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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' 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 T. B. Armstrong

"The energy situation is extremely fragile and there is a critical need for energy conservation in the short term and development of new technology in the long term. To keep tabs on energy flow, separate electrical metering is being installed in major industries: The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production. IntroductionThe current energy situation is best described as being extremely fragile. The supply of hydrocarbons is being exhausted, while energy needs remain the same in the industrialized nations and are increasing in the Third World.The petroleum era is expected to come to an end in this century, as there will no longer be enough. Gas, coal and electricity will replace some oil as it becomes scarcer during the transitional period before new sources are gradually developed.There is a critical need for energy conservation in the short term. Conservation will, to a great extent be justified by savings, as it will be less expensive to recover energy that is being wasted than to acquire additional energy.The gradual development of new technology is expected to increase energy self-sufficiency in the long term. New energy sources are attracting a growing interest and are becoming more economically feasible and marketable. These rediscovered energy sources include peat, wood, solar, domestic waste and windmills.The sudden increase in energy prices has forced users to look at the way energy is used and find ways to save it. Major industries have found that using separate meters for each department spurs energy responsibility. Energy trim programs now rank in importance with profitability, productivity and with quality.Energy has become such a crucial cost element that better data are essential for planning. Data acquisition varies with plant size and number of utilities. Meters can be read manually or monitored by computerized energy management systems.The relationship of electrical energy input and production levels for a typical uranium mine will be used to demonstrate two steps to energy management. The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production.The direct benefit of saving energy is the money saved."

Jan 1, 1980

By L. Xu, A. Vazirnegad, R. Bhaskar

"Respirable dust samples are collected by the U.S. Mine Safety and Health Administration (MSHA) and mine operators as part of the requirements of the Federal Coal Mine Health and Safety Act of 1969. About 1.85 million respirable dust samples collected by MSHA were analyzed for the period 197/ to 1990 f or the states of Alabama, Colorado, Illinois, Kentucky, Ohio, Pennsylvania, Utah, Virginia, and West Virginia. Trends in coal dust exposure of continuous miner operators, helpers, and roof bolters were analyzed. The reported production in each of the nine states was analyzed. The relationship between average dust exposure and coal production was evaluated. In addition, the differences between mine operator and MSHA inspector collected data were examined for each state. The average exposure levels for the three occupations - miner operator, miner helper, and roof bolter were also compared.The data indicate that the dust exposure levels have significantly dropped over the last twenty years, while at the same time production has increased, attesting to the ability of the U.S. mining industry to adapt successfully to the stringent dust control regulations. In terms of coal production, it appears that some states, such as Utah, Illinois, and Kentucky, were able to increase their production significantly relative to others, such as Pennsylvania and Virginia. The relationship between dust concentration and coal production is, at best, tenuous. 11 implies that existing dust control technology has assisted several mine operators to meet dust standards, while, at the same time, enabling mines to produce more coal than before. Comparison of data collected by MSHA and mine operators indicate that MSHA inspectors recorded higher dust levels and lower shift production relative to mine operators."

Jan 1, 1993

By J. G. Mieczaniec

"A key element for a successful mining operation is the grinding circuit. It must be reliable, maintainable and must perform to or above design capacities for 95% of the time. Once chosen and installed major modifications are either prohibitive or very expensive. The selection of mills and drive system is probably the single most important decision made within the concentrator complex.Economic studies revealed that the optimum expansion for Lornex was a 68070 increase in plant throughput. This indicated that the new mill would have to be larger than the existing 32-foot mills.The Lornex ore has a wide range of hardness which has a great influence on mill throughput (Table 1). To smooth out the fluctuations, a blending program ha s always been in effect at the mine, coupled with ball charge control for the primary mills. As can be appreciated, an optimum balance is difficult to achieve. A sudden change to so ft ore with a high ball charge puts the mill into jeopardy with a constant speed drive.The blending philosophy would continue after the expansion, but with the addition o f a second primary crusher, it was envisaged to have the new mill process harder ore to maintain production targets. In theory, the new mill would have a proportionally larger ball charge than the existing mills.This scheme makes the new mill vulnerable for processing so ft ore for a fixed speed mill. Thus, variable speed seemed an attractive option.In early 1979 prominent mill manufacturers were invited to submit proposals for the new mill. Fixed and variable speed mills were to be considered with mill size ranging from 32-foot to 36-foot diameter. The proposals were to include the drive, ball charge, % critical speed, couplings, gears, etc. - the complete package. The recommendations were variable with most manufacturers submitting several alternatives. Recommended mill speed varied from 68% to 78% critical. The concept of variable speed was becoming more attractive, Lornex ultimately decided on a twin DC motor drive system for a 34-foot diameter mill.After start-up it was discovered that best throughput occurred at 78% mill critical speed or top speed of the drive motors. The motor speed range was adjusted upward from 186 to 200 rpm. Today the mills operate at 79% critical or a motor speed o f 193.5 rpm.Clearly the selection of the variable speed mill allowed the flexibility to optimize throughput."

Jan 1, 1985

By V. Casaletto, L. Widdifield

"Equipment capable of higher production requires more sophisticated maintenance programs. Reduction in fleet size through use of modern technology requires higher utilization and greater availability.Caterpillar's maintenance control system program will display the capabilities of computer technology in a maintenance environment. Utilizing such resources will provide increased availability, lower inventory requirements, fewer last minute disasters, and lower cost per ton for maintenance. High production from premium quality equipment can only be realized by good planning and care, executed by maintenance and operating personnel acting cooperatively.IntroductionAccurate information is the key to accurate decision making. In today' s competitive industry, mines' teeter on the brink between profit and loss. Their survival will be predicated on their ability to retrieve, analyze, and formulate accurate information regarding the operating costs of mobile equipment. This paper deals with the impact of managing equipment operating costs and introduces a computerized control system designed for monitoring and controlling these costs.Caterpillar has produced several computerized modelling programs designed at taking the guesswork out of equipment selection and management. The most basic of these is Equipment Investment Analysis (E.LA.). This model compares two or more alternatives over the projected life of the machine. It deals with comparisons in the owning and operating costs of equipment, as well as, their productivity over a specified period of time.In order for a computer model to be credible as a decision making tool, it must be simple to use and the results measurable and accountable . Caterpillar has introduced several computer models to assist the industry in selecting and managing an equipment fleet. These programs allow the user to predict and monitor the operating costs and production capability of mobile equipment. They are often used in the machine selection process to select equipment which best suits a mine's long-term cost per tonne objectives. After the optimum selection is made, Caterpillar's Maintenance Control System provides an efficient management tool to monitor and verify the data used in the selection model."

Jan 1, 1992

By A. P. Annan, C. Vaughan, J. L. Davis

"The increase in the price of gold in the last decade has led to a significant increase in placer gold exploration around the world. Placer deposits occur in depressions along ancient river and stream channels. In many instances these channels have been covered with materials from subsequent erosion or glacial activity. Use of high resolution geophysical techniques to map the bedrock and soil strata can improve the probability of finding the placer deposits. The buried bedrock surface is often the factor controlling deposition. Ground probing radar (GPR) and seismic sounding are complementary high resolution techniques which can be used to map bedrock topography and soil stratigraphy. Current radar instrumentation operates best in coarse grained soils. Penetration depths of up to 30 m have been attained, and stratigraphic horizons have been mapped with spatial resolutions of better than 0.5 m. Although its spatial resolution is considerably lower than radar, seismic refraction achieves superior penetration in electrically conductive environments such as clay and silt soils. The seismic refraction method thus proves to be complementary to the radar method. The seismic reflection technique is analogous to the radar method. Current seismic instrumentation and energy sources however, provide limited spatial resolution and do not permit the use of seismic reflection for reflecting horizons lying at depths of less than 20 m to 30 m.This paper discusses the use of both radar and seismic methods for placer exploration. The radar and seismic techniques, the field methods, equipment, and data displays are presented. Field examples illustrate the high resolution mapping capabilities of the techniques. The radar data agree very well with the borehole data, and in fact they sometimes detect subtle soil strata variations associated with changes of permeability that were not initially observed in the borehole logging. The radar method can give a continuous profile of water table and bedrock in sandy soils. Seismic soundings are necessary to differentiate between the bedrock and water table reflections observed by the radar."

Jan 1, 1987

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 D. A. Whitley, R. P. Carone

"Shaft sinking by blind shaft drilling offers the mining industry a different approach to shaft construction. Benefits of blind drilling include less construction time and fewer safety problems while being economically competitive with conventional sinking methods under certain conditions in soft water-bearing formations.Technology and equipment are available today to drill blind shafts up to 20 feet in diameter and over 3,000 feet deep. This is an extension of proven techniques developed by private industry. To drill a 20-foot-diameter hole requires a specially manufactured hydraulic drill unit with a rated hoist of 2 million pounds and 500,000 ft/lb of torque.This brief paper will discuss the following aspects of blind shaft drilling:• benefits;• equipment f or drilling;• circulation methods;• circulation rates;• shaft alignment; and• progress rates to be expected.IntroductionShaft sinking by blind shaft drilling offers the mining industry an improved alternative method for shaft construction. The drilling of 20-foot -diameter shafts to depth s in excess of 3,000 feet can be done by utilizing equipment specifically designed and manufactured for this purpose. This equipment is a distant cousin, at best, to standard oil well drilling rigs. The actual configuration of these drilling machines will be discussed in a following section.To most people, the drilling of mine shafts may seem to be a new or at least recent technology. However, as early as the I920s, German and Dutch miners were drilling 24-foot diameter shafts for coal mines. What led ""miners to resort to drilling were ground conditions that still plague conventional shaft sinkers today: soft , water-bearing formations. The major drawback to these early attempts was a low penetration rate due to poor bit cleaning. This, however, was not the end of early shaft drilling activity. From 1936 to 1950, a number of mine shafts were drilled by the calyx or coring method. This method employs a bit with a large bore down the center. As it drills, it cuts a core. When 5 to 10 feet have been drilled, the core is removed. The disadvantages of this method were the time and work involved in breaking the core free, and it was inadequate in soft water-bearing ground."

Jan 1, 1981

By Yve Potvin, Fabien Miller, Deni Jacob

"The potential benefits of measuring the volume of excavated open stope are enormous. By comparing the actual excavation of the orebody contour, it is possible to determine the “real"" amount of mining dilution, as well as the ore left in place. Noranda Technology Centre (NTC) has developed a stope survey system that allows for rapid and efficency measurement of underground excavation volumes. The system comprises a commercial laser (EDM), a set of lightweight rigid rods, an automated scanning device, a hand-held data logger and Autocad-based software for volume calculation and drawing manipulations. This system is an invaluable tool for the optimization of mining practices and could also be used for the delineation of caving areas. IntroductionDilution is the most common and devastating mine disease. In some instance s it could systematically cripple the profitability of an operation, and in more extreme cases, it could become a terminal disease. The Northern Miner Journal (1988) reported that many operations have closed because of uncontrolled dilution, before they even completed their first year of production. Other studies (Pakalnis, 1986) have shown that approximately 40% of all underground operations have more than 20% dilution. Considering that every ton of dilution costs in the neighbourhood of $25, the total direct cost of dilution to the Canadian mining industry may easily exceed one hundred million dollars per year.The dilution problem has been a major concern to Noranda Minerals for some time and in 1984 a comprehensive geomechanics research program was created at the Noranda Technology Centre (NTC) in Pointe-Claire, Quebec. New empirical stope dimensioning techniques, improved cable bolting techniques and sophisticated ground control instrumentation projects (among others) contributed to improve the situation at Noranda.A new tool which can have a very significant impact on the ""war"" against dilution is being developed by the Noranda Technology Centre. The laser stope survey system allows for a rapid, precise and efficient measurement of underground excavation volumes. By comparing the excavation planned with the volume excavated, it becomes possible to determine the ""real"" amount of mining dilution as well as the amount of ore left in place. This system may also allow for early detection of a dilution problem. As with other diseases, early detection provides the best chance for a fast and total recovery."

Jan 1, 1992

By B. J. Morash

"The six mineable coal seams at Quintette Coal Limited (QCL) vary in thickness from 0.75 m to 9.5 m and include waste partings varying from 0.2 m to 4 m thick. They are highly faulted and folded in some areas and dip angles range from horizontal to 90 degrees. Due to this variation in thickness and orientation of the seams, several different mining methods are used to economically recover the coal. The key to QCL's success depends on the timely collection of detailed geology data and on mine planning which allows selection of the best method of mining to achieve maximum coal recovery with minimum dilution and cost. This paper will describe the various methods of mining employed and the criteria used to select them. IntroductionThe unique challenge in mining at Quintette Coal Limited (QCL) is to extract large quantities of coal and waste, from a complex orebody, with constricted areas for mining and waste disposal. This challenge dictates the mining of as many as four 10 m benches per year, through complicated coal structures, from three separate open pits with a total of 19coal seams. This total is made up of 5 seams at Mesa, 8 seams at Wolverine and 6 seams at Shikano (Fig. 1).Computer-aided mine planning and scheduling is carried out on a routine basis to ensure the prudent selection of methods of mining, to produce plans which ensure delivery of appropriately blended raw coals to the plant, and to coordinate disposal of large quantities of waste.Several methods of mining are applied simultaneously, and as a result coal and waste are released at different rates from twelve to fifteen mining faces. In order to ensure uninterrupted flow of material from excavators and to utilize the truck fleet effectively, a computerized dispatch system is employed.GeologyGenerally, geological conditions can be related to structural (faults and dip angle) and/or stratigraphic (coal seam and interseam waste thickness) sources"

Jan 1, 1988

By Margaret Armstrong, Normand Champigny

"In order to define the ore/waste boundaries as finely as possible, mine planners often insist on kriging very small blocks (relative to the spacing of the available sample grades). These kriged estimates for very small blocks are over-smoothed. For several simple sampling patterns the relation between the block size and the variogram range, and the kriging results were studied. Meaningful estimates of individual blocks are obtained when the variogram range is large compared to the block size and the sample spacing. For a variogram range less than half the sample spacing, the kriged block estimates were found to be uncorrelated with the actual grades. The over-smoothed estimates should not be used for calculating recoverable reserves IntroductionNowadays most mining companies know that the geostatistical estimation method, kriging, gives the best estimates of individual block grades. As most mining engineers and geologists are now familiar with the basics of geostatistics, the definitions of the variogram and of kriging will not be repeated here. An appendix containing definitions and mathematical formulae is given at the end of the article for the sake of completeness.Even though it is well known that the kriged estimates for very small blocks are over-smoothed, mine planners who want to define the ore/waste limits as finely as possible (especially for selective underground mines), are often tempted to krige very small blocks. So we ask how small is ""too small""? Clearly ""smallness"" cannot be expressed in yards or metres. It depends on the nature of the orebody under study and this is reflected in the variogram model chosen to describe it.To answer this question, the changes in four key parameters are studied:• the kriging variance,• the variance of the estimated grade,• the slope of the linear regression of the true but unknown grade on its estimated value, and• the coefficient of correlation between the actual grade and the estimate;as a function of the block size and the variogram range. (Definitions of these four parameters are given in the appendix). Ideally the kriging variance should be low while the variance of the estimated grade should be equal to the dispersion variance of the blocks. The slope should be l.0, and so should the correlation coefficient.These four parameters have been calculated for several simple 2-D examples. The effect of using kriged block estimates that are too heavily smoothed are shown, and the dangers of trying to use these estimates for predicting recoverable reserves for small blocks (i.e. selective mining units) are pointed out."

Jan 1, 1989

By M. Fiset, D. Belley

"In a country with a cold climate like Canada, the cost of replacing worn parts of snowplow equipment such as blades, is enormous. A great number of these blades, with different mechanical properties, are available, but few studies have been made to facilitate their selection according to the optimum ratio of wear resistance to cost. To attain this objective, laboratory tests have been developed to correlate test results with those obtained from various road service conditions. The test materials, 1080steel in different states, 1090steel and D2 tool steel were chosen to cover a large range of hardness as well as different microstructures. Comparison of the wear resistance of the test materials relative to a standard revealed, in general, a good correlation between the results of laboratory cement wheel tests and those from the road tests. Scanning electron microscope wear scars of laboratory and road test specimens were compared to determine if similar removal material mechanisms were active. IntroductionThe correlation between wear tests in service and those performed with the aid of laboratory equipment is of considerable interest. As a matter of fact, when correlation exists, it allows one to evaluate the performance of a material more easily and at a lesser cost.Several authors have been interested in simulating various wear circumstances, especially abrasive wear. In particular, the simulation of wear on implements working the soil, has been tried by several researchers. The test using the rubber wheel seems to be the one giving the best correlation if the soil is rather sandy and its moisture content is low.However, little work has been done concerning wear on the plates, skids and blades attached to trucks used for levelling and snowplowing; such equipment being used to keep roads clear in northern countries. Now each year, companies introduce new steels by claiming that the y are superior in performance to their predecessor s and invariably they demand a higher cost. The study presented here was undertaken with the goal of making a more judicious choice of these steels by simulating their behaviour in the laboratory. Moreover, the examination of abraded surfaces in situ and in the laboratory have also allowed us to compare the respective wear mechanisms."

Jan 1, 1988

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'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 J. C. Farge, G. G. Jacox

"The effect of composition and process variables on the properties of Marmet and modified Ni-Hard slugs will be discussed. The effect of shape on the over-all performance of grinding media and its relationship to the end product has been investigated and these results will be reviewed. The cost effectiveness of slugs versus balls has been well documented under various operating conditions and this will be presented in detail. IntroductionThe selection of grinding media is based on the cost-to-wear ratio combined with the consideration that it must stand repeated impacts without breaking. The specific consumption of grinding balls and slugs depend s primarily on their microstructure, which, in turn, is controlled by chemical composition and by the processing condition s to which they are subjected during manufacture.The first part of the paper deals with the optimization of alloy content to produce cast iron slugs with microstructures consisting essentially of marten site and carbide and with hardnesses of 650 B.H.N. and higher. The second part deals with the use of cast iron slugs to replace forged steel balls in milling operations. Tests indicate that the shape of the media has little if any effect on results and that the hardness and microstructure of grinding media will determine its relative performance. Where price is equal under similar milling conditions, the cost effectiveness of cast iron slugs is superior to that of forged steel balls.In order to understand the role of alloying elements in cast irons, it is judicious to consider the microstructural changes which occur during solidification and subsequent cooling, particularly from above the eutectoid temperature. Solidification starts with the formation and growth of dendrites of austenite. As temperature decreases, an eutectic nucleates and grows between the primary austenite dendrites. Immediately after solidification, the structure consists of proeutectic austenite and of the eutectic of primary carbides and austenite. On cooling to room temperature, austenite can be retained or transformed to pearlite, bainite, martensite or mixtures thereof. It is necessary to prevent decomposition of austenite into pearlite by adjusting the alloy content. For best abrasion resistance, the microstructure at room temperature should consist essentially of martensite and carbide. These types of irons are known as ""martensitic irons "". Depending on the carbon content of the austenite and cooling rate, "" as-cast"" martensitic irons will contain different proportions of martensite and retained austenite. The more carbon in solution in the austenite, the lower is the temperature at which the martensitic transfoemation takes place. By the time room temperature is reached ., the transformation is seldom complete, and the end product is generally a mixture of martensite and retained austenite."

Jan 1, 1980

By V. G. Papangelakis, G. A. Moldoveanu

"One of the major commercially viable secondary sources of rare earths elements (REE) are the ion-adsorption ores, containing 0.05-0.3 wt% rare earth oxides physically adsorbed on the surface of clay minerals formed during weathering of lanthanide-rich rocks. Although ion-adsorption deposits are substantially lower grade than other types of rare earth sources, this is offset by the easier mining and processing costs and the low content of radioactive elements. The conventional extraction route involves ion exchange leaching with monovalent cation solutions (usually ammonium sulphate) under ambient conditions; during the process, the lanthanides are desorbed and transferred into solution, being replaced by the exchange cation. While ammonium ions are known to perform best, there is increased environmental motivation to investigate alternative lixiviants in order to minimize ammonium sulphate impact via reducing its use without considerable loss in rare earth recovery.The objective of the present research was to screen various cations salts (alkaline and alkalineearth metals present as sulphates, chlorides and acetates, respectively) in order to comparatively evaluate lanthanide desorption levels and provide a theoretical base of the observed behaviour. The overall aim is to gain a better understanding of how both cation and anion may affect the overall REE recovery from ores in order to provide alternative processing routes. The distinct exchange capabilities of cations were explained in terms of differences in cation hydration enthalpy within the same group, i.e. Cs+ > NH4 + > K+ > Na+ > Li+ and Ba2+ > Ca2+ > Mg2+. For the same cation, it was determined that REE extraction efficiency follows the order sulphate > acetate > chloride, based on dissimilarities in constants of formation values. Various process implications scenarios will be described based on using mixed cations and anion matrices."

Jan 1, 2016

By Jack A. Caldwell

"This paper describes the nature of stage curves and the ways in which they have been used to study the construction sequencing of tailings impoundments. They provide a simple graphical method of representing the main aspects of an impoundment, studying the effects of different impoundment geometries, and predicting and controlling impoundment operations.IntroductionThe designer o f a tailings impoundment must, at all stages of design, examine the relations hip between the volume o r tones of tailings that will be deposited at the impoundment, the height to which the tailings will rise in a given time and that rate at which the various parts of an impoundment will be constructed. The operator, if he is to keep proper control of this impoundment, must be able to examine the effect of different deposit ion rates at the impoundment, and the consequences of placing more or less tailings at different parts of the impoundment. The operator must be able to see at a glance just how much higher an impoundment will be at the end o f the following yea r of deposition, and this he must be able to do in order to formulate his annual construction and operating plans.Stage curves are the best way o f modeling an impoundment. The simple impoundment model represented by the stage curves enables the designer and operator visually to examine the relationships between the various factors, such as impoundment volume, deposition rat e and construction rate, that dictate the essential features o f impoundment design and operation.A design model for a tailings impoundment may be formulated as in Figure 1. This involves the site selection process, the definition o f a geotechnical model, a water balance model and a height-capacity-production model."

Jan 1, 1983

By W. B. Evans, D. P. Taylor

"ANFO has two major disadvantages as an explosive low density and no water resistance. ANFO can be blended with more dense, water-resistant explosives to form ""Heavy ANFO"". The use of emulsion or slurry explosives as a matrix is an attempt to correct deficiencies on the basic ANFO while maintaining cost effectiveness.This paper will deal with four main topics.1. Product description, general properties, strengths and weaknesses.2. Technology of ""Heavy ANFO"" , the emulsion matrix concept and the range of products.3. Relative performance and economics in various rock types.4: Current status in Canada and worldwide, and the expectations for future use.IntroductionANFO has been in use for roughly three decades and remains the explosive of choice where conditions are suitable. ANFO has not become the universal explosive - despite its undoubted superiority in cost effectiveness - because it has serious flaws as an explosive. The most serious of these flaws are the low density and absence of water resistance in ANFO. Low density implies a low concentration of explosive energy in a borehole. The amount of energy available can be increased by. adding an energetic material, generally aluminum, or by drilling more holes. Both solutions add considerably to the over-all costs. The absence of water resistance in ANFO is inherent and can only be overcome by changing the fundamental nature of the product. The history of research in the explosives industry since the advent of ANFO is mainly the story of the various attempts to overcome one or both of ANFO's shortcomings.Many technologies and products have been developed to compete against ANFO, among them slurries, emulsion s and the so-called Dense ANFOs (ANFO with an extra nitroalkane liquid sensitizer). None of them was cost-competitive with ANFO so their market penetration has been limited to wet holes or specialized applications. The exception to this result is Heavy ANFO. The aim here was to have the best of both worlds, the high density and water resistance of emulsions or slurries plus the low cost of ANFO. The concept is very simple. ANFO is about 50 % air, roughly 30% in the prills and 700,10 between the prills . The air in the prills provides the necessary sensitivity but the air between them is waste volume. If one fills this volume with a high density, waterproof product, the resultant mix will be denser and more water resistant than ANFO and with very little added cost. This basic concept was invented by B. Clay in the 1970s and is covered by Canadian Patent No. 1 115 959. The water resistance and/or density of the mix can be tailored by changing the emulsion level. Because Heavy ANFO is relatively inexpensive while providing good in-hole densities and water resistance, it can compete successfully against all other bulk products. It competes against ANFO and Al / ANFO by allowing pattern expansion and reducing over-all costs while it competes against emulsions and slurries on price."

Jan 1, 1987

By E. Kouame, M. D. Everell, D. Hodouin, T. Kasongo

"To properly operate a mineral beneficiation plant it is often necessary to have systematic evaluations of the performances of a piece of machinery, of one of the processes or of the plant as a whole. The state of a process can be described from a collection of data: flow-rate values, particle-size distributions and chemical compositions of the various streams of the circuit studied.However, the obtained raw data are generally redundant and inconsistent from a material balance viewpoint, due to natural disturbances, sampling errors, unreliable instrument readouts and laboratory analysis inaccuracies.BILMAT is a computer program which upgrades mineral processing data, adjusting them in a least-squares sense in order that they may verify material conservation constraints. As a consequence, the program is able to calculate the best values of unmeasured flow rates (such as. circulating load ratios). BILMAT is a very general program in the sense that it can accept various data from any type of mineral processing circuit, and its usefulness is demonstrated here for three processing circuits of a pyrochlore concentration plant: grinding, desliming and flotation.IntroductionComputation of the material balance around any metallurgical process, that is the determination of stream flow rates and composition. is a basic step in its evaluation. In spite of the simplicity of the physical laws of mass conservation. and of the very fundamental need to know the flow of components around a process. it appears, particularly in mineral processing. that material balancing is quite often a difficult task. The reasons for this involve the problems associated with collecting measurements in mineral dressing plants. They are generally poorly instrumented and invariably data must be collected through hand sampling. Due to the heterogeneous nature of the streams, segregation occurs and samples are often not adequately representative, a problem made worse by the lack of circuit sampling facilities, which are often completely ignored by plant designers. Furthermore, the nature of the measurements to be performed (for instance particle size distributions in the fine range) frequently gives inaccurate results."

Jan 1, 1981

By G. Morin, S. Amira, F. Chiesa, B. Duchesne

"The effect of heat treatment on the stress corrosion cracking (SCC) resistance of A206 aluminum alloy was studied using the breaking load test according to ASTM G139 standard. This method uses the residual strength measurement from tensile test specimens previously stressed axially while exposed to the 3.5% NaCl alternate immersion test according to ASTM G44 standard. The residual strength after exposure is used as a measure of SCC degradation. The performed heat treatments were conventional natural aging T4 and artificial over-aging T7 tempers in addition to artificial aging T6 temper added for purpose of comparison. These tests showed that A206-T4 has a good resistance to SCC at pre-loads up to 100 MPa while A206-T7 was prone to SCC for pre-loads of 0 and 50 MPa, and showed total loss of ductility for pre-load of 100 MPa. Both A206-T4 and A206-T7 pre-loaded at 150 MPa showed poor resistance to SCC. A206-T6 showed the worst susceptibility to SCC as it lost all mechanical resistance at the end of the alternate exposure to 3.5% NaCl solution, even without any applied pre-load. Metallographic analyses showed that the resistance of A206 alloy to SCC is mainly affected by the transition from pitting to intergranular corrosion attack. The initiation and the propagation of intergranular corrosion attack might be assisted by high stress concentrations in the pits, particularly for pre-loads higher than 100 MPa. The transition from pitting to intergranular corrosion can also be promoted by hydrogen evolution from growing pits, which supports crack propagation by hydrogen embrittlement.INTRODUCTIONA206 aluminum-copper casting alloy offers ones of the best mechanical properties among the aluminum foundry alloys. This alloy belongs to the 206 alloy family developed in 1970s as an alternative to the highly resistant and ductile but more expensive Ag containing 201 alloy (Tiryakioglu, 2009). A206 showed also good low cycle fatigue strength (Sigworth, 2002). These properties made A206 alloy a strong candidate for some vehicles lightweighting applications. In spite of its interesting properties cited above, A206 alloy was rarely used in critical applications due to its tendency for hot tearing and stress corrosion cracking (SCC). Studies of A206 hot tearing behavior showed that this tendency can be reduced by improving castability and by grain refinement (Sigworth, 2002). The optimisation of the casting design can also prevent the tendency toward hot tearing. The SCC of aluminum-copper alloys arises from their susceptibility to localised corrosion, especially in chloride containing media. The localised corrosion in Al-Cu alloys is caused mainly by their heterogeneous microstructure, where the presence of many phases leads to the formation of micro-galvanic coupling with the matrix. However, most of the studies aimed at investigate the SCC of Al-Cu alloys concerned the wrought ones, especially the AA2024 alloy tempered in the T3 (or derivative) condition and few studies concerned the 206 and cast alloys in general (Alexopoulos, 2009 and 2016; Campestrini, 2000; Vogt, 1998; Talamantes-Silva, 2008; Zhang, 2003). Jean et al. (Jean, 2009) studied the effect of heat treatment on SCC behavior of B206 alloy using alternate immersion technique and grain boundary attack method and did not find direct correlation between the two methods. Another study showed that aging of 206 alloy may make a casting susceptible to SCC (Sigworth, 2002). However, these studies did not provide detailed information about the mechanisms of SCC of 206 alloy, as well as the effect of the stress intensity. The present work investigates the effect heat treatments and the stress intensity on the morphology of the corrosion attack and the SCC behavior A206 aluminum alloy. This study uses the breaking load test according to ASTM G139. This test utilizes direct tensile specimens and was developed to give a more quantitative measure of the SCC performance of resistant aluminum"

Jan 1, 2016

"Classification is an important aspect of mineral processing plants and if not operated efficiently can compromise both the throughput and recovery of the circuit. This paper discusses the benefits that can be obtained by operating the classification section efficiently. Since classification in most of the mineral processing plants is carried out using either hydrocyclones or screens the case studies involving the two classifiers were used in the discussion. Studies where optimising the classification circuit lead to improvements in the performance of the overall plant were included to show the benefits of paying attention to this area that is usually neglected. The data from the mill-float mill float, SABC, and Single Stage SAG mill circuit case studies have shown that the circuit throughput can be increased throughputs without compromising recoveries by optimising the classification circuit. On plants where the ore can be liberated at coarser sizes the experimental results have shown that replacing hydrocyclones with high frequency screens can lead to improved plant throughput with no detrimental effect on recovery. The discussion has included information on design and operational considerations for plants that have or intend to use hydrocyclones or screens that could make the classification circuit more efficient.INTRODUCTIONComminution is the process in which particles are reduced in size mainly by crushing and grinding. In most mineral processing circuits comminution is carried out to liberate the valuable minerals prior to subjecting the ore to a recovery process were values are concentrated or extracted (Mainza, Evertsson, Benzer, Tavares, Powell, Rule, Plint, …, 2013). Comminution is an energy intensive process and is usually achieved in stages. It is common to have a classifier after each comminution stage. The process is carried out in stages to minimize energy usage and to improve the efficiency of the process in meeting product specifications. Although comminution circuits have been advancing classification has remained a key component of the comminution circuit design (Mainza et al., 2013). Mainza et al. (2013) summarised the role of classification in mineral processing circuit as: to allow material in the equipment that can promote high operational efficiency, to improve material handling in the circuit, to minimize overgrinding of material to reduce energy consumption, and to provide the best quality product to enter the next stage of processing. Classification stages are employed in comminution circuits to ensure that high throughputs are processed without compromising the product quality or to produce a fine product at relatively low energy. If neglected classification can have detrimental effects on plant performance."

Jan 1, 2016