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By André R. Laplante

"This presentation reviews three continuous centrifuge separators, the Falcon-C, Kelsey Jig and Multi-Gravity Separator, two units with possible synergies with centrifuges, the In-Line Pressure Jig and InLine Leach Reactor, and two centrifuge prototypes, the continuous Knelson and the Altair Jig. Typical applications, potential problems, promising synergies and general concentration mechanisms are discussed. A general discussion on the sources of information, the nature of the fluidization water flow, test work for potential centrifuge separations and the early life cycle of new technology, using centrifuge separators as an example, follows. ResumeCette présentation fait le point sur trois appareils de séparation centrifuge, le Falcon-C, bac oscillant Kelsey et Multi-Gravity Separator; deux appareils qui présentent des synergies possibles avec les centrifuges, le bac oscillant In-Line Pressure Jig et le système InLine Leach Reactor; et deux prototypes centrifuges, le Knelson continu et le bac oscillant Altair. On y présente des applications typiques, les problèmes d’opération les plus usuels, les synergies les plus intéressantes. On discute également des sources d’information disponibles pour le minérallurgiste, de l’impact du mode d’ addition de l’eau de fluidisation, de la démarche expérimentale a faire avant d’installer un centrifuge, et du cycle de développement de nouveaux équipements, en utilisant les centrifuges comme exemple.IntroductionGravity separation has undergone a renewal in the past 20 years, much of it attributed to centrifuge units. The semi-batch Knelson Concentrator has probably received the most exposure, but it is rather because of the shear number of units sold than the width and breath of its applicability. It excels at recovery very dense minerals in very small concentrations, and has become a workhorse of the gold industry. Falcon’s Super-Bowl comes a distant second, with very similar applications. A number of copycats are available, and are generically identified as “Knelsons.” This presentation will not address these units explicitly, but as possible synergies."

Jan 1, 2000

By A. T. Robles, L. Buckingham

"When certain gold ores containing pyrite, graphite, humates and fumates or other organic compounds which are possibly adsorbed on clay minerals are treated with aqueous solution of alkali metal cyanide, the aurocyanide complex which forms is adsorbed by these ""pregrobbing"" species in the ore. Dissolution of gold proceeds until chemical equilibrium is attained between the cyanide complex adsorbed by the ore and in solution.Sodium lauryl sulfate (SLS) may be used to decrease the adsorptive activity of the prerobbing species in the ore. It has successfully improved recovery at Holt-McDermott Mine and is highly biodegradable.INTRODUCTION Precious metal-bearing ores containing preg-robbing species are poorly amenable to standard cyanidation techniques. The ""preg-robbing"" impurities with which they are associated tend to adsorb the gold cyanide complexes and carry them through the circuit to the tailings. Also, gold may be associated with these impurities in the ore and not be leached by the cyanide complexing agents used in conventional cyanidation techniques.Methods for dealing with the problem of preg-robbinq impurities in gold ores are:Carbon-in-Leach (CIL), in which granular activated carbon is added to compete with preg-robbing species for the gold cyanide complexes. This process can improve gold recovery in mildly preg-robbing ores especially when fine carbons with very high activities are used. However, with severely graphitic ore, the carbon associated with the rock may outperform the granular activated charcoal.Chemical Oxidation, in which the ore is subjected to preliminary chemical oxidation techniques to oxidize carbonaceous materials using such diverse reagents including ozone, sodium hypochlorite, calcium hypochlorite, chlorine gas, permanganates, perchlorates, chlorates and oxygen prior to subjecting the oxidized ore to standard cyanidation treatment to extract gold. This is generally a costly method and will not prove economical on a low-grade ore."

Jan 1, 1992

By Toshiharu Teranishi

When we apply the Au nanoparticle superlattices to the future nanooptical and nanoelectronic devices, the control of not only the particle size but also the interparticle spacing and symmetry of superlattice are the essential problems to be overcome. First, we developed a novel synthetic procedure to precisely control the particle size of alkanethiol-protected Au nanoparticles in the range from 1.5 to 9.7 nm by the heat-treatment of the small Au nanoparticles in the solid state. Absorption and nonlinear optical properties of alkanethiol-protected Au nanoparticle superlattices were then elucidated. On the other hand, the hexagonal close-packed (hcp) 2D superlattices of small Au nanoparticles with the interparticle spacings from 1.9 to 3.6 nm were successfully prepared by using newly-synthesized ligand for single electron tunneling devices. Finally, various symmetric superlattices, such as chain (2-fold), quasi-honeycomb (3-fold), square (4-fold), etc., of Au nanoparticles could be produced by making use of interligand interactions or some templates.

Oct 1, 2003

By Franciszek Letowski

"Recovery of precious metals associated with a sulphidic matrix has an important economic effect on complex sulphide ore treatment. Because a significant portion of the silver and gold associated with pyrite are rejected to tailings during the selective flotation of pyritic Zn-Pb-Cu ore, other ore processing routes have been investigated.Direct leaching of the ore with ferric sulphate solution followed by separation of different components of the solid suspension by flotation was applied. Such a process allows for: production of zinc sulphate solution; transformation of galena PbS to insoluble lead sulphate PbS04 and its accumulation in a non-floatable fraction together with precipitated geothite FeOOH and other insoluble gangue minerals; separation of sulphur and non-decomposed sulphides in the floatable fraction.Laboratory research carried out on the ore samples from an Alaskan pyritic deposit from the Delta region indicates that gold and silver are mostly concentrated in the floatable fraction.Total recovery of zinc from the ore attains 96.5% to 98.5% after ferric sulphate leaching. About 85% of copper was extracted from the ore during the ferric sulphate and subsequent ferric chloride leaching, and about 6% of the copper remained in the non-floatable fraction and mixed fraction during the leaching-flotation step. Extraction of lead from the nonfloatable fraction during leaching with calcium chloride solution attained 98% to 99%. Extraction of silver from the floatable fraction during leaching with ferric chloride solution reached 98%. Extraction of gold from the floatable fraction during cyanide or thiourea leaching reached 71% to 89%. The results of the laboratory research are presented."

Jan 1, 1987

By V. J. Southey

"This paper reviews the history of Wabana Mines from 1892 to 1966, inclusive, in order to clarify the essential factors involved in the achievement of maximum mining efficiency and in the search for an acceptable product to satisfy changing market requirements. It outlines the progressive steps taken to mechanize both surface and underground operations, and forecasts anticipated performance, assuming certain major plans could have been implemented. Results of extensive research in the treatment of the complex ore are discussed. Selected processing alternatives are compared as to feasibility in a summary incorporating estimated capital and operating costs."

Jan 1, 1969

By H. C. M. Gordon

IN every co al mine the hazard of fire in the underground workings is always present. This is a natural condition, because the mineral wrought and the supports normally used are themselves combustible and because underground fires may originate in many ways. The main causes of such fires are: (1) Open lights igniting gas, wood, or the coal itself. (2) Explosives igniting gas or coal. (3) Spontaneous combustion. (4) Electric sparks or short circuits igniting gas, wood, or coal. (5) Frictional heat igniting wood, coal, or oily material. (6) Explosions of gas or coal dust igniting wood or coal. The relative importance of these causes, and others less frequently met with, varies from district to district and depends to a great extent on the equipment and supplies used, on the training and discipline maintained among the workmen, and on the liability of the coal to heat spontaneously. It follows, therefore, that open lights should not be used and that, wherever there is the least possibility of methane being encountered, only locked safety lamps of an approved type should be allowed underground. It follows, also, that only 'permitted' explosives should be used and that the it use should be confined to conditions under which they are ?permitted?. Similarly, electric equipment should be of an approved type and all such stationary equipment should be housed in well-ventilated fireproof rooms. Such rooms, as well as entries to stables and oil houses, should be provided with doors which can be easily and tightly closed and which should be constructed of fireproof material.

Jan 1, 1937

By Stuart McTavish

"Noranda's Goldstream mine is a 1,500 ton per day copper-zinc mine located 60 miles north of Revelstoke, B.C.The orebody which contains 3.69% Cu and 2.69% Zn was discovered in 1972. Construction of the mine and mill was completed in 1982 and start up was achieved in May 1983.This paper will describe the flowsheet and related equipment 1n the crusher and mill, in particular the dewatering section where two Larox pressure filters are used to dry the concentrates.INTRODUCTIONThe Goldstrean property is located 60 miles north of Revelstoke, B.C. in the Goldstream River Valley (Figure #1). It was discovered during the construction of a logging road in 1972. Exploratory drilling was initiated in 1975, and in 1976 a portal was collared to obtain underground data. Samples were taken from the underground workings for metallurgical bench tests and a pilot plant run.Due to economic conditions, the decision to go into production was delayed until 1980. Concentrator construction and underground development progressed simultaneously during 1981 and 1982, and were completed by the end of 1982. Start up of the concentrator was achieved on May 2, 1983.The ore typically contains about 40% sulphides, which are chalcopyrite 10%, sphaletite 6%, pyrrhotite 20%, and pyrite less than 5%. The sulphide minerals have a high degree of intergrowth. Most of the sphalerite occurs in micron sized grains which are embedded in the pyrrhotite and, to a lesser degree in the chalcopyrite. There is also some fine dissemination of the chalcopyrite in the pyrrhotite. The waste rock is made up of about 30% quartz, 20% guange silicates, and 10% carbonates. The ore has a specific gravity of 3.2 and a natural pH of 8.2."

Jan 1, 1985

By C. J. Farrow, M. J. Smith

"Bechtel has managed a number of test programs to establish autogenous grinding circuits and size mills for a variety of ores.This paper outlines the procedures whereby bench scale and pilot plant tests are used to select the most efficient circuit to grind the range of ore types likely in the mill feed.Five examples are given. One example traces the development of a pilot plant program; it shows how bench scale Work Indices provide a target for the grinding power requirement of an ore and relates how pilot circuits are changed as a result of interpreting the results from preceding tests. The other four examples relate the interpretation of the test results to provide design data; they illustrate a number of phenomena which have been observed.INTRODUCTIONConventional crushing and grinding circuits can be specified from bench scale test data such as Bond Work Indices. Bond Indices have been correlated to a large number of rod and ball mills and by applying appropriate factors, mills can be sized from the Work Index.For autogenous or semi-autogenous grinding circuits, bench scale tests are not sufficient to be able to specify the circuit and size of the mills for a commercial plant. It is necessary to run a series of pilot plant tests to determine how the ore behaves."

Jan 1, 1982

By A. Ansari

Standard depressants used in the selective separation of copper and molybdenum sulfides by flotation pose environmental and health hazards. In this investigation, the floatability of chalcopyrite and molybdenite in the presence of six well-characterized lignosulfonates was studied through Hallimond tube flotation and adsorption tests. The results define the physicochemical conditions (lignosulfonate type and dosage, pH, presence of ions) for possible selective separation of molybdenite from chalcopyrite. Dissolved polyvalent cations, especially calcium, appear to serve as "activators" for lignosulfonate adsorption on both minerals. From this point of view, the use of lime as a pH modifier is not always appropriate for optimum selectivity since high dosages of lime to produce high pH tend to indiscriminately depress both sulfides.

Jan 1, 2006

By D. H. Pulling

Statistics for copper consumption from various sources show that the importance of production from secondary sources is increasing. This is of particular significance in countries such as the U.K ., which are totally dependent on secondary material for the supply of home-produced copper . A brief survey of the U.K . industry has shown that, in the past, relatively high profit margins have not encouraged high cap ital investment and development of more sophisticated operating techniques, but set against a background of rapidly escalating operating and energy costs , together with an ever-increasing demand for copper , it is anticipated that the industry will see great changes in the near future. With this in mind, an experimental furnace for continuous smelting of copper-bearing scrap has been developed at the University of Birmingham. This furnace has been designed to combine the conventional operations of blast furnace or reverberatory smelting, usually followed by converting, in one unit, and incorporates many of the features of the WORGRA and Noranda processes for continuous primary smelting . It consists of a long, sloping channel, constructed of chrome-magnesite refractory, which: is churned near the top. Smelting is effected lower down the channel, where iron, etc., in the scrap is oxidized and slagged off with silica flux. Provision is made for oil injection to increase oxidation rates. Metal can flow from the bottom of the channel, and slag from the top, so that a counter-current flow is achieved. The approximate molten capacity is 25 kg. Various scrap charges, both synthetically made and as supplied from various industrial smelters, have been successfully smelted. Metal of 83 -95 % Cu has been produced, depending on the original charge composition and the degree of air injection, although air injection was also found to considerably increase the copper loss to slag. Recoveries of other valuable non-ferrous metals in the charge were good.

Jan 1, 1977

By Y. Q. Li, Z. C. Shen, C. H. Zhao

The dynamic properties of the KYF-320 flotation machine which is one of the biggest flotation machines in the world were tested and then verified by the computational fluid dynamics (CFD) simulation. Aeration rate testing results suggest that he effect of circulation holes on the fluid state in the flotation cell becomes gradually smaller when the impeller speed is greater than 107 r/min. The dispersion degree of air increases with the increasing impeller speed and decreases with the increasing aeration rate, and it increases slightly when the impeller speed is greater than 107 r/min. The air holdup increases and then decreases with the increasing impeller speeds and aeration rates, and a maximum value of air holdup is shown at impeller speed of 107 r/min with a value of 12.28%. CFD simulation results are in good agreement with these testing results. Industrial-scale test was carried out using this 320 m3 flotation machine for the flotation of Cu ores. The results show that Cu content increases from 0.06% of crude ore to 1.26% of concentrate, and the S content increases from 0.39% of crude ore to 24.83% of concentrate. Moreover, using this flotation machine 5.88% of power is saved compared to other type of large flotation machine.

Jan 1, 2015

By J. Ran

"Hangingwall sloughing is the major cause of unplanned stope dilution in open stope mining. Understanding the hangingwall sloughing mechanism can guide the stope design in minimizing the unplanned dilution and improving muck size and therefore mining efficiency. Though the modified stability graph method may be adopted to assess the hangingwall stability, for mines where the hangingwall is inaccessible, the required parameters are difficult to define and some of the simple influential factors cannot be easily identified in this method to provide guidelines for stope design.This paper identifies the main modes of hangingwall sloughing in open stope mining based on detailed data collected from mined-out stopes and site observations, applies the voussoir beam notation to analysis of the hangingwall sloughing mechanism, compares the model analysis results with the collected stope data, and provides guidelines for stope design and model application. Finally, examples are given to demonstrate the time and support effects on hangingwall sloughing and stope dilution."

Jan 1, 2002

By Yves Breau, François Robichaud, Jacques McMullen, Pierre Pelletie

"The Expert System implemented at the East-Malartic mill controls a copper flotation circuit while taking into account the profitability of the operation. Treating a copper gold ore, the flow sheet integrates a flotation step followed by cyanide leach circuit. The Expert System calculates the actual net smelter revenue of the flotation circuit, accounts for the operating cost of the cyanidation circuit and defines optimal metallurgical operating targets to optimize the overall economics. To be realistic, probability factors and circuit diagnostic analysis are integrated in the operating target selection. The Expert System has proved to be successful, robust and flexible. It resulted in a copper concentrate grade increase of 1% while maintaining recovery. The use of the expert system also gives plant personnel good training procedures and promotes interaction with mill operator.INTRODUCTIONThe East-Malartic mill is located close to the town of Malartic in North Western Quebec, Canada. The mill was first commissioned in 1938. The flowsheet at that time was a straight cyanidation circuit. The original mill operated in that mode up to the discovery of the Bousquet 2 orebody in 1989 which consisted of a massive copper sulphide ore averaging 0.7% copper and 7 g/mt of gold. Bousquet 2 ore processing required a mill retrofit. Grinding capacity, gravity and copper flotation circuits were inserted prior to the cyanidation circuit in order to achieve 3,200 mtpd capacity at expected recoveries.In 1993, control of the flotation circuit was deemed critical in order to maintain the flotation performance and to reduce cyanide consumption since the copper sulphide contained in the Bousquet 2 ore is a high cyanide consumer (50% bornite). (McMullen, 1999) As a result, the implementation of advanced control strategies in order to increase the metallurgical profitability of the plant and to reduce the operating variability between operators was the most rational alternative. In 1994, Corem was chosen to perform the plant audit and to write the first set of rules. The on-line control strategy was started up in 1995. In 1996 and 1997 the Expert System was shutdown because of an extreme change in the ore coming from the milling of Bousquet #1 ore and the departure of the projects engineer. In 1998, the system was restarted and the results were impressive. Furthermore in May of 2001 a patent was issued for the control strategy. (Breau, 2001)"

Jan 1, 2002

By B. McLaughlin

Agnico-Eagle started sinking the Penna Shaft in 1995 with a hoisting plant capable of hoisting 1800 t/d from a 2240 m deep shaft. Exploration success led to the installation of a hoisting plant capable of extracting 4500 t of ore and 1800 t of waste per day from a depth of 2210 m. The introduction of a new rope selection factor, permitting a safety factor of 4.0, represented the breakthrough required for a cost-effective project. This paper describes the process taken to introduce the new rope selection factor and its impact on the original plant.

Jan 1, 2004

By Poul Hansen

"In mid-I986, a world-class mining enterprise was created in British Columbia, bringing together the copper mining operations of Lornex Mining Corporation and Cominco into a new entity, Highland Valley Copper. This paper will focus on the background and rationalef or the formation of this partnership, the resulting operating changes and the productivity and cost benefits expected.IntroductionSince the deterioration of the world copper industry in 1981, the future of the Lomex and Cominco mines in the Highland Valley of British Columbia looked increasingly precarious. Despite the expansion of the Lomex mill complex in 1981 to one of the world's most cost-efficient copper and molybdenum concentrators, persistent low metal prices had made the remainder of the ore in the Lomex orebody marginally economic and had significantly foreshortened the expected life of the operation.Cominco had acquired the Bethlehem Copper assets in order to consolidate its ownership of the Valley orebody, and was feeding the Bethlehem concentrator with ore from the Valley orebody. In spite of the high grade and free-milling character of the Valley ore, profitability of the Cominco operation was severely restricted because of the limited capacity of the Bethlehem concentrator and its location high on the hillside above the orebody."

Jan 1, 1987

By Chris Larsen, Ian Orford, Marilene Renaud, Mike Cooper, Lud Strah

"An AG/SAG mill was commissioned at the Brunswick Mine concentrator in October 1998. This paper discusses the original liner design for the bi-directional AG/SAG mill, observed liner wear, subsequent liner design modifications, and benefits. The mill liners consist of steel top hat shell liners, steel end liners, rubber grates and rubber lined pulp lifters.Lifter profile and plate thickness measurements were made on monthly shop days. Trends of wear rates were used to project when liners required replacement. Wear profile information was also used to determine where liner re-design would improve wear life or minimize scrap loss. Charge motion simulation modeling was used as a tool to assist liner design modifications.INTRODUCTIONThe Noranda Inc. Brunswick Mine, located near Bathurst, New Brunswick, is a large underground lead/zinc mine with a nominal production rate of 10,000 tonnes per day. The Brunswick concentrator commenced operation in 1964 with conventional surface crushing followed by rod and ball milling to reduce run of mine ore to the required size to feed two parallel flotation circuits. In 1998 the AG (autogenous grinding) mill was commissioned for a nominal production rate of 9,000 tonnes per day replacing two surface crushing plants, three rod mills and three primary ball mills. In May of 1999 the mining rate was increased and the service of the AG mill was changed to a SAG (semi-autogenous grinding) mill with the addition of 4 inch steel balls (4% volume). The mill operates in closed circuit with 0.66 meter horizontal cyclones to produce a ground product of 90 micron nominal size (K80).The AG/SAG mill liners are steel (Norsteel, chrome-moly steel) with rubber grates and rubber lined pulp lifters.The original liner design was based on studying other AG/SAG mills and recommendations by the liner supplier, Norcast. Design factors taken into account were total cost of lining (including material and labour), protection of mill shell and impact on operating performance."

Jan 1, 2003

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

"This paper presents the principles and application of a recently developed control system for flotation columns and column circuits. The control system uses triple pressure sensors to directly sense zone densities and level, and to help in characterizing the operating behaviour of individual columns. A complete expert system development tool is incorporated within the system to provide the facility for full operator training, an on-line help facility, and a means for incorporating the column behavioural patterns into decision making rules. The system is configurable to any number of columns and can be run in either supervisory mode or direct control through Minnovex PIO software. Additional process signals can be interfaced with the system if desired. Operating experience with the control system on the Minnovex column flotation pilot plant is described.I: INTRODUCTIONColumn flotation cells have been more difficult to control and are more sensitive to circuit changes than conventional mechanical flotation cells. This susceptibility encouraged investigation into various types of control strategies ranging from inexpensive single pressure measurements for level control, to highly instrumented and expensive systems using a variety of control methods.Flotation operators, in many cases, do not understand the column control philosophy or have not been trained to deal with this new type of cell. Some operators have subsequently absolved themselves of the column operating responsibility. Some column cells, for example, have run for hours with slurry pouring over the top or no froth coming off at all.ColumnEX is a column flotation control system which was developed from the viewpoint of an operating shift foreman who wishes to:1. control the column or column circuit through a genuine understanding of what is happening inside the columns, and2. to share this knowledge with his flotation operators and be available to guide them through times of circuit upsets."

Jan 1, 1990

By Z. Xiao, A. R. Laplante

"The recovery of gold by gravity from grinding circuits differs from virtually all other gravity recovery approaches in that stage (or unit) recoveries are low, typically 1 to 5%, but overall recoveries are generally in the 25 to 60% range, because individual gold particles are presented many times to the primary recovery unit. Full-scale recovery is difficult to predict from benchscale tests. Further, maximizing metallurgical recovery is generally not the primary objective, because of the more efficient downstream recovery circuit (flotation or cyanidation). In this paper, a new concept, the gravity recovery effort, is presented to assist in the design and optimization of gold gravity circuits. It is defined as the product of the circulating load treated by the primary gravity unit, the recovery of this unit, and the gold room recovery. Predicted gold recovery is linked to the recovery effort, the size distribution of gravity recoverable gold (GRG), the dimensionless retention time in the mill and the partition curve of GRG. A case study of economic optimization of the gravity recovery effort is presented. INTRODUCTIONPredicting gold gravity recovery from grinding circuits has always been a hit-and-miss exercise. The problem can be traced to the highly variable gravity recoverable gold (GRG) content of ores, high GRG circulating load, fraction of the circulating load treated, and performance of primary gravity recovery and gold room. Typical values, highs and lows are shown in Table 1."

Jan 1, 2001

By B. M. Monaghan

"The developments of the 1950s and 1960s have seen Canada become an international source of iron ore, trading in all geographical areas and hemispheres. Canada's production will significantly increase in the next decade. However, although the outlook is bright, iron ore producers cannot be complacent. Instead, we must weigh the factors that have brought about the development of the Canadian iron ore industry and examine the probable changes for the next decade that will affect its growth.Looking ahead in the 1970s, competition will be severe. Pressure on prices will continue as new sources of supply bid for long-term contracts and as steel producers continue to develop mines around the world. Rising costs will make these pressures more severe. We must balance the growth of the world's steel makers with the emergence of new supplies of iron ore. The iron ore industry is now part of a new era in which demand and supply are intermeshed, and the developments around the world are acting on the Canadian iron ore industry's future."

Jan 1, 1971

By J. Paraszczak

High productivity of the equipment employed is one of the key factors in reducing production cost in mining. One of the best ways to achieve it is an increase in availability. In order to identify the most adverse factors affecting it, reliability and maintainability analysis ought to be done. If its results are to be conclusive, a sufficient amount of credible and consistent data is indispensable. Following the experience from field studies done in underground mines, this paper discusses the current status of failure and maintenance reporting, as well as problems encountered. It exposes the importance of adequate data collection on failures and downtime as well as it reviews the benefits of analysis based on such material. The question of benchmarking with regards to mining equipment performance is also addressed. The paper concludes with some recommendations on data collection and processing, as well as stresses the need to standardize the most important reliability/ maintenance notions throughout the mining industry and to conceive appropriate performance measures.

Jan 1, 2000