone. During the following years, the niobium formations were defined by diamond drilling and a decline was driven down to obtain adequate samples for metallurgical evaluation. In 1974, after seven hundred (700) bench scale tests, eleven (11) months of pilot plant 2nd a worldwide market research, a joint decision was made to initiate the development of a fifteen hundred (1500) T.P.D. mine and mill, under the management of Teck Corporation. The construction of the facilities were completed in early 1976 both on time and within budget. Niobec is now the only north american and the world second largest producer of Niobium concentrate.The process used at Niobec was developed from the pilot plant programs and later modified by the mill personnel. A simplified flowsheet for this process is shown on the slide. Briefly, run of mine ore is crushed to minus 5/8"" and fed to a rod mill, ball mill, and screens circuit where the ore is ground to 92% passing 65 mesh. The ore is deslimed in two stages of cycloning, and the final cyclone overflow at 85% minus 11 microns goes to waste. The cyclone underflow is conditioned prior to flotation of a carbonate-apatite concentrate. This concentrate is cleaned twice and then pumped to tails. The carbonate rougher and cleaner tails are cycloned in two stages to change the process; water and then sent to the pyrochlore rougher flotation. The rougher concentrate is cleaned five times. This is followed by pyrite flotation to remove sulfur, leaching with hydrochloric acid to remove phosphorus and dyed to less than 0.1% moisture before being sent to packaging."

Jan 1, 1978

By G. R. Morin, A. R. Leggett

"In 2010, closure of the Kidd metallurgical plants along with structural changes to the historical ratios of payable metals led to a review of the operating targets at the Kidd concentrator. Previous operating data indicated that there was potential for significant revenue improvements by lowering the copper concentrate grade. An extended plant trial was carried out in 2011 to quantify the revenue impact of lowering the copper concentrate grade from 27% to 25% Cu, as well as identifying the lower practical grade limit. Analysis of plant trial results indicated that the process change increased net concentrate revenue by $2 M to $3 M per year at current metal prices. Increased revenue was primarily associated with higher silver recovery to the copper concentrate while the marginally improved copper recovery was essentially offset by increased zinc losses and concentrate treatment and freight charges.Details of the plant test work results are discussed in this paper along with the financial analysis of the revenue improvements. Key areas covered are major revenue components of the concentrate contracts, validation of revenue improvements through plant testing and sensitivity of the improvements to market and plant variables.INTRODUCTIONThe Xstrata Copper Kidd concentrator commenced operations in 1966 treating ore from the nearby Kidd volcanogenic massive sulphide deposit in Timmins, Ontario (Thompson, Amsden, & Chapman, 1978). Over the years, up to four separate grinding-flotation divisions were in operation with throughput rates peaking at 4.5 million tonnes per year of ore in the mid 1980’s (Kelebek & Francis, 1998). As the mine production decreased and grinding throughput increased operations were gradually cut back and since 1998 two grinding-flotation divisions have been in operation treating Kidd mine ore. One of the idled divisions has been intermittently used for custom milling operations, most recently the Ni/Cu Montcalm deposit, which was milled from 2004 to 2009. Current budget throughput rates, limited by mine production, are 2.4 million tonnes per year of ore grading 1.7% Cu, 4.3% Zn and 55 g/mt Ag. Concentrate production rates are 150,000 tonnes per year of copper concentrate and 155,000 tonnes per year of zinc concentrate. The Kidd mine is the deepest base metal mine in the world with mining operations down to 9600 ft.In 2010 the copper and zinc plants at the Kidd metallurgical site, which previously treated the concentrate production from the Kidd concentrator, ceased operation due to a variety of economic and market factors. Following the metallurgical plant closures the Kidd concentrator products were shipped off site for treatment. The majority of the copper concentrate is now processed at the Xstrata Copper Horne smelter in Rouyn-Noranda, Quebec with the majority of the zinc concentrate processed at Canadian Electrolytic Zinc Ltd. in Valleyfield, Quebec."

Jan 1, 2013

By Bjarne Ekberg

"The most economical way of dewatering mineral concentrates is to reduce the moisture content in the material mechanically using thickeners and filtration avoiding thermal drying whenever possible. In addition to this, the dewatering process should produce low moisture cake, high quality liquid, consume little energy. As a part of the dewatering process The CERAMEC® capillary filter has been proven to meet these requirements in over 150 installations. The filtration is done by using ceramic filter plates with small pore size. The capillary action in the filter medium secures gas free filtration at good vacuum, and due to the small pore size, filtrate is crystal clean and particle free. The energy consumption is minimal. This paper will discuss results from filters installed at Pyhäsalmi Mine Oy in Finland where the technology was has been used since the late eighties. Results from more recent installations like Los Pelambres in Chile and LKAB in Sweden will also be discussed.INTRODUCTIONThe idea of using capillary action to make dewatering more efficient was invented two decades ago. Ten years later whole filter plants utilizing the new method were delivered to mine concentrators for dewatering of mineral concentrates. Further development of ceramic membrane plates made it possible to meet the requirements in iron ore dewatering. In addition to low cake moisture, high capacity and low energy consumption, there are a lot of environmental advantages in using capillary filtration. One of them is clear filtrate.CAPILLARY FILTRATIONA micro-porous filter material behaves like a large number of narrow capillaries forming a network. When a pressure difference ?P is applied over the wetted, hydrophilic pore with the radius r, the force striving to push the water out of the pore will be ?Ppr2. The force striving to hold the water in the pore will be the vector force 2pr? cosa of the surface tension ? and the cosine contact angle a. Equalling those forces and solving the equation for ?P we get"

Jan 1, 2003