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By M. E. Kuchta

Research is currently being conducted at the Colorado School of Mines in the use of high-pressure water for scaling down loose rocks in underground mine openings. In addition to improved miner safety, the adhesion strength of shotcrete applied to rock surfaces treated with high-pressure water is also increased significantly, resulting in improved overall support capabilities of the shotcrete layer. Test results show an increase in adhesion strength by a factor of four on a concrete test wall cleaned with water at 21 MPa (3,000 psi) as compared to a surface cleaned at 0.7 MPa (100 psi).

Jan 1, 2003

By Andrzej M. Wala

The purpose of this paper is to describe a data-acquisition system for mine ventilation surveys that was developed by the author. The system is composed of "off-the- shelf" components integrated through a custom- designed software package to form a user- friendly data acquisition system capable of operating as an unmanned base station or as an operator-triggered roving station. To demonstrate the applicability of the developed system, the author will present data collected during an experimental ventilation survey, which was performed in one of the underground coal mines of eastern Kentucky.

Jan 1, 1993

By K. J. Stratford

The selection of a suitable mathematical function for the modelling of the partition curve for dense medium cyclones is discussed. Five alternative functional forms (based on three basic functions) are compared using data obtained in the course of an extensive experimental study of a 610mm dense medium cyclone. It is shown that, for adequately sampled tests, any of the functions may be used, although a function based on the Weibull distribution is shown to be most suitable. Where data are poor, all functional forms exhibit disadvantages, notably in terms of a reduction in precision of parameter estimation. This problem can be exacerbated by the complex numerical techniques required to estimate the parameters of a particular function. A set of ideal requirements for a performance function are presented.

Jan 1, 1986

By V. Srajer

Several optimization approaches used by the Canadian mining industry were used to investigate surface coal mining operations in Canada. A computer program was applied to optimize truck/shovel systems and for various analyses. The analyses of eight western Canadian coal mines included investigation of fleet imbalance, under- and over trucking, development of the criteria for automated truck control, comparison study of various haulers and loaders, haulage layout, driving conditions, machine performance, servicing and maintenance policies, etc. Each property investigated supplied background information for the study and cooperated in optimization studies. The truck/shovel system optimization study is demonstrated as a test case to outline procedures and step-by-step instruction for similar applications.

Jan 1, 1989

By Bobby J. Timmons

A few years ago during a Southeastern Geological Society field trip, a professor from a Florida university and a trip leader commented at one of the stops, "... and for those of you interested in the geology ...." More recently, in the spring of 1992, a geology faculty member at another southeastern university remarked, "Field geology is obsolete." An interviewee for a potential assistant position stated that he had never had a field class. This would-be assistant was to receive his Master's Degree in-Geology within a few days from yet another southeastern university. A regional illness? Certainly not, but symptomatic of the general health of Geology. Capitulation to Environmental Sciences seems to be in vogue with Geology becoming a four-letter word and the mention of mining is verboten regardless of the company. These trends permeate our profession while the demands for earth materials increase, often exponentially. Efficiency, a word lost on government, is routinely expected of business--our business--the industrial minerals geologist and producer. Our stage shrinks but the responsibilities expand. Are we meeting the challenge? Stay tuned!

Jan 1, 1995

By G. J. Olson, K. S. Noah

A combined physical/biochemical process for the precombustion removal of 13 inorganic hazardous air pollutant (HAP) precursors, i.e., Sb, As, Be, Cd, Cr, Cl, Co, F, Pb, Hg, Mn, Ni and Se, from coal was tested. Biochemical processes for removal of HAP precursors from coal potentially offer advantages of deeper cleaning, more specificity and less coal loss. The slurry column is a second-generation process for the beneficiation o f fine (60-mesh x µm coal by a combination of physical separation of mineral matter and biooxidation of pyrite. Sixty-seven percent removal of pyrite from a 60-mesh Pittsburgh #8 coal was achieved at a 35% (w/w) slurry concentration and a five-day reactor residence time. Ninety percent of the heating value of the feed coal was recovered. Among the HAP precursors of most concern, over half of the Se, As and Hg were removed from the feed coal. From 40% to 70% of most HAP precursors were removed from the feed coal. Hg in the feed coal was reduced from 0.12 to 0.054 ,µg/g in the product coal, while waste coal contained 0.24 µg/g.

Jan 1, 2000

By B. H. Johnston, B. N. Doyle

Introduction In 1983-84 Kidd Creek Mines Ltd. in Timmins, Ontario, started up the second commercial zinc pressure leaching plant based on the Sherritt Gordon Mines Ltd. direct pressure leach process. The first plant using this process was that of Cominco Ltd. at Trail, B.C. (Parker, 1981; Martin and Jankola, 1985). The Kidd Creek plant was constructed to take advantage of the extra capacity that the existing purification, cell house, and casting sections had in comparison to the roasters and acid plant. In addition, pressure leaching offered flexibility in other areas such as treatment of low grade concentrates. The process design package and the detailed design of the autoclave system was carried out by Sherritt. The detailed design for the rest of the plant was done by Cambrian Engineering Ltd. under the direction of Kidd Creek. The construction and commissioning of the plant has been described earlier (Johnston, 1983). This paper describes, briefly, the application of the direct pressure leaching process for zinc concentrates and the integration of the pressure leaching plant into the main zinc plant at Timmins. It discusses some of the startup problems and changes made to the plant. The damage following a fire inside the autoclave is also described. The subsequent repair and recommissioning of the autoclave has enabled Kidd Creek to demonstrate the originally sought advantages of pressure leaching and to take advantage of their extra cell house and casting capacity to produce more zinc. Process principles The development of the direct pressure sulphuric acid leach of zinc sulphide concentrate at temperatures above the melting point of sulphur and at elevated oxygen pressure has been described in detail in a number of papers (Martin and Jankola, 1985; Johnston, 1983; Veltman and Bolton, 1980; Kawulka et al., 1975; Veltman et al., 1977). The most significant aspect of the process is that sulphidic sulphur is converted to elemental sulphur in reactions that can be simply described as follows. ZnS + H2SO4 + 1/2 O2 ZnSO4 + S° + H2O (1) FeS + H2SO4 + 1/2 O2 FeSO4 + S° + H2O (2) The rates of these reactions increase with temperature. However, at temperatures above the melting point of sulphur, 12°C, the presence of a surface active agent is necessary to prevent the occlusion of the sulphide particles by elemental sulphur and the consequent termination of the leach. The zinc that is dissolved according to reaction (1), after purification of the zinc sulphate solution, is recovered by electrowinning. The iron that is put into solution according to reaction (2) is further oxidized to the ferric state and subsequently, as the acidity decreases, precipitates as hematite or as complex basic sulphates. MINERALS AND METALLURGICAL PROCESSING Examples of this are shown below. 2FeSO4 + H2SO4 +1/2O2 - Fe2(SO4)3 + H2O (3) Fe2(SO4)3 + 6H2O - Fe2O3.3H2O + 3H2SO4 (4) 3Fe2(SO4)3 + 14H2O - 2(H3O)Fe3(SO4)2(OH)6 + 5H2SO4 (5) (hydronium jarosite) The conventional zinc process involves concentrate roasting to produce zinc oxide and sulphur dioxide. The zinc oxide is leached in spent electrolyte and the sulphur dioxide is converted to sulphuric acid. The pressure leaching process can, therefore, be used to replace conventional concentrate roasting, sulphuric acid manufacture, and atmospheric leaching with one unit operation. Integration of pressure leaching into the Kidd Creek flowsheet A block flowsheet of the Kidd Creek zinc plant, with the new pressure leach plant integrated, is shown in Fig. 1. The original plant, which was constructed in 1972, is based on standard electrolytic zinc plant technology. Concentrates are dead roasted to remove sulphur. Sulphur dioxide is treated to produce sulphuric acid. Calcine is leached in two stages using cell-house electrolyte. Iron is precipitated as jarosite and solutions are purified to remove copper, cobalt, nickel, and cadmium to provide feed for a partially mechanized cell house. Zinc is plated from solution, and the resulting cathode is melted, cast, and sold (Johnston, 1983). The integration of the pressure leaching plant into the main zinc plant is relatively simple. As shown in Fig. 1, a fraction of the zinc concentrate and spent electrolyte are fed to the pressure leach autoclave, while part of the calcine from the roaster is used as a neutralizing agent in the oxidation step along with. the jarosite leach solution. The main exit stream from the pressure leaching plant is the slurry discharged from the oxidation step, consisting of neutral zinc sulphate solution, unreacted calcine, and precipitated iron oxides. This slurry is transferred to the neutral leach step in the main zinc plant. The leach residue from the pressure leaching step, consisting of basic iron sulphate and elemental sulphur, is washed and impounded along with the jarosite residue produced in the main plant.

Jan 1, 1987

By Yannis Faitakis, Brenda M. Wright, Gail Powley, Mark Hamblin, Randy B. Paine

A major problem in the ore extraction process is the high degree of variation in the mix of materials that make up the mined ore. The mix of materials makes it extremely difficult to characterize and predict the behavior of the ore when it is processed. The Quality Production Database (QPD) tool was successfully used in order to predict and transfer information regarding the quality characteristics of the ore being delivered to the extraction and hydrotransport unit operators. This paper describes the use of QPD in the prediction of various ore characteristics at the Syncrude Canada Ltd. processing facility.

Jan 1, 2000

By Gustavo G. Pilger, Jair C. Koppe, João Felipe C. L. Costa

Information for mine planning requires to be close spaced, if compared to the grid used for exploration and resource assessment. The additional samples collected during quasi-mining are frequently located in the same pattern of the original diamond drill holes net but closer spaced. This procedure is not the best in mathematical sense for selecting a location. The impact of an additional information to reduce the uncertainty about the parameter been modelled is not the same everywhere within the deposit. Some locations are more sensitive in reducing the local and global uncertainty than others. This study introduces a methodology to select additional sample locations based on stochastic simulation. The procedure takes into account data variability and their spatial location. Multiple equally probable models representing a geological attribute are generated via geostatistical simulation. These models share basically the same histogram and the same variogram obtained from the original dataset. At each block belonging to the model a value is obtained from the n simulations and their combination allows one to access local variability. Variability is measured using an uncertainty index proposed. This index was used to map zones of high variability. A value extracted from a given simulation is added to the original dataset from a zone identified as erratic in the previous maps. The process of adding samples and simulation is repeated and the benefit of the additional sample is evaluated. The contribution of an extra sample was evaluated locally, globally and in terms of variation in total measured resources. The procedure showed to be robust and theoretically sound, mapping zones where the additional information is most beneficial. A case study using coal seam thickness illustrates the method.

Jan 1, 2002

By M. J. Howes, C. A. Nixon

INTRODUCTION A safe heat stress control strategy for an underground mine has three elements: Application of an environmental measure which reflects physiological strain with sufficient accuracy for the range of conditions encountered underground. Acceptance of a functional relationship between the environ- mental measure and human performance which is used to optimise the environmental conditions achievable with either ventilation or ventilation and refrigeration. A management control strategy based on the environmental measure which is designed to ensure that work in environments where excessive physiological strain may occur is prevented and corrective action is initiated. The environmental measure that reflects physiological strain is the link between the three elements and, since the turn of the century, the discussion of the merits of various indices has been prolific. One problem in selecting a suitable measure or index is the ease with which it can be physically obtained relative to accurately reflecting the physiological strain. For example, wet bulb temperature is simple to measure and, for a particular mining sys- tem, it may adequately represent physiological strain, however, it would not necessarily provide the same relatively safe measure in a different mining system. The acceptance of a measure which can be universally applied has been confounded by both development and predisposition. That is not to say that there is only one "correct" measure and all others are unsuitable. It is self evident that if the application of a particular index has resulted in adequate control, then that mea- sure is correct for that situation. However, an understanding of the limitations is necessary to ensure that adequate control is maintained as mining conditions change. Almost 100 years after the question of heat stress in mines started to be dealt with in a collective manner, an analysis of the available information is leading towards a general strategy to control this problem. In the paper, the developments in heat stress assessment are briefly examined and followed since the earliest published observations on the effect of heat in mines (Haldane, 1905), efforts to determine a relationship between an environmental measure and human performance are reviewed and summarised and the benefits of control strategies such as acclimatisation and shortened shifts are discussed as they relate to Mount Isa Mines. The results of testing the prototype air cooling power instrument are discussed and a heat stress control strategy outlined. HEAT STRESS AND AIR COOLING POWER The operation of the human engine is analogous to other engines where the conversion of chemical energy from the oxidation of fuel to useful mechanical energy is not 100% efficient. In a diesel engine it is about 33% and in a human engine less than 20% resulting in at least five times as much heat produced by the meta- bolic process as useful work done. Metabolic energy production is related to the rate at which oxygen is consumed and is about 340 W for each litre of oxygen per minute. Using measured oxygen consumption and an average body surface area of 2.0 m2, the approximate metabolic energy production associated with different mining tasks is (Morrison et al. 1968):- • Rest, 50 W/m2 • Light work, 75 to 125 W/m2 (machine, LHD or drill jumbo operators) • Medium work, 125 to 175 W/m2 (airleg drilling, light construction work) • Hard work, 175 to 275 W/m2 (barring down, building bulkheads and timbering) • Very hard work, over 275 W/m2 (shovelling rock) Heat balance is achieved when the rate of producing heat (the metabolic heat production rate) is equal to the rate at which the body can reject heat mainly through radiation, convection and evaporation. Heat exchange between the lungs and the air in- haled and exhaled is normally less than 5% of the total and there- fore usually ignored. Any heat not rejected to the surroundings will cause an increase in body core temperature. Since heat stress is related to the balance between the body and the surrounding thermal environment, the main parameters required to be known when determining acceptable conditions are those associated with the heat production and transfer mechanisms. These can be summarised as follows: Metabolic heat production rates (M - W) Skin surface area (A3) (and effects of clothing) Dry bulb temperature (t[ ]) Radiant temperature (t[ ]) Air velocity (V) Air pressure (P) Air vapour pressure (e [ ]) The rate of heat transfer to or from the environment depends on the equilibrium skin temperature t, and the sweat rate S,. These in turn depend on the response of the body to the imposed heat stress and the effect of thermoregulation (Stewart, 1981). Thermoregulation The body contains temperature sensitive structures which send impulses to the brain at a rate depending on the temperature. Both hot and cold signals can be differentiated and the thermoregulatory response ahivated according to which signal pre- dominates. If "cold" signals are dominant, body heat loss is re-

Jan 1, 1997

By Krzysztof Filek, Anindya Sinha, Bernard Nowak

Taking into account the possibilities of condensation of water vapour contained in air during its flow through a series of spot coolers as well as formation of eventual fog, a mathematical model of dynamic temperature, specific humidity and degree of fogging of cooled air in underground mining galleries with direct ventilation is developed. In the gallery, there situated various heat and moisture sources both of distributive and of concentrated types. The model is represented by three partial differential equations of first order. The cases of cooling with full as well as with partial air stream flowing through the gallery are distinguished. Owing to the fact that in practical designing, more important is the knowledge of the variation process of temperature, specific humidity and fogging of cooled mine air at steady state conditions, the model is modified to a set of three nonlinear ordinary differential equations. Based on the presented algorithm, the obtained set of ordinary differential equations is numerically solved which enables to: • specify the distribution of temperature, specific humidity, relative humidity and degree of fogging of cooled air in mining galleries; • determine the cooling powers of air coolers for their known locations; • indicate the locations of the coolers of known cooling powers. Exemplary calculations of temperature and relative humidity of cooled air are presented graphically in a number of diagrams.

Jan 1, 1997

By H. Huang, J. C. Wu, C. H. Xie, Y. Q. Xue, Z. H. Zhang

Two- and three-dimensional groundwater flow models for dewatering in an opencast mine are presented. The models focus on solving special problems using numerical simulations of the opencast mine. These special problems include the drying up of the aquifer, the movement of the model boundary and the prediction of the residual head. These models are used to describe the decline of groumdwater levels in the Yuanbaoshan Opencast Coal Mine and its neighboring area during the drainage period. The simulated values agree very well with the field data. There have been several successful examples of numerical simulation for predicting the mine dewatering rate in an opencast mine. However, The hydrogeological conditions for the Yuanbaoshan Opencast Coal Mine are rather complex, and the total average flow rate is very high at 4.676 x 105 m3/d (1.65 x 107 cu ft/day). In this paper, a two-dimensional groundwater flow model in a large area is developed to solve the problems mentioned above by simulating regional drawdowns of the groundwater level during the drainage period, and a three-dimensional flow model is developed to simulate the residual head near the open-pit bench.

Jan 1, 1998

By Naj Aziz

Both bolt profile shape and spacing (rib spacing) have been found to influence the bonding capacity of the grouted rock bolt. The bolt surface profile configuration has greater importance to rock bolting in strata reinforcement in mining than the steel rebar used in civil engineering construction. This is because a rock bolt in mining usually is subjected to greater dynamic loading than the steel rebar in civil engineering construction. The increased bonding capacity of a bolt is important when supported ground is either heavily fractured, faulted or the supported ground is of soft formation, typically that of coal measure rocks. Past laboratory studies have identified the bolt profile spacing as of significant relevance to bolt resin rock bonding increase, however, no attempt has been made to determine the optimum spacing between the bolt profiles spacing. Accordingly, a series of laboratory tests were carried out on 22 mm core diameter bolts, commonly used in Australian mines, installed in cylindrical steel sleeve. The study was carried out using both push and pull testing methods. The push test was carried out in 150 mm long sleeves while the pull testing was conducted in 115 mm long sleeves. Profile spacing tested include, 12.5, 25.0, 37.5, and 50 mm lengths. Additional studies undertaken include modelling the profile of the load-displacement data of pull testing. Bolts with a profile spacing of 37.5 mm were found to provide optimum load bearing capacity as compared to other tested profile spacings.

Jan 1, 2009

By Sidney A. Hiscock

INTRODUCTION Lead and zinc are today rightly regarded as sister metals. Historically, however, they differ markedly with lead being known and in general use as the metal since 3000 BC in several countries which at that time might have been described as mature economies. Zinc was not isolated and recognised as an industrial metal with distinct and valuable properties until about 1700-1800 AD (although it had, of course, been utilised as a constituent of brass for some 2000 years previously). c were well established as industrial metals by the beginning of the 20th century. The present paper briefly reviews the overall growth in world consumption* and changes in the three main areas which today represent the mature economies or industrialised areas, i. e. ,Europe, the United States and Japan. Trends in overall consumption are determined by the demands for the individual uses for lead and zinc. The changes in patterns of use in the industrialised areas - individually and collectively - and reasons for the changes are also considered. LONG TERM TRENDS - 1900-1984 The World Picture Since the beginning of the 20th century the world consumption of lead (3.9 million tonnes in 1984) has grown almost fourfold and that of zinc (4.7 million tonnes in 1984) is almost eleven times greater than it was in 1900. The long-term trend has been one of continuing expansion for both. metals although the pattern has been severely disturbed at certain periods, for example in the years immediately following the two world wars and at times of marked recessions in industrialised areas. Apart from such major setbacks, new levels of consumption have been established regularly every few years. Recently, however, consumption has not really grown and with only a slow and partial recovery in world industrial activity at present, there is no indication that the peak consumptions of 4.2 million tonnes of lead, set in 1979, and 4.8 million tonnes of zinc (1973) will be exceeded in the near future. The growth in world lead and zinc consumption since 1900 is shown in Figure 1 (which includes changes in copper and aluminium for comparison). Overall growth over the period 1900-1984 has been about two percent a year for lead and about three percent for zinc (annual growth for copper has also been about three percent). Increases in tonnage consumption and growth rates by decade for lead and zinc are summarised in Table I. However, the growth rates by decade sometimes conceal very large individual annual increases and decreases. For example, in 1975 lead consumption fell by 13 percent and zinc consumption by 22 percent compared to 1974. In 1976 consumption of lead and zinc rose by 12 percent and 18 percent respectively. The years when consumption 'landmarks' (ie one, two, three and four million tonnes) were first reached are shown on Table 2. Clearly, the pattern of shorter periods being required to attain each extra million tonnes of consumption was broken in the 1970's. The consumption of zinc overtook that of lead for the first time in 1940, and since 1946 has always been at a higher level. Trends in industrialised areas The two key industrialised areas at the beginning of the 20th century were Europe and the United States which between them accounted for some 95 percent of world lead consumption and 98 percent of zinc consumption. Figures 2 and 3 show the growth in lead and zinc consumption, and Figures 4 and 5 the percentage shares for each area since 1900. Europe has usually been the major user of lead except for some years in the 1920fs, most of the 40's and early 50's when consumption in the United States was larger. In 1900, Europe accounted for about 64 percent of world lead consumption and the United States just over 30 percent. Currently, Europe takes about 40 percent of world consumption and the

Jan 1, 1986

After World War I the character of my mining life changed completely. Before then adventure had been almost continuous but after the war there was lots of staying in one place, almost routine, between adventures. This was fortunate, as the life of a mining engineer is a wandering one at best, and it makes being married and raising a family a dubious enterprise. It would have been quite out of the question with the life I led before I got married at the end of the war The new sort of life made the "beaconing of adventure" intermittent. There were often long stays at home between wanderings which was fortunate as it changed the continuity of these memories. I must briefly summarize accounts of the more or less routine periods and concentrate on the occasional exciting episodes between. After a long visit with friends in Berkeley, Dorothy and I left for Arizona and three days at the Grand Canyon that were nearly too much for both of us. One beautiful day we started a short walk down the Bright Angel Trail. It was such a lovely bracing day that we kept on walking and enjoying the views until, just after noon, we reached the Colorado River. We were the only ones in sight so to enjoy the sunshine to the full we both had a good swim in the swiftly flowing river. So far, everything was wonderful. But when we started to climb back up the trail, rising nearly 6000 ft in ten miles we soon found that we had counted too much on our youth and endurance. We had to rest every few hundred feet and it seemed as though we would never reach the top. It was eight o'clock in the evening when we finally staggered into our room and collapsed. On the way to the Canyon I had seen John Greenway and Dr. Ricketts in Santa Barbara and they had little idea what would happen to them or to me. Before resuming his position as general manager of Calumet and Arizona and New Cornelia, Greenway must recover from the effects of the gas he had suffered in the war. Dr. Ricketts

Jan 1, 1976

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By G. Gillet

The conventional magnetic separators used in the mineral industry are equipped with traditional magnetic circuits or with permanent magnets (rare earth). However, when processing minerals with low magnetic susceptibilities or minerals of very small particle size, these separators are inefficient. In addition, a conventional magnetic circuit consumes considerable amounts of electricity to generate a relatively low magnetic field (~2 Tesla). The equipment is also heavy and bulky. Only a separator having a superconducting sole-noid can contribute in solving most of the problems, and only such a separator can produce high magnetic fields with low energy consumption. New developments and the experience acquired in using these superconducting magnets have resulted in new uses for magnetic separation, especially in the environmental area. This paper presents and describes a separation system that uses superconducting magnets that are completely automated and able to treat raw materials (concentration or purification) or liquid industrial effluents containing heavy metals in solution. Several industrial sectors, including the surface treatment, petroleum and chemical industries, are generating effluents containing heavy metal ions. The main metal ions found in such effluents include zinc, copper and vanadium. The purification of such effluents by magnetic separation is performed by adding a certain proportion of iron salts. The treated effluent passes through a superconducting separator where the flocs are retained in a matrix of steel wires. The efficiency of supraconducting high-gradient magnetic separation (HGMS) by this method exceeds 95%. Mineral separation and industrial-waste purification results are presented.

Jan 1, 2000

By J. A. Finch, G. Labonté

The behavior of four different redox sensing electrodes (gold spiral, black platinum, chalcopyrite, and galena) during roughing and cleaning flotation is examined and related to the flotation response. The results indicate that the solution potential (as sensed by the black platinum electrode) is different from that of the pulp potential (gold spiral) and is only weakly related to the flotation response. On the other hand, the potentials exhibited by the gold spiral and the two mineral electrodes are related to the flotation response. The transient responses of the gold and mineral electrodes during the addition of flotation modifiers could indicate the conditioning time required to achieve flotation or depression of selected minerals.

Jan 1, 1991

By I. S. Parrish

Gold deposits in Hebei and Shanxi provinces in northeast China fall into three general classes: - Placer deposits - Mesothermal vein deposits - Disseminated and oxidized volcanogenic deposits Past production dating back thousands of years has all but exhausted the placers. The mesothermal vein deposits, incorrectly considered in the past as "greenstone" type, are small to moderate in both size and grade. Production from these granulite and amphibolite hosted veins, at present greater than from the newly discovered disseminated deposits, is expected to wane. Production from the disseminated deposits achieved only in the past few years is expected to equal or surpass that from the other classes of deposits. At least one "world-class" deposit is being developed.

Jan 1, 1989

By C. M. Brown

Common operating philosophies and objectives of mining and mineral process operations are discussed. Key information required to effectively monitor operations is reviewed in conjunction with the impact of external variables such as commodity price changes on strategies adopted.

Jan 1, 1991