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By P. W. Godbehere, B. W. Brooks

"This paper deals with the milling of slags from the Noranda Continuous Process, in the Horne Concentrator at Noranda, Quebec. The milling of copper slags, although novel in Canada, has been practiced in Japan, Finland and elsewhere for several years.In March 1973, the prototype Noranda Continuous Process furnace or ""reactor'' came on-stream, In this pyrometallurgical process (see Fig. 1), copper concentrates mixed with flux are smelted in a single furnace to form copper or a high grade copper matte, and slag. The conventional smelting processes of roasting, reverberatory smelting and converting have been incorporated in a single 17 ft. diameter by 70 ft. long steel cylindrical vessel (l, 2). The copper or matte is tapped from the bottom of the reactor into ladles and transferred to an anode furnace or to a converter. The slag is tapped into 40 ton capacity ladles, cooled, crushed, and conveyed to the concentrator. At the present time, the reactor is operated to produce matte. The matte slag* containing 4 - 6 per cent copper is milled using conventional ore dressing grinding and flotation equipment. A flotation concentrate assaying 35 - 40 per cent copper is recycled to the reactor. Crushing, conveying, grinding, flotation and tailing disposal equipment previously used for over 40 years for ore treatment has been modified to suit the requirements of slag milling. Slag milling is essentially the scavenger circuit of a pyrometallurgical process."

Jan 1, 1977

By C. Veloo

"Westmin Resources operates a mine/mill complex at Myra Falls and produces copper and zinc concentrates with minor associations of gold, silver, and lead from a polymetallic, massive sulphide orebody. Mill flotation performance has historically been plagued by periods of uncontrollably high circulating loads leading to low concentrate grades and below optimal tails. This led to a detailed examination of the flotation process. Locked cycle, and pilot plant data are presented. In addition, plant design and commissioning as well as column cell details are discussed. A comparison between actual and predicted performance is also presented. Benefits realized are higher copper and zinc concentrate grades and improved zinc recovery.PrologueMyra Falls Operations is located at the south end of Suttle Lake in Strathcona Provincial Park. Access to the mine site is provided by a 90km paved highway from Campbell River, British Columbia.The mine and facilities are nestled in the narrow Myra Valley between Mount Phillips and Mount Myra. These steep, rugged mountains rise from the valley floor at an elevation of 900 feet to summits of almost 6000 feet. The steep slopes are heavily wooded with fir, hemlock, and cedar up to an elevation of 4000 feet where a transition to alpine meadows occurs. The summits are craggy rock faces, some capped with glacial ice. Myra Creek winds its way through the valley dropping several hundred feet into Suttle Lake. Precipitation is high at over 100 inches per year, which may include up to 10 feet of snow in winter. The temperature may range from 32° C in summer to -18° C in winter.Myra Falls began production in 1965 with the Lynx Mine, supplemented by the Myra and Price Mines in subsequent years, feeding a 950 stpd mill up until the middle of 1985 when the H-W mine was brought into production. The original mill was replaced with a modern 2700 mtpd facility and in 1988 production was increased to 4000 mtpd.Both Lynx and H-W mines are currently in production producing an average of 3650 mtpd. The ore minerals are primarily sphalerite and chalcopyrite, with subordinate galena, bornite, tetrahedrite, and electrum."

Jan 1, 1992

By Rudy Tenbergen, Dick Farmery, Jim Naylor

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

Jan 1, 1997

By B. Hart, J. Marois

"Scanning electron microscopy coupled with energy dispersive X-ray spectroscopy (SEM–EDX), and Time of flight secondary ion mass spectroscopy (TOF-SIMS)were used to analyze pyrochlore grains from Niobec froth flotation plant. Samples from the mill feed, pyrochlore rougher feed, pyrochlore rougher concentrate, and tail were analyzed in order to identify a potential relationship between pyrochlore matrix composition and selective separation. SEM-EDX analyses indicate that pyrochlore grains with high Fe content appear to be less recoverable than those with a lower Fe content and that the flotation response is related to matrix Fe rather than Fe occurring as inclusions. TOF-SIMS was used to analyse the surface of high Fe and low Fe pyrochlore, in order to identify a potential relationship between pyrochlore matrix composition and the related effect on cationic collector adsorption (tallow diamine). Surface analyses of pyrochlore from a conditioning test show that species indicative of the collector favour the surface of Fe poor pyrochlore relative to the Fe rich variety. The results from the conditioning test illustrate a negative relationship between collector loading and the presence of Fe and related oxidation species. This study identifies the link between pyrochlore matrix Fe content and, the presence of surface oxidation products, to collector loading. Poor Fe rich pyrochlore recovery noted in the Niobec froth flotation plant may be partially explained by this mechanism.INTRODUCTIONPyrochlore, the most abundant niobium containing mineral (Perrault and Manker, 1981) is economically recovered in Brazil, China, and Canada (Niobec) (Kouptsidis et al., 2008). Since 1974 Niobec has been extracting niobium from the St-Honorécarbonatite deposit. Niobec is the only operating Nb extraction facility in North American and one of the three major world producers of ferroniobium (Dufresne and Goyette, 2004; Belzile, 2009).Flotation is the preferred and most cost-effective method for pyrochlore recovery (Burks, 1958; Shapolov and Polkin, 1958; Zakharov et al., 1967; Razvozzhaev and Nikiforov, 1972; Bogdanov et al., 1973; Abeidu, 1974; Raby and Desrochers, 1975; Wilson, 1979; Biss and Nadeau, 1982; Biss, 1985). With the aim of optimising the flotation process for the Niobec ore, numerous experiments were performed on the influence of surfactants and type of water to optimise pyrochlore recovery and grade during flotation (Gomez et al., 1987; Rao et al., 1988; Dufresne and Goyette, 2004; Belzile, 2009)."

Jan 1, 2013

By Fernando Romero Lage

The purpose of this paper is to analyze detailed crushing and screening configurations and the interrelationship as they relate to mass flow balance, operating efficiency, plant availability and final production. The entropy of ore bodies limits the standardization of crushing and screening plants, however, consideration for a continuous and sensible assessment of the mass flow balance of the circuit will yield the best achievable results. Since energy costs have continuously risen in the last few years, it is imperative that these circuits offer low power requirements and low operating expenses. The paper will detail how the efficient implementation of continuous mass flow balance in the application of crushing and screening plants will generate quality improvement and process improvement, which will translate into bigger production and energy savings.

May 1, 2010

By Daniel D. M. Adams

Over the last 100 years industry in general has been trying to advance the ability to maintain systems through management processes. Mining has looked to utilize onboard data systems in our mobile mining fleets to compliment these processes. Combining the two concepts build the concept that Rio Tinto refers to as asset health. These programs can generate a cost shift between $20,000 and $100,000 per piece of equipment annually. Even with the documented savings many operations struggle to justify the systems, why?

May 1, 2011

By R. J. McLeod

"CONTROL STRATEGYThe Lake Dufault Division of Falconbridge Copper Limited has been concentrating a complex copper zinc ore since late 1964. An ARL X-ray on stream analyzer was installed in 1966 for process control. A Digital Equip-ment Corporation PDP8 was added in 1968 for more complete control. To improve the system the computer was exchanged for a Moded PDP-81 in 1971.The new computer is time sharing with 8K core memory and 32K disc storage. A new interface was installed to accommodate 14 X-ray streams and additional analog signals for process control. The system gave us additional capacity and flexibility to make changes in operating programs and control strategy while on line. As the computer is used for process control a modified focal language was chosen as the operating language for ease of operation. This language is more suitable for process control than for statistical work.The main components of the control system are(a) Computer(b) Interface(c) Production Control X-ray Quantometer(d) Teletype terminals(e)Reagent feeders(f) Sample SystemThe operation of the system consists in the computer selecting the various streams to be analyzed by a PCXQ Quantometer. When the flowing sample is presented to the X-ray the intensities of the various elements are read by the computer from analog to digital converter, the assays are then computed. From the tonnage and assays the computer is programmed to determine the amount of reagent added to various parts of the circuit. The computer drives the feeders via the interface for the required amount of reagent. After following a certain sequence the computer prints out the performance and control data on an ASR35 Teletype. The data is printed out approximately every seven minutes. Any des¬ired changes in the program or strategy can be made during operation via an ASR33 Teletype connected to the computer."

Jan 1, 1973

By R. Patrick

As a unique South African organisation, CSIR: Division of Mining Technology (Miningtek) strives to support the mining industry by improving health, safety and productivity through developing, adapting and transferring technology. Considerable effort is currently directed towards the development of potential non-explosive mining methods with the potential to allow continuous mining in narrow reef conditions. Besides improving productivity, continuous mining results in concentrated mining which has a significant impact on reducing infrastructural costs. Impact Ripping and Water Pulse Rock Breaking have been tried on gold mines, whereas Activation, Mini-Disc? cutting and the Plasma Holemaker are currently being tested on platinum mines. The organisation?s many years of experience in novel rockbreaking, together with its world-wide collaborator network, has placed it in a unique position to assist the mining industry and manufacturers in developing and/or adapting existing technologies for application underground. Currently a range of technologies are being investigated, which have reached different stages of maturity. The main technologies under investigation are; - Impact ripper system - Water pulse rock breaking - Activated undercutting rollers - Mini disc cutters - Electric rock breaking

Oct 1, 2001

By Edwin G. Sharp, Frank S. Gaunce, George M. Freeman, Denis J. Kemp, Eric Dulson, Paul Salmon, Ernest R. Hamilton, Carl E. Parden

"THE ECSTALL ZINC PLANT was built for the treatment of the Kidd Creek zinc sulphide concentrates. Construction proceeded throughout 1970-71, with operations beginning in April 1972. The two concentrate types treated differ principally in minor impurity and precious metal contents. Processing flexibility was required to allow efficient recovery of the by-products, especially cadmium, copper and silver. The conversion of the zinc sulphide concentrates to end products follows a sequence of operations that are common to most electrolytic zinc plants and typical of the ""Tainton- Jephson"" flowsheet.1. Roasting in air at an elevated temperature to convert the metal sulphides to soluble metal oxides (calcine) and sulphur dioxide. The sulphur dioxide is converted to sulphuric acid in a contad plant and shipped by train to market.2. Leaching of the calcine in spent electrolyte to effect extraction and recovery of zinc as zinc sulphate solution, leaving an insoluble residue.3. Purification of the clarified zinc sulphate solution by cementation with metallic zinc dust to remove impurities detrimental to the electro-deposition of zinc or detrimental to product purity. The solids from cementation are the starting materials for secondary metal recovery.4. Electrolysis of the purified zinc sulphate solution to reduce the zinc to the metallic state while coincidentally regenerating the sulphuric acid for recycle to the leaching step."

Jan 1, 1974

By Y. Shu, A. R. Laplante

"A laboratory Knelson Concentrator (LKC), 7.5 cm in diameter, was used to study the performance of a similar unit at plant scale, 76 cm in diameter, This paper reports on sampling and assaying procedures and problems, and discusses the performance of the plant unit.The sampling procedure was based on the extraction of large samples, 5-15 kg, and their complete processing with a LKC, to characterize size-by-size performance. The procedure yielded reliable estimates of overall gold assays, after blatant assaying errors were corrected. The plant centrifugal unit concentrate was sampled in situ, which yielded very erratic grades, with a large overall negative bias. Due to the very heterogeneous nature of the concentrate, which can be linked to flow patterns in the riffles of the unit, it is recommended that for future testwork, the concentrate be recovered separately, blended and then sampled.The plant centrifugal separator, when operated at a feed rate of 25 t/h, yielded total and free gold recoveries of 41 % and 62%, respectively. This is much lower than previously reported data, and suggest that at 25 t/h, the unit is overloaded. However, recovery did not decrease with time over a loading cycle of 2 hours, which suggests that present loading cycles, generally 1 hour, can be significantly increased. INTRODUCTIONThe study of gold gravity circuits is plagued with the statistical problem of determining the behaviour of particles occurring with a very low frequency. This problem becomes particularly acute with increasing particle size. For example, considering the average weight of a gold flake in the 840-1200 µm (20/28 mesh) to be about 5 mg (Banisi et al, 1991), a reasonably high grade feed of 10 git would require, for a relative error (standard deviation) of 10% on grade, 100 particles, that is 0.5 g of gold, or 50 kg of 840-1200 µm material (a Poisson distribution is assumed). Gy's formula, for the same data, yields a weight of 33 kg, with a shape factor of 0.1 and a size distribution factor of 0.5. Both results are comparable; to obtain the suggested weight of material, a stream containing 5% weight in the 840-1200 µm would require 1 tonne samples (of unsized material). The problem rapidly disappears with decreasing particle size. Still with Banisi et al's data, a gold flake in the 300-420 µm range averages 0.5 mg in weight; the same 10% error on grade corresponds to a weight of 5 kg. Gy's formula for the same size distribution and particle shape factors as above predicts a required mass of 1.4 kg. This would correspond to an unsized sample weight of 10 kg with 14% weight in the 300-420 µm, for example."

Jan 1, 1992

By D. M. Wyslouzil

"This paper was prepared for presentation at the 16th Canadian Mineral Processor's Annual Meeting, Ottawa, Ontario, Jan. 17 to Jan. 19, 1984.This paper gives an account of the development of treatment procedures for the differential flotation of refractory copper-zinc ores. Special references are made to the effectiveness of various zinc depressants and their combinations.Studies carried out on refractory copper-zinc ores demonstrated that conventional zinc depressants such as cyanide and zinc sulphate were not particularly effective when secondary copper minerals were present in the ore; the flotation activity of sphalerite was enhanced in the presence of cyanide.Among the various zinc depressants evaluated, the following depressant combinations were found most effective: Zn(OH) 2-S02 , 802-CaO, Na2S-Na2S03-S02, Na 2 S-S02 and Na 2S-S0rDS22.The practical application of any of these systems depends on the type of ore treated and other conditions such as pH, levels and points of collector and depressant additions.In this paper the effect of each depressant combination on different ores is described with reference to the selection of collectors."

Jan 1, 1984

By K. V. Konigsmann

"The results of Mr. Klymowsky's research confirm the essential role played by oxidation products in the flotation of base metal sulphides. They invite a discussion on the subject of aeration which I may open by a brief description of the development and current practice of aeration. The first installation of pre-flotation aeration in Canadian milling practice was undertaken during the years 1928 and 29 at the then new Noranda Mill. C.G. McLachlan and his collaborators demonstrated that the flotation differential between chalcopyrite and iron sulphides was enhanced by aeration in soda ash alkaline pulps. The ores treated contained more than 80 per cent sulphides. It was proposed already then that the products of mineral oxidation assisted copper and retarded iron sulphides, at least pyrrhotite flotation. The concentration of thiosulphates was used as a measure of the progress of oxidation. Several papers by C.G. McLachlan elaborate on this subject.The aerator was also used as classifier since it was held that aeration, i.e. controlled oxidation, should go hand in hand with creation of new mineral surfaces. Grinding mills and aerators were placed in closed circuit. Figure 1. This aeration was essential for successful separation at Noranda.Following the application of aeration at Noranda the technique was applied to existing mills in N.W. Quebec. Impressive improvements in concentrate grades and recoveries were obtained. It was observedthat the zinc flotation benefitted from aeration as much as the preceding copper float. In the late forties and fifties the aerator classifier was installed in new mills associated with the Noranda group. A diagrammatic sketch of the Quemont flow-sheet (figure 2) may serve as a typical example."

Jan 1, 1973

By Peter A. Eshun

This paper studies two methods that account for risks associated with gold mining projects such as mineral risks and country risks. The first method is sensitivity analysis, which investigates the effects of a change in any economic parameter such as operating cost, capital investment, gold price or taxes on the viability of the project. The second method is Monte Carlo Simulation technique, which takes into consideration the simultaneous random variation of the economic parameters on the project's viability. An automated simulation cash flow model has been developed as part of this paper. Sensitivity analysis and the simulation model have been used to investigate the economic viability of the Dokrupe Gold Project (DGP) serving as a test case. It is concluded that, the Monte Carlo Simulation technique gives a more realistic judgement of the economic viability of gold mining projects.

May 1, 2004

By A. C. T. Bigg

"The mineral potential of the Nanisivik area was discovered as far back as 1910 when a prospector of the Geological Survey of Canada, on an expedition under Captain J. G.Bernier, recorded a deposit of pyrite with minor sphalerite and galena. An outcrop of the present orebody was discovered by Texas Gulf Company in 1957. Subsequent drilling outlined the orebody consisting of 6,000,000 tonnes of ore containing 13 % Zinc and 1.5 % Lead. The orebody appears to have been formed by the filling of a limestone cavern with sulphides. The sulphide minerals consist of pyrite, sphalerite and galena. The limestone bedrock is mostly dolomite with some calcite and magnesite. A cutoff grade of about 6% Zinc has been calculated.In 1972 Mineral Resources International Limited took an option on the property. Upon completion of the feasibility study by Watts, Griffiths and McOuat, the then new independent project management group of Strathcona Mineral Services Ltd. were engaged to develop the project through the financing, construction and operating phases. Strathcona Mineral Services still manage the project from their head office in Toronto and through the Mine Manager and Department Superintendents at the Mine site. The interest for the project is presently split five ways: Mineral Resources International Ltd. 53.5 %, Government of Canada 18 %, Metallgesellschaft A.G. of Germany 11.25 %, Billiton B. V. of Holland 11.25 % and Texas Gulf Co. 6 %."

Jan 1, 1980

By W. D. Gould, A. Kapoor, K. Morin, P. Bédard

This paper presents the results of a research study that is being conducted by the Mine Effluents Program, Mining and Mineral Science Laboratory (MMSL), Natural Resources Canada aimed at evaluating biotechnology processes for the treatment of nitrogen compounds such as thiocyanide (CNS) and ammonia (NH4-N) which are present in gold mill effluents. A sequencing batch reactor (SBR) technology, commonly used for the biological treatment of municipal and industrial effluents, was used in this study. In the SBR process, the micro-organisms were able to degrade CNS to NH4-N and NH4-N to nitrate (NO3-N) at operating conditions of two 12 h treatment cycles per day, with pH maintained in the 7.4 to 7.6 range, and at room temperature (approximately 21°C) and also at 12 °C. The end products of CNS and NH4-N biological oxidation were NO3-N and sulphate (SO4) that are relatively non-toxic. Partial removal of NO3-N was achieved by biological denitrification reactions in the SBR process. The SBR process effluent was measured to be non-toxic to rainbow trout based on the 96 h acute toxicity test. The microbial consortium isolated from the SBR treating a simulated effluent was able to effectively oxidize CNS and NH4-N to NO3-N in water samples (under batch conditions) collected at three mine sites located in Quebec, the Northwest Territories, and Yukon.

Jan 1, 2004

By Jessica Egan, Daniel Hodouin, Claude Bazin

"The recovery of gold by leaching is influenced by the size of the gold carrier particles and the chemical environment. Using the relationships between the leaching rate, infinite recovery and particle size for a given set of chemical conditions, it is possible to simulate gold recovery for various feed size distributions. However the use of these particle size relationships is only valid if the leaching of gold from a given size class is independent of the production history of this size class. For instance, to use these relationships it is necessary to verify that gold recovery from the leaching of the 200-270 mesh size class remains the same if the size fraction is taken from a coarse or finely ground ore sample. Several laboratory tests were carried out to answer that question. Results of the tests show that within the large error margin associated to gold content measurements, the history of the production of a size class does not affect the gold leaching from that size class. Emphasis of the discussion is put on the difficulty associated with the processing of data containing significant measurement errors as it is the case for gold assays of leaching tests.INTRODUCTIONMore than 75% of the gold produced in the world is extracted by cyanide leaching of gold ores. The ore is ground to expose the gold carriers and the ground material is reacted in an oxygenated cyanide solution that dissolves gold, leaving solids gangue minerals that are disposed to tailings ponds or returned in underground mines as paste fill. The gold contained in the solution is recovered by either the carbon in pulp process or by cementation onto zinc powder, before being melted and casted into a brick (Marsden and House, 2006)."

Jan 1, 2009

"It is projected that the reactivity of acid generating tailings and waste rock can be minimized by reducing the access of oxygen by placing these materials under water in natural or manmade lakes. A project under the MEND and B.C. AMO Task Forces has been in progress for four years to evaluate active and historic tailing sites in B.C. and Manitoba. The sophisticated geochemistry studies conducted on the interstitial pore water has illustrated encouraging results which are discussed. The additional studies which are needed to complete the scientific database for the study sites, for achieving regulatory and public approval and for the engineering of application of this method of control are outlined.IntroductionWorld wide research effort is in progress to find the solution to the acid mine drainage problem. A great deal is known about acid mine drainage but much work is required to find economic and effective approaches to the problem. Two coordinated programs have been initiated recently in Canada: the National Mine Environment Neutral Drainage (MEND) program and the British Columbia Acid Drainage Task Force.Disposal of tailings materials under a water barrier such as a natural lake, is thought to be one solution to prevent acid generation by preventing oxygen and bacterial action on the sulphide surfaces. The deposition of reactive tailings materials under water in natural lakes has been used historically at various mines in Canada. However, there has never been a thorough review of the long and short term impacts of historic sites which may provide crucial information to understand the physical, chemical and biological processes which are taking place.This paper provides a general overview of a series of studies which examine the general biophysical characteristics and the sediment geochemistry of four lakes where subaqueous disposal has been or is currently being used for acid generating tailings. The subaqueous disposal studies have been conducted from 1989 through 1991 and were initiated and supported by both research task forces. The objectives of this paper are to summarize the history and the background information of the studies conducted, the general results obtained from each lake, an overview of the major collective results and the additional work which is required to achieve the final objective of developing criteria for applying this technology in the industry."

Jan 1, 1992

By K. L. C. Gonçalves, V. K. Alves

"CVRD’s Department of Development and Technology – GETEK has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. This methodology has been applied in several prospective CVRD mining projects. It encompasses a procedure for variability analysis based on bench scale tests, pilot scale grinding and flotation tests and modeling and simulation for scaling up the industrial flowsheet.The bench scale tests include the determination of the Work Index, simplified DWT tests, rougher flotation and locked-cycle tests with samples that are representative of the ore body variability. After completion of these studies, continuous, pilot-scale flotation tests are carried out at a Mini Pilot Plant with drill core samples representative of different stages of the mine life. Finally, studies are carried out with large volume samples using a conventional grinding and flotation pilot plant. This sample is also used for liberation characterization, grinding kinetics, flotation kinetics and industrial circuit scale up. This methodology has been successfully employed in the development of CVRD’s new mining projectINTRODUCTIONCVRD has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. The definition of the processing route and the sizing and scale up of the processing system are some of the products generated during the pre-feasibility of a greenfield project. This paper is focused on the scale up of grinding and flotation circuits, operations known to be of the most importance in a mineral processing plant.The methodology for the definition of the flotation and grinding circuits is divided in the following stages: bench scale ore variability studies, pilot plant operation and modeling followed by simulation. The product of this work is a calibrated simulator that can be utilized for plant operation as well as mine planning, looking to achieve optimal operational conditions.The current models used for scale up of the grinding circuits are aimed at the generating information about the size distribution of the circuit product. However, it is well known that the liberation state of these products may be of critical importance for the performance of the concentration processes downstream. Nevertheless, CVRD has been developing a methodology for liberation kinetics during grinding, aimed at optimizing grinding circuit throughput together with downstream flotation recovery."

Jan 1, 2006

By Ernie Doell, Trung Nguyen

"In early 1995 Cluff Mining commissioned a newly constructed high density tailings thickener at its Cluff Lake uranium mill. The 85 ft diameter high density thickener is one of the first of its kind in North America, and incorporates the novel paste technology in its design to produce densified (in the range of 60 -65% solids) mill tailings suitable for stable deposition in the tailings management area. This paper discusses the impetus behind the installation of such an unconventional thickener, and the many trying experiences in making it operational.IntroductionCluff Mining, a wholly owned subsidiary of Cogema Resources Inc., is the operator of a uranium mine and mill complex at Cluff Lake which is located in northern Saskatchewan. Since 1980, the company has developed and mined out several different ore bodies within the Cluff Lake region for the production of uranium concentrates, commonly known as yellow cake. Based on the known ore reserves, the Cluff Lake operations are projected to last till the early 2000's.Tailings from the Cluff Lake mill have been discharged to a tailings management area consisting of a tailings pond and two facilities for primary and secondary treatment systems. The tailings pond has an impervious bentonite lined dam, and it is divided into a solids and a liquids holding areas. Until the beginning of 1995, the mill tailings were discharged in two separate lines to the solids pond; one line carried the slurry residue of 45-50% solids (w/w) from the CCD circuit, and the other discharged raffinate tails containing less than 1 % (w/w) gypsum precipitate. Supernatant in the solids pond is treated in the primary treatment system prior to discharging to the liquids pond. Figure 1 shows the site plan of Cluff Lake tailings management area.It has been estimated that by the end of 1995 the solids pond would be filled to about 60% of its design capacity, and probably would be full before the year 2000. Therefore, Cluff Mining had to consider a more efficient tailings disposal method that could extend the service life of the existing Cluff Lake tailings management area. A high density tailings thickener was believed to be the most practical and economical system for tailings disposal at Cluff Lake mine."

Jan 1, 1996

By V. Coffin, T. Cienski

"(Figure 1) Reagentized pulp enters into the column at about 1/4 of the distance from the top. Inside the column it meets with a downward flowing stream of water and with an upward flowing mass of air bubbles. Flotation takes place in this counter current system while at the same time fine gangue particles are washed off the froth before it reaches the overflow lip at the top of the column. The main operating features of the column flotation are as follows:a) No moving parts. Solids are kept in suspention not by an agitating action of a propeller but by settling down against the rising air bubbles.b) Wash water. Spray pipes are installed a few inches below the froth surface and deliver required volume of water. The purpose of this water is twofold:(i) it keeps in balance the flow of material through the column and(ii) it washes rising froth in the froth layer as well as removes unwanted fines attached to the bubbles at the lower flotation section. As a result the concentrate produced is of higher purity than in the conventional flotation. Column is characterized by very high up-grading. However, even temporary disruption of wash water supply results in a drastic drop of concentrate grate.c) Flow balance. (Figure 2) For the correct operation volume of the underflow (tail) is larger than that of the feed. The difference is made up by the wash water. This difference called bias is predetermined and ranges from some 4 to 15 USGPM depending on the size of the unit and process requirements."

Jan 1, 1981