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"The ore currently being milled at Delnite is a high pyrrhotite ore with the gold intimately associated with the pyrrhotite. To get maximum recovery of the gold with the lowest reagent cost, it is necessary to keep the pH as low as possible In the laboratory the best results were obtained with a pH of 10 but in the plant it was found that a pH of 11 was the minimum practical pH. With the n1anual control it was difficult to maintain this figure.The Mines Branch had a conductivity probe (see Fig. 1), developed by Dr. G .G. Eichholz that they wished to try out in plant operation.On June 17, 1963, the probe with all necessary control equipment was installed in the regrind circuit (see Fig. 2) at Delnite Mines. Crushed quick lime is fed at this point but it is also necessary to feed hydrated li1ne at two other points in the cyanide circuit in order to maintain the pH close to 11. 0 throughout the circuit. Only the feeding of the quick lime at the regrind mill was controlled in this test.One thing that concerned me with the operation of the probe was whether the reading on the probe chart would correspond to the pH. A look at the graph accompanying this paper will show that except for very minor variations the graphs of the chart reading and the pH reading are parallel. The upper graph shows that the pH reading and the lime titrations also correspond and that the cyanide titrations appear to have little effect on the pH reading.Table 1 gives the results obtained for the two months before and after the instal1ation of automatic control as well as the two months of control. Unfortunately there was a decided change in ore just after the automatic controls were removed. There was much less pyrrhotitc in the ore than usual thus changing the need for lime and cyanide."

Jan 1, 1964

By D. Shishin

Continuous integrated experimental and thermodynamic modelling work is conducted in Pyrometallurgy Innovation Center (PYROSEARCH). The target is to develop a tool, which is capable to predict the compositions of phases, phase equilibria and enthalpy balance for pyrometallugical processes with higher accuracy. All these can be achieved by thermodynamic calculations based on Gibbs energy minimization algorithm. Recent developments made it possible to predict important factors, such as the chemical solubility of Cu in slag as a function of matte grade, temperature, P(SO2), Fe/SiO2 ratio in slag, wt% Al2O3, CaO and MgO in slag. In industrial practice, it is not only chemical solubility, but the overall amount of slag, which affects the recovery of copper and other valuable metals. It will also impact the energy balance. The presence of solids in the slag matrix influences the physical entrainment of matte in slag. Oxygen enrichment which is sometimes used to control the temperature of the furnace, may also affect the volatilization of minor elements into the gas phase. Same is true if additional fuel is used to compensate for heat balance. In other words, all key process parameters mentioned above are interrelated. Present study demonstrates the ways thermodynamic modelling can be used to link those process parameters on the example of copper smelting. INTRODUCTION Research methodology involving integrated thermodynamic modelling and experimental approach has been recently reported by the authors (Shishin, Hayes, & Jak, 2018 Shishin, Hidayat, Jak, Decterov, & Belov, 2016). Present paper reviews the updates and discusses challenges in the development of a large self- consistent thermodynamic database. The database covers the Al-Ca-Cu-Fe-Mg-O-Pb-S-Si-Zn-(Ag, As, Au, Bi, Ni, Sb, Sn) chemical system. The database works in the environment of FactSage software Bale et al., 2016 and related family of products: ChemApp (Petersen & Hack, 2007), ChemSheet and SimuSage. A list of phases, and thermodynamic models is shown in Table 1. For the copper industry, most important phases are Gas, Slag, Spinel and Liquid matte/metal/speiss. Black-colored elements represent the base system, elements in green can be considered as slag modifiers (slagging elements), and elements in red are minor. Ni is highlighted because it was included recently, in 2019, and currently under development. In addition to copper, the database covers the conditions typical for lead and zinc pyrometallurgical extraction. It is developed in consistency with the experimental results for high concentrations of Pb and Zn. Examples of applications in the present study include the thermodynamic analysis of fluxing for an agitated copper reactor to show the effect of Fe/SiO2, and calculations of copper slag cleaning using reductant.

Jan 1, 2019

By L. A. Nazarova, M. Vandamme, L. A. Nazarov, R. I. Rodin

"The paper describes a new method of interpreting the “canister test” data for the quantitative estimation of gas-kinetic parameters of coal bed based on the developed model of gas emission from coal sample placed in a sealed container. The model accounts for grain-size composition of the sample, initial gas content S, as well as coefficients of diffusion D and mass transfer ?. The inverse coefficient problem is formulated to find S, D and ??from the data of pressure measurements in the sealed container. It is shown that the inverse problem is uniquely solvable if the pressure is dropped during the experiment. The method has been tested using the data from an underground coal mine in the Kuznetsk Basin.Based on the poroelastic model of a coal bed and Khristianovich’s concept about origination and evolution of the filtration zone F in the neighborhood of a degassing borehole the authors considered the corresponding direct problem on rock mass geomechanical state. It was offered interpretation method for data on pressure measurements in a plugged degassing borehole allows evaluation of the diffusion coefficient, the horizontal component of natural stress field and parameters of F. The related inverse problem of the mixed type is formulated, and the cost function is constructed. The analysis using synthetic input data has shown that the cost function has a number of local minimums arranged on the same straight line, one of the minimums being the inverse problem solution. INTRODUCTIONPre-mine degassing is an integral part of coal mining. It allows mitigation of induced accidents in mines (coal and gas outbursts, rock bursts) (Guide, 1990; Malyshev & Airuni, 1999; Seidle, 2011) and also furnishes an alternative and ecology-friendly energy source—methane (Rice, 1997; Nambo, 2001). Choosing an optimized pattern of degassing boreholes, as well as estimating degassing time and amount of gas recovered requires data on gas-kinetics indices of coal beds out of which the key parameter is a gas content S. According to the latest guidelines for coal mines in Russia (Ruban et al, 2012), mining is allowed if S in a coal bed section is under 7 kg/m3."

Jan 1, 2015

By J. Cueto, M. Torres, J. Fernandez, J. Nuñez, C. Salgado, G. Zarate

Anglo American has used the extractant M 5640 in all the Divisions, which is an aldoxime modified with TXIB formulation, except in Los Bronces, where an aldoxime modified with tridecanol (TDA) formulation is used. It is a well-known fact that the aldoxime degrades while the modifier does not. With time the aldoxime/modifier ratio decreases significantly depending on the operating conditions of each plant. In order to evaluate the effect of this increasing modifier content in the organic of the plant, especially in the organic losses, a complete metallurgical test work program was started at laboratory and pilot plant scale. Results showed that the addition of aldoxime without modifier does not affect the chemical performance of the organic phase, but there is a significant decrease in the viscosity and density of the organic phase. Based on the results obtained in the metallurgical program, it was decided to run plant scale tests at both SX plants, Mantos Blancos and Mantoverde. In these tests, instead of adding the usual formulation of aldoxime modified with TXIB to maintain the organic inventory, CuPRO MEX™ 3406 which is an aldoxime without modifier is going to be added until the original aldoxime/modifier ratio is reached. The plant tests started in April 2013 at Mantos Blancos and in June 2013 at Mantoverde and will last 18 months. The results obtained at the laboratory and pilot plant scale, together with the preliminary plant test results, are discussed in this paper.

Jan 1, 2014

By Raj K. Rajamani, Sanjeeva Latchireddi, Dave Plummer, Stieger. Julius

"The Cortez SAG mill liner design was recently modified to improve the mill operating energy efficiency. In addition to the expected benefits, the design change unexpectedly played a significant role of verifying the proposed grinding circuit expansion for the Cortez Hills project as well as illustrating current SAG mill bottlenecks. This paper will discuss the evaluation that addressed operator and maintenance observations, historical PLC data, and “crash stop” observations to calculate charge motion, packing between lifters, and slurry transport through grate and pulp lifters. The performance of the new shell and discharge end liner design will also be discussed as well as their role in the possible grinding circuit expansion.INTRODUCTIONCortez Gold Mines commenced operation in 1969 with its Mill #1 rated at a nominal 2,000 tons per day. Mill #1 was a conventional rod mill/ball mill circuit followed by an 8-stage CIP circuit. In the early 1990s the Pipeline deposit was discovered and subsequently Mill #2 was built to process the increased ore volume. Mill #2, commissioned by mid-1997, was a nominal 10,000 tons per day SABC circuit followed by an 8-stage CIL circuit for oxide gold ore. Operating and commissioning details of this plant have been published elsewhere (Powell, 1998). By late 1998, there was not enough ore volume to support both mills so the smaller, higher operating cost Mill #1 was put under its current care-and-maintenance status. A simplified flowsheet of the current mill is shown in Figure 1."

Jan 1, 2007

By J. Arzúa, L. R. Alejano, P. Rodríguez

"After six years of successful excavation with a support pattern, some roof failures have happened in a room-and-pillar underground mine excavated in a stratified rock mass. In this paper the authors give explanation to the failure mechanisms and redesign the support in order to solve the recent problems of roof collapse.The mine magnesite bed is approximately 14 m thick, and runs in the E-W direction with an average dip of 18º. Due to mine constraints and the inexpensiveness of the ore, the excavation is performed in the mineral itself, avoiding the need for mining dumps and waste excavation. Rooms and drifts are oriented in the mineral bed to form an angle of 19º to the direction of the strata, reducing the roadway gradients to less than 10%. Rooms are 11 m wide and pillars are 7 m wide. They are designed to leave a 1 meter thick magnesite bed in the roof of the rooms to act as a supporting beam. More often than not there is a layer of very poor quality marly shale up to 1 m thick resting on this magnesite beam. Overlying the poor quality marly shale there is a better quality, self-supporting sandy shale.First, a characterization of the involved rock masses (magnesite, marly shale and sandy shale) is performed, collecting the geotechnical data from previous studies and performing new in situ geotechnical surveys to check the soundness of the previous data. A distinction is made between different quality zones inside the mine by means of the RMR classification, associated to the occurrence of faults.Back-analysis of some of the failed rooms and data from exploration boreholes were used to determine the failure mechanisms (associated to Voussoir type failure or occurrence of sub-horizontal joints crossing the magnesite beam). Based on an understanding of these mechanisms, the method used to design the support and to analyse the stability of the different possible thicknesses of the magnesite beams of the rooms and drifts of the mine were improved.Numerical models of the most relevant conditions found in the mine were performed in order to better understand the behaviour of the roofs and to simulate the supporting solution obtained in the previous analyses.The most relevant conclusions with regard to the main elements causing the instability of the roof include the appearance of sub-horizontal discontinuities crossing the magnesite bed, which causes total instability of the roof and usually leads to collapse; excessive thinness of the magnesite beam in the problematic rooms; and the relaxation of the deepest (around 270 m in depth) zone of the mine because of existing faults."

Jan 1, 2015

By Glenn Lyle

CEMI Introduction ? Introduction of the CAF VOD Project ? Who are our partners ? What were the Outcomes What is next

May 1, 2011

By L. A. Wolff

"Second only to profitability (but not always), productivity is the most sought after measurement for any operation. Whether it is in mining, manufacturing, construction, etc. managers always strive to know how productive various plants are. Even the Federal Government publishes, on a monthly basis, the nation's factory productivity.Just what is productivity and how is it measured? This is an important topic that needs definition. There could be as many definitions as there are people who use it. In the mining industry, the problems of defining productivity would be analogous to the definition of equipment availability. Each mine uses its own definition to suit its needs. Webster's Dictionary primary definition for productivity is ""the quality or state of being productive,"" with productive defined as ""yielding or furnishing results"". This is a definition open to a broad spectrum of interpretation. A discussion of how productivity is currently used in the mining industry and an introduction to a potentially different approach to its measurement and meaning is outlined in this paper.Defining ProductivityTypically, mining companies track several productivity measurements. These are partial measurements and can include the following: tons per specific equipment class per hour, tons per man-hour, and tons per shift. None of these, however, track the over -all performance of an operation, that is, a unit of output per unit of input.For a mining operation, output can be defined as the ore for mill processing. Input includes the required components of production to get to the first step of processing.The Wesbter's Dictionary definition exposes a major fault with the popular conception of productivity. The critical issue of accountability is not addressed. Anyone can produce, but how effectively and efficiently are resources used? Quantifying an inclusive measurement is difficult. Statistics referring to an increase in productivity are often published. For example, the statement, ""gold productivity has increased by 10%"" has many potentially dangerous interpretations. One must ask, what does this statistic really mean? Did gold companies produce 10%” more gold? Were less manpower or equipment required to produce an equivalent amount of gold? Did ore grades increase resulting in more ounces of gold produced, etc. Obviously, production and productivity can have similar implied applications and meanings, but also two different, distinct meanings.To mitigate a problem like this, a standard needs to be developed within the mining industry. Because of varying ore grade, changing waste to ore ratios, different mining methods, etc. all of which have a direct impact on mining, development of a standard is difficult. Kendrick(1) discusses the use of Total Factor Productivity as a productivity measurement. Summarized, the total factor productivity measure is an indication of the change in efficiency in the use of basic labour and capital factors associated with an operation.This measurement has little meaning when calculated at a given point in time. It should be an on-going calculation, tracking an operation over a period of time, to evaluate trends. These trends can be used to determine if an operation's productivity is what it should be, and alert management to potential problems before they become major and adversely affect an operation."

Jan 1, 1990

By V. K. Alves, P. Lacoste-Bouchet

"In order to improve ball milling efficiency in pellet plants, Companhia Vale de Rio Doce (CVRD), experiments new technologies in Santa Luzia (M.G. – Brazil) research center facility. Pilot grinding tests have been conducted with small grinding media, size distributed from 5 to 12 mm (millpebs). Tests have been conducted on an ore from CVRD’s Usina 5 Nibrasco, in Vitória, Brazil. A sample of the regrind circuit feed was obtained and used in comparative grinding tests with two types of grinding charges: 100% balls and a blend of 70% balls and 30% millpebs by weight. According to the balls charge, the grinding media mixture has showed the following benefits: reduction of the power draw, increase of the mill capacity, reduction of the circulating load and a slight increase of the ore blaine. Following the results obtained, CVRD is going to further explore the regrinding potential of millpebs.INTRODUCTIONCompanhnia Vale de Rio Doce (CVRD) is the largest mining company in Brazil. CVRD is involved in iron, base metals, gold and bauxite industries. Moreover, CVRD operates several pelletizing plants in Brazil and in other countries.CVRD’s Center for Mineral Development (CDM) is equipped with a comminution laboratory where studies are carried out that are aimed at the optimization of existing plants at several mining sites as well as the technical and economical feasibility of new projects. The laboratory is equipped with a complete pilot plant with several mills and classifiers, enabling continuous grinding tests in grinding circuits of almost any capacity and characteristics.In January 2004, CVRD undertook to conduct pilot tests at CDM for evaluating the effect of small grinding media on ball milling efficiency. Generally speaking, it is not unusual for operators to add oversize grinding media in a conventional regrind mill. This fact originates mainly from prohibitive prices of grinding media below 25 mm due to a production capacity decrease of grinding media producers. The size of the grinding media is an important variable affecting ball milling efficiency. Use of oversize balls decreases the number of contact surface between which particles can be caught and broken."

Jan 1, 2005

By B. K. Gorain

"Mining of low grade deposits along with the necessity to design and operate flotation plants with high throughput presents unique metallurgical and environmental challenges that need to be addressed to reduce project risks, improve project economics and to sustain the profitability of operations over the life of a mine.Some of the key challenges include poor recovery of valuables and concentrate quality issues due to complex mineralogy, the need to use poor quality water such as brackish or sea water, reduce environmental impact from the process, integrate flotation with hydrometallurgy to maximize precious metals recovery, escalation of capital and operating costs, use of unproven technology and constant use of larger equipment, high ore variability, worldwide constraints on resource availability and an ever increasing need for reliable metallurgical input to financial models to evaluate economic viability of projects.The need to develop effective solutions to these problems in a short time frame is becoming a key to the success of any new project and for sustaining the profitability of a mining operation. An integrated multi-disciplinary approach to problem solving is crucial with the need for a more lateral thinking along with strategic partnerships and collaboration involving the right players. This paper presents some of the key findings based on experiences from a large number of challenging projects and operations focused on solving complex flotation problems.INTRODUCTIONThere are various challenges to successful development of mineral resources. Most of the ore bodies being developed worldwide are of lower grades requiring high throughput to be economical. Many of these lower grade ore bodies have significant occurrences of deleterious and penalty elements that could render a deposit unworthy of development unless there are technological solutions to process them economically in a sustainable manner with minimal environment impact. Even in a high metal price scenario, the economics for some of these low grade ore deposits is marginal. From a metallurgical perspective, these ore bodies have complex mineralogy and use of any off-the-shelf technology is not always effective. Development of innovative processes are often required, which is expensive, time consuming and sometimes risky.This paper will provide a broader perspective on the steps involved in solving flotation problems for operating plants and for greenfield development projects. In addition, some of the key challenges in flotation of complex low grade ore bodies along with the economic benefits in developing solutions to these problems will be presented."

Jan 1, 2013

By Nur Hossain

Analysis of Linear Elastic Fracture Mechanics (LEFM) parameters via the Virtual Crack Closure Technique (VCCT) has been obtained for SENT specimens of AM6OB magnesium alloy. It is found that the ratio of initial crack length to initial monotonic plastic zone size is one of the critical factors in establishing the limit of using LEFM concepts. If this ratio is below a certain limit (i.e. when crack lengths are small in comparison to the plastic zone size), loss of similitude stress intensity factor range would occur. This ratio, therefore describes the limit below which the use of LEFM would no longer be admissible. In this study, the above limit was established for AM6OB magnesium alloy. In addition, fatigue crack propagation tests were conducted to ascertain the fatigue response of the alloy and determination of the coefficients of Paris model.

Jan 1, 2010

By George Langford

Introduction The manufacture of ductile beryllium is one of the most interesting problems confronting the metallurgist today, and one which becomes more acutely in need of resolution as the uses for beryllium increase. Beryllium belongs to the hexagonal group of metals along with zinc, magnesium, and cadmium. The melting point of the metal is about l,280°C. ± 5°C. Due to its brittleness, the metal's chief use is as an alloying agent in copper and iron. The very low absorption coefficient of beryllium for x-rays makes the metal ideal for x-ray windows, and, similarly, because of its low capture cross-section for neutrons, the metal holds a preferred position in the field of nuclear physics. Modern x-ray diffraction methods were responsible for explaining the mechanism whereby zinc metal, although brittle when cast, could be made ductile at room temperature. The diffraction patterns of zinc and magnesium indicated the importance of the parts played by recrystallization and preferred orientation in this mechanism. It was with this in mind that an Investigation into the recrystallization and preferred orientation of beryllium was undertaken. The investigation, after a series of experiments were conducted involving cold-working and annealing, yielded the recrystallization temperature curve for beryllium. Both the texture of preferred orientation and the effect of recrystallization on this texture when subjected to compressive stresses were also determined. W. A. Alexander, J. K. Swinton and L. M. Pidgeon, in their investigation on The Recrystallization Temperature of Beryllium (C.I.M., Trans., Vol. L, 1947, pp. 657-664), utilized the Brinell ball as a means of cold straining the metal. Sufficient strain was developed to estimate the recrystallization curve for the metal. However, in order to determine the lowest recrystallization curve, as well as to develop a preferred orientation structure, larger deformation than that possible with a Brinell ball was required. The plastic deformation of beryllium can be carried out at an elevated temperature above 700°C. Above 850°C., the metal is quite ductile and can be readily compressed or extruded.

Jan 1, 1948

By N. C. Clarke, T. Wu, M. le Page, K. J. Henley

In 1994, Gold Mines of Australia acquired the rights to the historic Mt. Lyell copper mine, and commenced a feasibility study into expanding production. The ore was shown in laboratory tests to be amenable to single-stage, fully autogenous grinding (FAG) of underground primary crushed ore, and this route offered the maximum cost savings. Flotation performance was expected to be enhanced. However, batch flotation tests done on pilot-scale autogenously ground ore showed an unexplainable recovery loss of between 1% and 3% Cu. A second pilot plant trial was set up, with the ground product feeding banks of continuous flotation cells. Single-stage FAG, FAG/pebble and FAG/ball circuits were tested. Results confirmed a loss of about 3% recovery with single-stage FAG milling. The FAG/ball circuit eliminated the loss. Analyses showed that autogenous milling probably produced more chalcopyrite/pyrite composites and that recovery of these composites was lower than with ball milling. In size fractions below about 53 microns, where chalcopyrite is well liberated, copper recovery was less affected by milling method and flotation selectivity. It was concluded that for the Mt. Lyell situation, a FAG/ball circuit was preferable to single-stage FAG.

Jan 1, 2002

By S. Vézina

"Environmental issues have evolved considerably over the last two decades. The main contributing factors, to name a few, have been the changing values of modern societies, the globalization of trade and its linkage to environment, evolution and accessibility of information/ communication technologies, better educated populations, climate change, longrange transport of toxics, and the increasing influence of non-governmental organizations (NGOs) and multistakeholders.As well, environmental management of mining operations had to be realigned in order to meet these new realities and to respond to more elaborated regulations oriented towards sustainability. Environmental management systems (EMS) have been introduced not only to meet the assurance of conformance to regulations but also to prevent pollution and to apply the continuous improvement of mining processes. EMS models have been developed and certification to international standards (e.g. ISO 14 001) is a common practice.Globalization has imposed challenges and additional requirements to environmental management. Developing a mine in emerging countries implies dealing with international agencies, local communities, and a variety of government infrastructures to obtain a “license to operate.” This paper presents the structure of Cambior’s EMS and how it was adapted and implemented in the company’s operations in South America."

Jan 1, 2006

By M-A. Thibault, C. Gauthierand, J-P. Gagné, J-Y. Carrier, P. Rochette, G. Dufour

"Anodes consumed in the electrolytic process for the production of primary aluminium are considered consumables. They need to be replaced at regular intervals of approximately one month. The so-called anode butts which are removed from the cells contain the metal supports (rods), cast iron thimbles, crust and residual carbon material. Cleaning is required to remove the waste material before the rods are fitted with new anodes. An inspection is required to ascertain if repairs to the rods are necessary. Presently, this inspection is carried out visually. ST AS, in 2004 in collaboration with Alcoa Deschambault, initiated a project to develop an automated inspection system. The objectives were to find a technical solution to perform these functions with a very simple system. Such a system was designed around ""artificial vision"". The inspection system developed uses only one fixed camera and does not require a separate step in the process for the anode rod assembly to perform the inspection. This paper presents the initial steps of the work performed to develop the equipment and software as well as a summary of the results obtained."

Jan 1, 2006

By Victor F. Hollister

Although none of the deposits discovered to date are of commercial interest, significant Frasch or sedimentary sulphur deposits occur in Nova Scotia in Upper Mississippian Windsor Group evaporites. The evaporite sequence contains anyhdrite and is overlain by impervious shales. It also is •mderlain by locally petrolijerous formations that are potential hydrocarbon sources for reaction with anhydrite to produce sulphur. As well, stratigraphic and structural traps are known to exist in the Mississippian beds. Thus, all requisites for sedimentary sulphur genesis are present. Available hydrocarbon sources are numerous but small; resultant sulphur deposits therefore are probably also small. Evidence suggests, however, that sulphur deposition is not yet complete.

Jan 1, 1984

By J. G. Trowell, D. T. Magee

"IntroductionMANITOBA'S first recoverable oil was from the California Standard Daly 15-18 well in Lsd. 15, Sec. 18, Twp. 10, Range 27, W.P.M., completed early in 1951. This well also gave the first indications of commercial production in the Williston basin, which is centred in North Dakota and extends into southeastern Saskatchewan and southwestern Manitoba. Subsequent drilling in the Manitoba portion of the basin has revealed the presence of oil in more or less commercial quantities at various points in an area approximately four townships in width and eleven townships in length.Although completions are becoming fairlly widespread in the area,, the greatest activity is in the Virden district, with the focal point in the Daly field, approximately eight miles west •of Virden. This paper will deal with the various techniques used in the Daly field, where •completion problems have •been unusually troublesome. Techniques evolved there have already been used with success at other points in southwestern Manitoba.Daly FieldGeneralTo date, all production has been from the Mississippian limestone, which is encountered •at depths ranging from 2,300 to 2,500 feet. Pay thickness is •of the order of 100 feet. Core analyses indicate average porosity .of about 9 per cent but permeabilities, for the most part, are extremely low."

Jan 1, 1953

By M. J. Bennett

"The Quemont mill is primarily a differential flotation plant in which copper, zinc ·and pyrite concentrates are recovered. Native gold and silver are associated in the form of very small particles with all the minerals contained in the ore: chalcopyrite, sphalerite, pyrite, pyrrhotite and even with the gangue rock.The proportion of the precious metals associated with chalcopyrite is such that the best recoveries we can obtain in our copper flotation concentrate are about 68 percent of the gold and 51 percent of the silver in the ore. In order to improve the total recoveries of precious metals it is necessary for us to float a pyrite concentrate and treat it by cyanidation for the extraction of whatever gold and silver we can obtain in this manner. In spite of the fact that we regrind this concentrate to 75 percent minus 325 mesh, it is difficult to extract much more than half of the gold and silver present in the pyrite concentrate by cyanidation, due to the extremely fine association of the precious metals with the pyrite.In 1963 we produced 86, 853 ounces of gold by flotation alone, and 14, 581 ounces by cyanidation. These amounts are equivalent to 67. 74 percent and 11. 37 percent respectively of the total gold in the mill heads. Comparable figures for silver are 325, 598 ounces by flotation and 99, 448 ounces by cyanidation; and 51.37 percent and 15.69 percent respectively.The total annual value of these metals (ie. $4, 000, 000) qualifies Quemont as a precious metal producer of some note. On the other hand, it is evident that by far the greater proportion of our gold and silver production is contained in copper flotation concentrates. When the income from our copper and zinc values is also considered, the much greater importance of flotation compared with cyanidation in our operation may be appreciated. This explains why most of our work on controls has been carried out in our flotation mill, rather than in our cyanide plant. Some of the controls installed there could, of course, be employed in either type of operation."

Jan 1, 1964

By Ammar Henchiri, Godfrey McDonald

"Breakwater Tunisia S.A.’s Bougrine mine is located in a semi-arid region of Tunisia characterized with Triassic salt domes. Due to extensive agricultural land-use in the mine area, a zero-discharge tailings facility design and high water recycling rates to the concentrator were chosen to minimize the environmental impact to the surrounding area. The process water supply has an accumulation of naturally occurring salts due to high evaporation rates and the need to recycle 70% of the process water. The elevated salt content of the process water had an adverse effect on the efficiency of the flotation reagents, which led to the production of lower quality zinc concentrate.The impact of the water salinity on the zinc concentrate quality was documented by extensive laboratory testing. A feasibility study was prepared and a decision was made to install a reverse osmosis plant for desalination of the process water. A plant was constructed during 2000 which currently supplies up to 85 m3/hr of de-mineralized water, equal to approximately 45% of the total process water requirement. The salinity of the process water decreased from about ten to twelve grams per litre before installation to less than three grams per litre. Together with minor modifications to the concentrator circuit, Breakwater Tunisia has increased the zinc concentrate grade from 50% zinc to 55% zinc without compromising the zinc recovery rate of approximately 84%.This paper discusses the successful installation of a reverse osmosis desalination plant that allowed Breakwater Tunisia to improve the grade of the Bougrine mine zinc concentrate and thus increase the net smelter return.INTRODUCTIONBreakwater Resources Ltd., through a wholly owned subsidiary, Breakwater Tunisia S.A., acquired the Bougrine underground mine and related assets in September 1997. During the winter of 1997/1998, the company rehabilitated the mining equipment, modified the concentrator circuit, developed a new mine ventilation system and prepared a new mining production program. The mine was reopened on May 2, 1998 and was in commercial production as of June 1, 1998."

Jan 1, 2003

By Clifford Britch

"Kaiser Resources Ltd. has completed feasibility studies resulting in approval of $277,500,000 for the construction of the Greenhills Surface Coal Mine, located 35 km north of Sparwood, B.C. The mine is designed to produce 1,800,000 tonnes per year of metallurgical coal over a 20-year period by conventional truck /shovel methods.The project is in the detailed design phase, with some preparation work being carried out on the site. The major construction activities are scheduled to commence in the spring of 1981. Mine development and pre-stripping will commence in July 1981, with first coal production scheduled for mid 1983.This paper will be a general description of the project and will include a discussion of the geology, reserves, mine planning, coal handling and processing, surface facilities and related topics.IntroductionProject DefinitionThe Greenhills Project is designed to produce, by conventional truck I shovel surface mining methods, 1,800,000 clean tonnes of metallurgical coal per year for twenty years . The production/ life constraints were chosen as being meaningful for planning and evaluation purposes after appraisal of the market potential and are not limited .by economic reserves. Substantial additional strippable reserves are available which would facilitate expansion of the proposed mine life and/ or capacity. In fact , all the facilities are designed so that they can be expanded in the future to accommodate a doubling of annual coal production, and certain portions of the coal processing facilities are being installed with an annual capacity of 3,600,000 tonnes. The mine will also produce approximately 2,400,000 clean tonnes of oxidized thermal coal, mostly in the first decade."

Jan 1, 1981