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By Paul G. Dunn

Safe rapid drift development results principally in reducing exposure of underground workers to the working face. This reduction in exposure is not achieved until either the rock walls have been suitably scaled or sufficient support has been put in place. Rapid drifting enables quicker access to orebodies (increasing returns), higher production rates and a resultant lower cost of mining. The Mining Innovation, Rehabilitation And Applied Research Corporation (MIRARCO) has been developing technology and a methodology for safe rapid drift development within the underground metalliferous mining industry. In terms of rock scaling for underground metal mines, there is still considerable use of manual and mechanical impact rock scaling techniques. Two main issues result from this approach. The first is that in highly stressed or highly fractured ground, a decision has to be made as to when sufficient scaling has taken place. This is not always easy to do and hence operator judgment is a crucial element in this approach. During a baseline study of current rock scaling practices, a wide range of scaling practices were observed within five different underground mining operations; this demonstrates the difficulty inherent in the rock scaling quality decision making process. The second issue is related to the exposure of workers to freshly created rock surfaces. In manual scaling techniques this is unavoidable and despite the best scaling practices, accidents still occur. In an attempt to provide an alternative scaling technology, MIRARCO has been investigating highpressure water jet scaling for removing loose rock as part of the drift advance process. This technology not only provides a means by which scaling practices can be engineered (remove the loose material you want to in a given rock mass condition), it is also quicker and easily automated. Automating the scaling process removes workers from exposure to freshly created rock surfaces and also enhances the application of spray-on liners through improved adhesion of these support types to the rock surface. This paper describes the water jet scaling technology research and methodology for safe rapid drifting within the underground metaliferous mining industry and highlights the key issues and major drivers behind their continued development at MIRARCO.

May 1, 2004

By L Thompson, R C. West, T S. Kho

In 1978, the Outotec direct to blister flash (DBF) smelting process was commercialized with the commissioning of the Glogow 2 smelting facility in Poland. Since then four further DBF facilities have been constructed, two at Olympic Dam, one in Zambia and most recently one at Glogow 1 in Poland. The DBF smelting process differs considerably from conventional smelting process in both chemistry and unit operations, presenting opportunities for flowsheet capital and operating cost savings particularly when the concentrate copper grade is high. WorleyParsons has amassed considerable expertise in the Direct to Blister process and in this paper the process is compared to the modern two stage Double Flash process. The processes are modeled, and the key operating parameters, consumables and production metrics are directly compared. Subsequent flowsheets, key equipment lists and capital costs are also presented for comparison and discussion. The results show the range of concentrates which are most economically treated by the DBF process relative to those best treated in conventional processes along with those concentrates which could be treated by either process. INTRODUCTION The Outotec Direct to Blister (DBF) process was developed and installed at Glogow in 1978 (Czernecki et al., 2006). Currently there are a total of five DBFs built on three sites: Glogow No.2, Olympic Dam No.1, Olympic Dam No.2, KCM Zambia and Glogow No.1, being the most recently constructed. The DBF process was the first flash furnace process to produce blister copper and paved the way for the subsequent development of the flash converter and other continuous converting processes. Today four flash converters are in operation taking the total number of flash furnaces producing blister copper to eight. In all DBF operations it is the unique ore and concentrate composition which made selection of the DBF process the most attractive. In all cases, the concentrate is relatively high in copper but low in iron because chalcopyrite is not the major constituent. When considering smelting options for high copper, low iron concentrates a key question is then often asked, “At what point does it become economic to switch from two stage smelting and converting process to the DBF process?”

Jan 1, 2019

By A. R. Laplante

"In this paper, the performance of the Knelson Concentrator and of the batch and continuous Falcon Concentrators are compared. The units are based on different operating principles, and as such they perform quite differently. The Knelson yields very high free gold recoveries over a wide size range of both the recovered and rejected species. It is ineffective in recovering very fine ( <15 µm) or flaky gold, as encountered in flotation feeds, concentrates or tails. It is the: preferred unit for gold recovery from grinding circuits. The batch Falcon is effective in the fine range, typically less than 37 µm for gold. Its recovery zone is quickly saturated when there is significant heavies in the feed. It works best as a pre-concentration unit. The continuous Falcon is a versatile unit which is capable of recovering fine heavies (density down to 5 g/cm3 ), especially from coarser gangue. INTRODUCTIONThe purpose of this presentation is to describe the mode of operation of two Canadian centrifugal concentrators, the Knelson and Falcon. It is hoped that it will prove informative to prospective users in the selection, testing and optimization for applications of either machine. The work is based on testwork both in Canadian mills and at McGill. I am indebted to many colleagues at these locations. Appropriate references will be made to articles further describing the work, either already in the open literature or in preparation.Although this work is meant to be more practical than scholarly, it is appropriate to begin by considering how the two units achieve selective recovery. This will become useful in establishing a suitable testing methodology, which will then be described, and interpreting typical results. In the last section I will attempt to define appropriate applications for the two units. This is a thankless task, as I am likely to miss some applications or to make projections based on partial results, as much of the work is still in progress. In deference to the reader and both manufacturers, areas where reasonable doubt still remains will be identified."

Jan 1, 1993

By P. Hanniala, S. Harkki, B. Andersson

"The Outokumpu Flash Smelting Process was put into commercial operation in 1949. The original process was based on the use of preheated air. The adoption of oxygen enrichment of the flash smelting process air in 1971 was a crucial improvement, especially for smelter capacity. Oxygen use according to the Outokumpu method means selection of the most suitable oxygen content for the process air on the basis of the concentrate composition, capacity and matte grade. The furnace is not bound to only one fixed degree of process air oxygen enrichment; the oxygen content may vary from 21% to 100%.The degree of oxygen enrichment and the matte grade are the most important process parameters in reducing the operating costs of the whole system, including smelting, converting and acid production. Raising the matte grade allows more efficient use of the reaction heat in the smelting stage. Sulphur is recovered in a high-strength, constant-flow gas stream. The blowing time in the converters is reduced considerably.Oxygen enrichment and high matte grade improve the energy efficiency of the plant complex as long as the surplus heat from the system is not wasted. When the operating costs and investment expenditure of the total plant complex are taken into account, a reasonably high degree of oxygen enrichment (35-80°/0) is the best alternative in practice. Existing operating smelters have achieved single-stream thoughputs of over one million t/yr of concentrates.HistoryLocal capital and operating costs as well as lower energy consumption per unit of metal produced are now the crucial factors in determining the choice of process. In the late I940s, the same reasons prompted Outokumpu to develop the original flash smelting process in Finland. Between 1900 and 1940, very few technical changes in pyrometallurgical copper production took place. However, the commercial use of the flash smelting technique began a new era in metal extraction from sulphide concentrates. The characteristic feature of the method is that the concentrate is smelted in suspension: this offers the most favourable conditions for heat and mass transfer, and thus high reaction velocities. Outokumpu was the first to exploit the known advantages of flash smelting in a commercial installation at Harjavalta in 1949."

Jan 1, 1982

By J. Châteauneuf, R. Dupéré, D. Mrabet

"In gold processing plants, it is common to start gold cyanidation in the grinding circuit by simple cyanide addition or as a result of recycling residual cyanide. In such cases, advantages such as a decrease in gold lockup and reduced leach time have been noticed. The prevailing conditions in grinding circuits such as intense mixing, particle breakage and friction or high temperature are quite different from those found in tanks and so is the resulting cyanidation behaviour. To gain a better understanding of the cyanidation behaviour in mills and of the pertinence of cyanide addition in grinding circuits, lab tests were conducted using an industrial ore from Canadian Malartic mine. In this study, the effects of different parameters on cyanidation in grinding circuits were examined; parameters such as such as temperature, dissolved oxygen, cyanide concentration and pH. The results showed that the kinetic benefits of high temperature in the mill can overcome the low dissolved oxygen concentration thus leading to an overall gold dissolution kinetics improvement for temperatures up to 50°C. However, gold leach in the studied grinding circuit at lab scale caused an increase in cyanide consumption with a higher generation of iron cyanide and SCN-. These results were confirmed by tests in the Canadian Malartic plant where the cyanide addition was moved from the SAG mill to the leaching tanks resulting in a 20% savings in cyanide consumption.IntroductionCyanidation in grinding circuits can benefit gold processing by decreasing gold lockup and by reducing leach time. However, other aspects regarding gold and cyanicide dissolution come into play in the conditions prevailing in grinding circuits. Indeed, high turbulence, high temperature, recycling load, friction and particle breakage make cyanidation behaviour in grinding circuits different from in-tank cyanidation (Marsden and House, 2006). In several cases, cyanide in grinding is limited to a very low level or completely avoided such as in circuits with gravimetric separation, for ores requiring oxidative pretreatment or for preg-robbing ores. In other cases, water management imposes a certain recycling of residual cyanide. This problem is generally addressed by a case-by-case approach. However, the chemistry of cyanidation in mill circuits is worth investigating in order to assess the advantages and disadvantages."

Jan 1, 2016

Adaptive Processing as proposed in geometallurgy has to rely on spatially interpolated information on geometallurgical parameters like phase composition, size distributions of particles of different phases, grain shape parameters, and portions of value elements in different grains. Using the geostatistically predicted values for adaptive processing, e.g. for the selection of milling diameters, thresholds in physical separation, or choices on using an extra pre-separation step, is typically not optimal. Mathematically this effect is introduced by two forms of nonlinearities: 1) The nonlinear scales of compositions, distributions, and shapes have special properties with respect to geostatistics. Classical geostatistics creates some artefacts for these nonlinear scales. On the other hand, modern geostatistical procedures adapted to these scales do not provide unbiased results with respect to linear transformations of the data (e.g. biased block estimates). 2) Neither economic nor ecological effects (e.g. monetary gain) of processing decisions are linear in the interpolated geometallurgical parameters. These nonlinear transforms are not unbiasedly estimated by the likewise transformed unbiased geostatistical predictions of the geometallurgical parameters. Furthermore, we need to optimize the conditional expectation of the gain, rather than obtain an unbiased estimate. Standard geostatistics as such does not provide the "right sort" of estimates for adaptive processing. A nonlinear kriging procedure is needed to approximate the nonlinearities mentioned before. We propose to solve these problems simultaneously using a nonlinear geostatistical technique for predicting the target function (the monetary gain), rather than to predict the geometallurgical parameters and compute the gain from them. The optimization can then be performed directly on this estimated function. It can be shown that this optimization performed on the conditional expectations, not on unbiased predictions, would yield the best possible processing choice. We propose a procedure choosing the processing parameters on an approximation of the conditional expectation. The difficulties with the classical approach and the effectiveness of this new approach are illustrated by a simplified simulation example with a single processing parameter and a simple dependence on the microstructure.

Aug 1, 2013

By Lynne Bowen

"On the da y shift in Number One Mine at Nanaimo on Vancouver Island, on May 3, 1887, William Griffiths, a coal miner with many years experience, fired a shot to advance the face of the crosscut. A flame ran out for a distance of fifteen feet along the roof and then extended up the slope for ten to twelve feet, and although many would have called this a blowing out shot, Griffiths said it was just a feeder. ""A blower"", he said, ""would have caused an explosion."" 1There was no explosion. Griffiths reported gas and the area was cleared for three hours and then work resumed. When the shift changed at 3:00 p.m., the fireman reported all places clear of gas and the ventilation good.2The ventilation in Number One Mine was the best anyone had ever seen. The air in the mine was almost too cold. The current was so strong that it sometimes blew out the flame on the miners&apos; fish oil lamps. It delivered five to six hundred cubic feet of air for every man and animal working in the mine. The law required only one hundred cubic feet 3The miners considered Number One to be relatively safe. The manager was proud of the mine and its up-to-date equipment"". Recently, there had been reports from France and America about the explosiveness of coal dust, but the manager did not think his mine was sufficiently dusty to be dangerous.5 The dust lay thick on the floor of the mine. Sometimes when a shot was fired the dust rose in great clouds. When a miner was shovelling it, the dust would billow up and be carried off by the current of air. The dust was also in the roof and in the stopings, put there intentionally to fill up the spaces and prevent the accumulation of methane gas."

Jan 1, 1992

By Michel Ruel

Most plants do not track how their systems perform. Optimization is the easiest way to benefit from your control systems. We suggest an approach in four phases. The first phase is to assess all equipment to determine the performances for instruments, valves, variable speed drives, and other equipment. The second phase is to remediate instrumentation problems, including the ones pertaining to valves, variable speed drives, analyzers, transmitters. When equipment is in order, loops are put back in automatic mode. If a loop oscillates, use standard default values for this type of loop. The third phase consists in validating the control strategies, properly configuring the systems (control system, programs and parameters, HMI, alarms), and then tuning the loops and control strategies. The fourth phase consists in analyzing the process and expected performances in order to optimize the control loops and control strategies. If performance is not sufficient, a decision tree will be presented to determine the right approach: advanced regulatory control, model predictive control, or fuzzy logic control. Once all steps have been completed, control performance monitoring software is used to sustain the results and pinpoint equipment or loops not performing as stated. Alarm management and rationalization could also be performed after auditing the alarm system, if needed. Examples from plants will be used to demonstrate how this step-by-step method is applied and produces results. INTRODUCTION Maximizing what you already have is the best investment to remain competitive. This article aims to show the benefits of a systematic approach using modern tools to improve plant performance. We at BBA suggest dividing the work in four phases, and a subsequent phase can begin when the previous one has progressed: Assess, Remediate, Stabilize, Optimize, or add advanced control. This approach has been successfully applied to several sites. Examples and illustrations from these plants demonstrate how this step-by-step method is applied and produces results. Under the BBA method, there are four questions to answer before proceeding to the next step. For each control strategy, you should ask the following questions: Assessment: Does it work? Remediation: Do you use it? Stabilization: Is it stable? Optimization: Does it perform? Do you need advanced control?

Jan 1, 2019

By Poul Hansen

"In mid-1986, a world-class mining enterprise was created in Highland Valley, British Columbia, bringing together the copper mining operations of Lomex and Cominco . In 1988, Highmont joined the new partnership.Highland Valley Copper is the largest copper mine in North America in terms of both tonnage mined and tonnage milled, and it is one of the largest in the world. Due to its relatively low ore grade its production of copper in concentrate ranks about ninth in the Western World.The current eighteen year mine plan is based on ore reserves of 761 million tonnes. The daily milling rate is 133 000 tonnes. At the end of mine life, a total area of approximately 6500 hectares will have been disturbed. At this time, vegetation has been established on about 600 hectares and while work will be done during the operating years, it will not be until the end of mine life, about 2008, that the last half of the land can be reclaimed and revegetated.Once final reclamation has been accomplished, Highland Valley Copper will have completed its "" temporary use of the land"". In the meantime, about 1250people have had employment for about 50 000 man years, at wages ranking among the best in Canada.IntroductionTo provide and sustain our way of life, we need steel, copper, aluminum, zinc and many other minerals.To sustain mining, we must have adequate ore reserves and replacement of reserves as they are mined. Each year the search becomes more difficult as more of the Earth&apos;s surface is investigated and depo sits are mined. This calls for aggressive and sustained efforts.In British Columbia, past and present mining activities have disturbed less than one-tenth of one per cent of the land.With today’s technology, mining operations can be substantially reclaimed and thus complete the cycle of ""A Temporary Use of the Land""."

Jan 1, 1992

By A. L. Mular

"On behalf of Bob Bruce and myself I would like to begin by stressing some key points.First, most of us know that the Canadian Mineral Processors was formed in 1969 and is an offshoot of the Canadian Gold Metallurgists. I attended that first meeting in 1969, and I haven't missed a meeting since! Pretty obvious that CMP serves the best booze! At any rate, we know that our meetings are supported by CANMET. Camsell Hall is always at our disposal; Bob Bruce acts as permanent secretary pretty much free of charge; we receive a financial grant from CANMET. This brings me to a second point.Second the Canadian Mineral Processors do not enjoy a free ride in this relationship. CMP provides a genuine service to the mining industry, to government, and to universities. For the industry, CMP serves as a forum for the rapid dissemination of new ideas and new technology developed at remote locations throughout Canada. New developments outside Canada are introduced by invited speakers at our meetings. The effect is to help reduce operating costs and increase efficiency.For government, problem areas and research priorities can be isolated through mutual discussions. Direct communication at our meetings can help CANMET to temper its goals and modify its national priorities.For universities, CMP provides a wealth of operating experience and practical know how. The yearly proceedings are a major current reference source fo1· those of us who teach mineral processing. Admittedly, the proceedings are publ-shed from CMP coffers. However, I note that our fees are up this year, which brings me to a third point.Third, we face some potential problems of a financial and political nature. From a financial view, we may lose our CANMET grant and other financial good will simply because of spending constraints. Our fees are expected to climb each year thereafter.From a political view, we know that the metallurgical society of CIM is anxious to incorporate CMP as a section. The pros and cons of such a merger have been described in a recent mailing. In th2 final analysis, the political and financial problems are interrelated. There are several alternative solutions to both problems, which brings me to a fourth point"

Jan 1, 1979

By R. Al-Hent, M. Aissa, J. Asfahani

"This paper focuses on determining the favorable geological and hydrogeological conditions for forming uranium occurrences in the Froqlos-Jabal Abou Rabah region, Palmyrides, Syria. Eighty rock samples from this area were analyzed by spectrometric gamma-ray technique to determine the radioactive concentrations of eU and eTh, and of %K. Uranium concentrations in shallow and deep groundwaters were also determined. High uranium concentrations were registered owing to the presence of phosphatic outcrops in the study region. The uranium migration trends and its remobilization were analyzed through the analysis of the behavior of eU, eTh, and their ratio. A positive relationship between eU and eU/eTh, and a negative relationship between eTh and eU/eTh in geological formations of different ages supports the secondary remobilization of uranium. The role of groundwater movement in transportation and redeposition of uranium mineralization is discussed: deep waters are less oxidizing than near-surface waters, and do not appear to have significantly remobilized uranium. This indicates that the source of secondary uranium mineralization is surficial or near surficial, and is mainly related to the leaching of outcropping phosphatic layers.Integration of the results with established radioactive and geological sections reveals four radioactive anomalous zones: north Froqlos, Kherbet Hannora, northwestern flanks of Jabal Abou Rabah (Kherbet Al-Hajar), and the northwestern limit of the Ghuntor depression. The high uranium concentration in north Froqlos is because of its original presence in the phosphate beds and extensive fracturing, which allows groundwaters to penetrate and remobilize the uranium. High uranium concentrations in the other three locations are caused by the presence of evaporites and capillary action, which draws solutions upwards and causes redeposition as surface crusts (gypcrete)."

Jan 1, 2009

By M. Amsden

"Since the computer program is far from complete this must be only a brief progress report on the computerization of flotation metallurgy at the Ecstall concentrator at Timmins.Ecstall is currently treating about 10,000 tons per day of two different ore types. One is a low sulphide copper-zinc ore, the other is a high sulphide lead-zinc-silver ore. The concentrator is divided into three individual grinding and flotation circuits. In effect there are three mills under one roof. Normally there are two circuits on copper-zinc ore and oneone lead-zinc-silver ore. At this point in time one of the copper-zinc circuits is under computerized metallurgy control.The first step in this control is the use of an ARL on-stream x-ray analyzer, coupled with a Honeywell 21 computer which types out assays and recoveries for each individual circuit. Our approach to a computer program was to keep it simple. The program was written to coincide as closely as possible to what an operator would do in the performance of his duties based on the in-formation he was getting from the x-ray analyser read-out. Initially, computer control was put on just the copper circuit of the ""A"" Division, controlling the collector addition at three points in the circuit; primary, secondary, and scavengers. The control was based on recoveries and the grade of the copper concentrate. Targets like 95% recovery and 24% copper concentrate grades that it was felt the circuit could achieve were set to start with. It quickly became apparent that this alone was not enough because if on a certain ore the best recovery that could be made was only 94.9% the computer kept calling for more reagents until the circuit was completely out of control. On another ore it might have been quite possible to get 96% recovery but when the set point of 95% was reached the computer stopped trying for any more. To overcome this problem a very simple optimum seeking program was devised by one of the Ecstall staff and it has worked very satisfactorily since being put into use in July 1970.In the copper float only the two collectors R-208 and the xanthate R-317 are under computer control. The addition of frother and depressant is left to the operator. The flotation cells have automatic level controls which can be adjusted by the operator. pH is controlled automatically through a lime loop-sensor arrangement."

Jan 1, 1971

By Vic Casaletto

"IntroductionThe Canadian mining industry is under immense pressure to compete with foreign mine operations. The very survival of the Canadian mining industry is threatened by lower ore grades, environmental pressures, high cost of labour, unfriendly tax laws toward exploration and the increasing cost of new technology. In order to compete in the world market, Canadian producers and their suppliers are rethinking age old customer-supplier relationships in favour of preferred supplier relationships which can improve the profitability and enhance the competitive position of both partners.This paper attempts to redefine partnerships with suppliers in the context of a ""win-win"" objective. It will outline the measurable benefits to both partners through a representative product which highlights the keynote structure of any successful supplier partnership The Status Quo...Mines all over North America are plagued with the same recurring questions, ""Who has the best product for my operation at the lowest price... How do we know a product will perform to a supplier&apos;s projections after the sale...What do we do if the product fails to perform as expected?"" These questions have been asked by underground mining operators for decades but rarely have long-term solutions with measurable results been achieved.The LHD industry is one example of the status quo where partnerships will provide large gains toward technological advancement, reducing mobile equipment costs. The underground mining industry is out of sync with their surface counterparts in the open pit business.Partnerships with suppliers in the open pit world are not only a standard practice but a culture which has existed for decades. Open pit operators enjoy concepts which fix their mobile equipment costs for the life of the mine and hold the equipment supplier accountable for performance and fleet economics for 60 000 service hours and longer.Manufacturers are attempting to transfer this culture into the underground mining industry. Figure I illustrates the significance of a capital driven decision."

Jan 1, 1994

By T. M. Siew

"This paper begins with a brief discussion of the geology of the Newfoundland Zinc Mine area. The results of groundwater studies and pumping tests are then presented. It was decided that groundwater in the mine could best be controlled by conventional sump pumping and grouting. The system is fully described, as well as plans for improvements on the method. LocationThe Newfoundland Zinc Mine is located in the Great Northern Peninsula of Newfoundland, six miles northeast of Daniel&apos;s Harbour. The mine is situated at an elevation of approximately 350 feet above sea level, 5 miles farther inland , west of the long range of mountains which rise abruptly with elevations of over 2,000 feet above sea level.GeologyThe ore deposits occur in lower Ordovician carbonate rocks that form part of the stable St. Lawrence platform, which is mainly composed of (1) the lower Cambrian Labrador group, (2) the upper Cambrian to lower Ordovician St. George group, (3) the disconformity overlying Middle Ordovician limestones and shales of the Table Head formation and, (4) two overlying klippen composed of Cambrian and lower Ordovician rocks.The ore deposit is a Mississippi Valley type of the Tennessee variety, having proven are reserves of 5.3 million tons averaging 9.70/0 zinc from eleven ore zones which exist in an area of 5.5 square miles.Among these ore zones, the L zone is the major underground mining area, and is by far the largest of the known deposits. It has proven are reserves of 3.4 million tons with 11.5% zinc; this underground development has been inprogress since 1974.Stratigraphically, the orebody, which occurs about 300 feet below the disconformities between the Table Head and St. George formations, is situated in the upper St. George formationof lower Ordovician age.The upper St. George formation of the L zone is divided into three main units (Fig . I) : the underlying lower limestone the middle grey dolomite pseudo breccia unit and the overlying siliceous dolomite unit. In this area of the mine, the beds of the St. George formation dip gently west to northwest to 10 to 15 degrees. The orebody, which is generally strata controlled with a southwesterly plunge and which has the form of thin, narrow, but elongate, en-echelon ore bed s, is contained in grey dolomite pseudo breccia. Its dimensions vary from 50 to 900 feet in width and 8 to 120 feet in thickness, and the over-alllength exceed s 6,000 feet.The surficial deposit of the area is a till comprising a heterogeneous mixture of sand, gravel, cobbles and boulders, with an average depth of 20 feet."

Jan 1, 1980

By Jan E. Nesset, L. Paul Staples

"High Gradient Magnetic Separation (HGMS) technology, although developed over 2 decades ago, has not been widely applied to the separation of sulphide minerals. Under conditions of strong magnetic field gradients, weakly magnetic minerals such as chalcopyrite (CuFeS2) and marmatite (Zn(Fe)S, sphalerite with iron in solid solution) behave paramagnetically and can be effectively separated from diamagnetic minerals such as galena (PbS). The application of this technology to the separation of Cu-Pb and Zn-Pb bearing streams in the treatment of fine grained complex sulphides may in some cases be more effective than separation by flotation. A 2 TPH Sala HGMS pilot plant was operated at Brunswick for a nine month period, ending in March 1995. The testing demonstrated the mechanical reliability of the technology, the consistency of operating results and the scale up from laboratory testing. HGMS is seen as a complimentary technology to flotation, best suited to performing those separations of galena (and other diamagnetics) which flotation does not do as well.IntroductionBackgroundSince the turn of the century, the magnetic character or magnetic susceptibilities of minerals have been determined with increasing precision and accuracy. Figure 1 illustrates specific magnetic susceptibilities of some common minerals. As long as susceptibility ranges do not overlap, separation is possible.In 1937, the principle of using a magnetizable matrix material in a magnetic field to increase both magnetic flux density gradients and number of capture sites for susceptible minerals was first applied (Frantz). It was not until the 1970's, however, that the genuine development of wet high gradient magnetic separation began with the introduction of very fine matrix materials to the magnetized volume of large solenoid magnets. Today, dozens of HGMS units purify, filter and separate minerals, primarily for the removal of iron stained minerals from clays used in the paper coating and china clay industries and for the recovery of fine titanium and iron oxides in the iron ore industry. Only the Garpenburg mill in Sweden is currently applying a continuous HGMS machine to separate complex sulphide minerals; upgrading a lead concentrate since 1988."

Jan 1, 1996

By J. Ji, A. Shi, B. Grcic

Floating copper minerals selectively against iron minerals such as pyrite has been challenging for most copper and copper/gold deposits, and many operations continuously strive to maximise copper recovery whilst meeting target concentrate grades. Maximising grade is achieved by selectively floating copper minerals and rejecting pyrite. This challenge is to minimise copper and gold losses in the roughers if this approach is used. A flotation testing program was conducted by Bureau Veritas Minerals on a high-level pyrite bearing copper/gold ore sample to establish the potential to reject pyrite in the rougher stage, but also maximise copper and gold recoveries, and ultimately final product grades. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. The main variables tested for achieving selective flotation of copper and gold were collector type and pH control, through lime addition. Grind sizes used in the rougher and cleaner tests had been established from previous test work. The collectors tested were SIPX, Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The two best performing collectors were then tested in open-circuit cleaner tests, with the optimum conditions examined further in locked cycle tests. INTRODUCTION Bureau Veritas Minerals Metallurgical Division was retained by Eldorado Gold Corporation to perform a metallurgical testing program on a high-level pyrite bearing copper/gold ore sample from a gold- copper porphyry deposit to optimize flotation reagent scheme to improve gold and copper recoveries and copper content in the concentrate. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. Initially, various collectors were screened in kinetic rougher tests, and over a pH range of 8.5 (i.e the natural ore pH) up to 11.5, using a primary grind size of 80% passing 120 µm. The collectors tested were SIPX (sodium isopropyl xanthate), Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The enrichment ratio, which is the ratio of copper to sulfur in the concentrate to that in the feed, was used as proxy for rating selectivity (i.e. degree of pyrite rejection or depression).

Jan 1, 2019

By G. A. Kosick, G. S. Dobby, O. Matwijenko

"A 100 kg/hr portable flotation pilot plant has been operated at Kidd Creek Mines with the objectives of:(a) determining performance expected by column flotation circuits on primary zinc and copper cleaning, and (b) evaluating column flotation circuit configurations.The self-supporting pilot plant, owned by Minnovex Technologies Inc. and operated by Minnovex and the University of Toronto, consisted of three 16 cm diameter by 6.5 m high columns and a bank of four Denver No. 7 cells. It was configurable to any arrangement using these four (or fewer) units. The columns were controlled through the use of multiple pressure transducer measurements. This paper presents the approach to testing and the results of more than 70 circuit tests. INTRODUCTIONThe integration and retrofitting of flotation columns within existing mechanical cell flotation circuits is becoming a common occurrence. The correct approach to this integration, however, is often difficult to define, due to a scarcity of previous experience and the difficulty in predicting column circuit performance. In most cases, the effect of column implementation has been experimentally assessed using a single pilot column, with the consequence that selection of an appropriate circuit to implement is difficult. Important considerations are the impact that column water addition will have on the existing operations, and the differences between mechanical cell and column performance in integrated cleaning circuits.To address the problem of circuit selection a portable flotation pilot plant, consisting of both columns and mechanical cells, was designed and constructed by Minnovex Technologies. The pilot plant was structured for complete flexibility of circuit configuration, with the intent to provide a means by which both research and plant personnel could determine the best possible column circuit configuration for a given application."

Jan 1, 1990

By V. S. Melamud, A. G. Bulaev

"Sulfide ore processing is characterized by significant losses of metals, for example, with flotation tailings. Therefore, flotation tailings can present an important source of metals. The treatment of flotation tailings is a technological issue due to low content of valuable metals (e.g. copper and zinc) and high iron content. The goal of the present work was to investigate extraction of base metals and gold from old flotation tailings containing 17% iron, 0.26% copper, 0.22% zinc, and 0.7 g/t gold. Chemical leaching in columns with 0.5 to 10% sulfuric acid solutions and bioleaching in air-lift percolators were used for the extraction of base metals. 1% H2SO4 provided high extraction of base metals and best selectivity. Recovery of copper and zinc reached 35% and 40%. Increasing H2SO4 concentration did not lead to significant increase in base metals extraction, but led to increased iron concentration in the pregnant solutions. The 1% H2SO4 leaching solid residues were used for further experiments. Acid leaching and bioleaching did not result in significant increase in base metals recovery. Leaching of the residue with 10% H2SO4 provided extraction of 4% copper and 20% zinc, and bioleaching for 100 days provided extraction of 16% copper and 12% zinc. The recovery of gold by cyanidation from the flotation tailings, the 1% H2SO4 leaching solid residue and residue of two-step acid leaching were 50, 49, and 65%, respectively. The results suggest that acid leaching can be effective approach for treatment of old flotation tailings.INTRODUCTIONAbout 80% of base metals (copper, nickel, zinc) are produced through pyrometallurgical treatment of sulfide ores. Mining and metallurgical treatments of sulfide ores are presently characterized by significant losses of non-ferrous and precious metals into the various wastes. For example, flotation of sulfidic ores yields formation sulfide flotation tailings. These wastes pose a problem not just because of their sheer volume, but some of them may affect local ecosystems. Nowadays, mining and metallurgical wastes can be considered as technogenical source of base metals."

Jan 1, 2016

By A. J. Williams

"Whenever a decision is made to bring a mineral deposit into production, the Owner in actual fact has decided to convert substantial sums of capital into productive capacity with the intenti6n of making a profit from such conversion.Escalating costs of construction, frustrating and costly delays in comp1etion, the increasing tempo of technological change and the increasing size and complexity of mineral processing projects have convinced Owners that the best way of implementing their capital must be sought and utilized. This requirement has stimulated a search for ways and means of applying modern management techniques to the design and construction of mineral processing plants.This search by consulting firms has resulted in the evolution of the concept of project management and the development of planning and con trolling techniques with the result that Owners can seek and obtain better assurance that their project objectives will be fulfilled.This paper will attempt to outline how a mineral processing plant can be brought into production through the implementation of project management techniques.Definition And Objectives Of Project ManagementThe design and construction of mineral processing plants is subject to the restraints of three inter-related parameters -Quality, Cost and Time each of which influences and is influenced by the other two.Project management is a technique where the Owner delegates complete responsibility to a single organization for all phases of a project (design, procurement, construction and commissioning) for the above parameters but retains freedom for major decision making. Normally, the consultant under-takes full responsibility for all phases, but partial project management services can be provided when the Owner so desires.The objective of project management is first to devise and define and then to implement a program in which the relationship between the three parameters of Quality, Cost and Time is optimized, i.e., the desired functional and other requirements are attained (but not exceeded) at the lowest practicable overall cost in the shortest practicable time. The measure of success of a project lies in the degree to which this objective is achieved."

Jan 1, 1974

By M. Jackman

"Froth stability is critically important for an effective separation of fine and ultrafine particles using column flotation. While a desirable froth stability is required to facilitate the drainage of entrained gangue particles for secondary cleaning, excessively stable froths could create problems in separation efficiency and down-stream processing. In this work, we demonstrate that a stable foam was produced using a poly (N-isopropylacrylamide) (PNIPAM) solution at the temperatures below the critical solution temperature. In contrast, very little foam was produced at the temperature above the critical solution temperature (~32 oC) of PNIPAM. We can also tune the foam stability by changing the temperature of wash water. When warm water of temperature of 50 oC was applied, the foam collapsed and diminished. This is probably attributed to the switching of PNIPAM molecules from coiled to a globule conformation as the temperature is increased. Our results indicate that the foamability and foam stability can be tuned by tempartureresponsive PNIPAM. This is the first time that thermo-responsive PNIPAM has been shown to have potential as a modifier of froth stability in column flotation.INTRODUCTIONIn mineral processing, column flotation has become a more prevalent process in the separation of fine mineral particles (typically < 100µm). Traditional flotation cells are sometimes less efficient in selective separation of ultra-fine particles (Kohmuench et al., 2009). The efficiency of flotation columns is dependent on the froth stability at the top of the cell to drain the unwanted gangue particles entrained in the froth. The greater froth height in flotation columns, compared to mechanical flotation cells, gives a longer froth residence time, allowing more of the drainage process to take place before the froth reports to the launderer. Secondary cleaning, by applying a wash water shower to the froth, enhances drainage of gangue particles entrained in the froth, and improves the recovery of the targeted minerals (Finch et al., 2007).While a stable froth is desirable to facilitate drainage of entrained gangue particles, greater amounts of fine particles have been found to contribute to excessively stable froths which could create problems for down-stream processes (Bethell et al., 2012). Operators have tried spraying additional wash water on the froth, and reducing frother dosage to destabilize the froth and to prevent backup at the launderer. The former is sometimes inefficient, and the latter results in unacceptable recovery loss. Currently, the best practice is often the installation of larger froth launderers, adding expensive defoaming agents, and/or using deaeration tanks (Bethell & Luttrell, 2005)."

Jan 1, 2016