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By S. K. Mukherjee, A. W. Laurito, M. M. Singh, N. I. Jayaraman

This paper describes the application of foam to control respirable dust in a double-drum longwall mining section. The feasibility of injecting an adequate quantity of foam through the shearer drum was initially established in the laboratory. Thereafter, underground testing of the foam system to determine its effectiveness on an active operating longwall was conducted in a mine in West Virginia. The foam system proved superior to the water spray system. The details of the laboratory and field investigations, the results obtained, and the degree of acceptance of this system by the mining company are discussed in this paper.

Jan 1, 1987

By K. Norrena, E. Magri

An important problem in mine operations is the classification of material as waste, low-grade stockpile and ore. This classification must often be made with blasthole data that are widely spaced and that have sampling errors. Geostatistical-simulation techniques combined with basic economic principles allow a procedure for classification that maximizes the expected profit. Geostatistical-simulation methods (Gaussian, indicator or annealing-based) allow the integration of hard and soft data in the creation of alternative, equally probable realizations of the mineral grades. At each location, for each realization one calculates the "profit" if the block were to be classified as ore or waste. The optimal classification is the one that maximizes expected profit (a maximum profit selec¬tion or MPS procedure). In this paper, the authors discuss the theoretical justification of the method and implementation details. The use of blasthole data from different types of mineral deposits, that is, different levels of continuity, are considered. The authors also show the efficacy of the procedure with different levels of sampling error. The increased revenue due to the MPS procedure and improved sampling is shown. The paper shows the geostatistical-simulation procedure and the uncertainty in block grades that result from incomplete and imperfect sampling. The optimal classification is presented. Optimal block classifications are transferred to realizable dig limits by hand-drawing polygonal boundaries. The results of the proposed method are compared to classification based on kriging. Limitations and areas of future work are identified.

Jan 1, 2000

By Tugba Kami?oglu

Hydromagnesite (Mg5(CO3).(OH)2.4H20), one of the sedimentary deposits in Salda Lake Basin, is developed under the control of active tectonics regime. Since the occurrence and deposition mechanism of hydromagnesite are unstable under atmospheric conditions, in order to evaluate the precipitation process, systematic host rock, lake water (surface, 20m, 40m, 60m) and hydromagnesite sampling were carried out. Macro, micro, XRD and chemical investigations were done in detail. Not only the results and formation and crystallization mechanism but also hydromagnesite's St values were evaluated. Main resources could be ultramaficmafic surrounding rocks for hydromagnesite formations. However, hydromagnesite development is considered in different phases, because it has showed various characteristics in Kayadibi and Köpekcayin areas. In Kayadibi area Huntide-Dolomite-Aragonite-Hydromagnesite paragenesis has been realized. Hydromagnesite has also formed in Köpekcayiri area, large amount of hyrdomagnesite occurrences and levels various are noticeable. Actual position of hydromagresite levels exhibits tectonic influences in and around the lake. Therefore, the levels seems to exhibit secondary formation properties. In addition, the formations are being still maintained sedimentation as products of a peripheral deposition phases. Resolving and redifferentiation sedimentation processes probably improve in periods of lake water chemistry especially Ca/Mg proportional changes, and also SI variations indicate the mentioned precipitation mechanism. Similarly, side and into the lake hydromagnesite is stored different levels and ultramafics-mafics rocks fallow indicated stages. It is postulated that hydromagnesite formation, which is existed by result of determined developments and multiphase sedimentation, could be originated from probable crack zones on the lake bottom.

Jan 1, 2005

By Donald A. Dahlstrom, Wai-Ping Kam

The mineral industry and individual processing plants are large consumers of energy. While energy conservation has received some attention in the present Federal Administration National Energy Strategy proposal it is obvious that the mineral industry should also continue to emphasize energy consumption reduction for important economic savings alone. As indicated by other studies, comminution is the largest single consumer of energy in the mineral processing plant (1.3% of total electrical energy consumption in the U.S.A.).1 Fine grinding is the largest portion of this amazing statistic. While substantial energy savings have been demonstrated by new machine developments and improved control methods, undoubtedly there are additional areas in comminution where substantial progress can be made. Some time ago, the U.S. Department of Energy published a report on "Comminution and Energy Consumption" by a committee composed of experts in the field of comminution from industry and academia. They identified the following potential for energy savings by specific areas? Classification device design 9-13% Comminution device design 3-16% Control 9% Grinding additive 3-6% Materials 3% It is apparent that the classification device had the first or second highest savings potential. Obviously, serious research consideration must be given to this technology area. A series of three papers have previously been written on the subject of two stage classification research work for ball mill operation by the authors. Accordingly, a brief review will be made. Previous Research Results on Two Stage Classification The hydrocyclone was selected as the type of classification equipment for this study as it represents by far the majority of the devices used in ball milling. While four different hydrocyclone circuits have been proposed by others,3 only two were considered in this work as illustrated in Figure 1. Flowsheet A is truly counter-current in nature in that the product from the ball milling circuit is taken from the first stage overflow

Jan 1, 1992

By J. R. Woolsey, B. G. Noakes

A prototype remote placer drill (RPD) has been designed and successfully tested in the offshore gold district of Nome, AK. The RPD system is unique in its design and in the cooperative support that was provided during its development by industry, academia and government. The RPD was conceived as a low-cost, rapid drill sampling system for the reconnaissance of marine precious metal placer deposits. The drill functions as a counter- f lush system. It incorporates a dual wall casing and a rotating percussion bit driven by way of a drill rod through the central eductor casing by a heavy duty pneumatic rock drill. The cost of drill sampling in the exploration of marine precious metal placers has traditionally been high due to the technical difficulties imposed by the environment and the remoteness of the sampling target. The objective was to reduce capital and operating costs of a drilling program through the development of a portable system incorporating drill compoments having a high power-weight ratio that could be deployed from work vessels of opportunity. A principal design requirement was to operate the system remotely from the sea floor with only flexible connections to the surface. This includes cable for deployment and retrieval and hoses for air supply and sample recovery, thus making the drill system semi-independent of the vessel's sea-induced motion. The prototype RPD underwent full operational test off Nome, AK in late summer 1988. It was deployed from the landing door of a modified LCT vessel, and rigged for a single pass of 5 m (16.5 ft). More than 50 holes were drilled in a wide variety of bottom materials typical of the off shore Nome submerged beach placers. The prototype RPD proved to be an overall success, recovering representative samples throughout the section in all but the heavy boulder ground. Compared with a conventional drill system on board that was assigned to a production drilling system, the RPD was capable of reliably sampling moderate ground less than 6 m (20 ft) in thickness, using a smaller crew and operating in higher sea conditions.

Jan 1, 1991

By Terry Ramsey, N. Finkelstein

INTRODUCTION The Cyprus Pima Mine is located 20 miles southwest of Tucson, Arizona and about fifty miles north of the Mexican border. The Cyprus Pima orebody was discovered in 1950. Feasibility studies were begun in 1952 employing underground development. Economic studies showed open pit operations to be feasible and stripping of overburden was begun in late 1955. The first ore was reached in October of 1956. A 3,000 tpd (tons per day) concentrator was completed and began operation January 1, 1957. From these beginnings, Cyprus Pima Mining Company has expanded into a plant processing approximately 54,000 to 56,000 tpd. Four expansions were required to bring the plant up to its present capacity. Each expansion incorporated what had been learned in previous operation of the plant. The first expansion was completed in August of 1963 doubling design capacity to 6,000 tpd. In reality, because of softer than average ore, capacity was closer to 8,000 tpd. Expansion I1 was completed in July 1966 bringing design capacity to 18,000 tpd. Expansion I11 followed immediately. Completed in the summer of 1967, this expansion increased the design capacity to 30,000 tpd. The actual dry tonnage of this portion of the operation in 1974 averaged 38,400 tpd, again, primarily as the result of treating softer than average ore. The latest expansion (Expansion IV), completed in December 1971 and utilizing a semi-autogenous grinding circuit, is located in a separate building (Mill 11) approximately 200 yards southeast of the original mill (Mill I). Expansion IV was designed for 14,000 tpd, but after circuit modifications, averaged 16,500 tpd in 1974.

Jan 1, 1976

By A. W. Khair

This paper presents an analysis of surface subsidence characteristics in room and pillar mining using physical models and laser holographic interferometry (holometry). The analysis included the effect of various geometric parameters and different overburden materials (sand and gelatin) and resulted in the formulation of a more realistic model material for laboratory simulation of typical geologic overburden. Previous work using this technique involved the use of sand to model over- burden and detailed results are presented elsewhere (1). The main emphasis of this work was to develop a gelatin based mixture which would provide a better model. Initial work during the development of this model included testing various ingredients and monitoring the time dependent characteristics of the model, such as the surface penetration resistance. This value indicated changing deformability/ridgidity of the model with time. Once several models had been developed, holometry techniques were applied. Initial tests showed that the stiffness of gelatin created a bridging effect over the mine models, thus restricting deformation. At this point, a 0.5 inch (1.27 cm.) layer of sand was placed directly over the mine model with the gelatin mixture on top. The approach was to model different zones of overburden movement that exist as an extraction occurs (2). The sand would represent the caved and fractured zones which occur above the excavated area. The gelatin models the continuously deformed zone which occurs between the fractured zone and the surface. Holometry techniques applied to this model yielded successful results.

Jan 1, 1986

By P. J. D. Lloyd

INTRODUCTION The mining of gold commenced on the Witwatersrand in 1884, and reached a mature state by about 1910, when nearly 20 million metric tons were mined. Mining has continued on the central Witwatersrand, the area of the first mining, at a similar rate until recent years. As a result, well over 1 billion tons of gold mining wastes have been dumped along a line over 80 km long, centred on Johannesburg and stretching from Randfontein in the west to Nigel in the east. (see Figure 1). The population of this conurbation, which straddles the line of mining, has grown to about 5 million. In 1922 uranium was first discovered as a constituent of the ore (and thus of the wastes). Investigations in the 1940's revealed that values near Johannesburg were low, but that, at the western and eastern ends of the central Witwatersrand line, and in the new areas centred on the towns of Carletonville, Klerksdorp and Welkom up to 200 km from Johannesburg, values were adequate to permit uranium recovery. Commercial production commenced in 1952, and by 1979 over 100 000 tons of uranium oxide had been produced. There thus arises on the central Witwatersrand nearly 1,3 billion tons of slimes containing an average of about 0,005% uranium and about 1,8% FeS2. The material grading is about 70%-75µm, and the primary mineral is silica (over 85%). The material is contained in dams constructed hydraulically. A typical dam has an area of about 100 hectares and there are over 6000 ha under slimes dams in the area of concern. The treatment of these dams for uranium recovery commenced in 1977. The situation has thus arisen whereby potential uranium ores have unwittingly been dumped in significant quantities within a major metropolis over a period of nearly a century. The purpose of this note is to review the experience arising from the potential hazard over a considerable time.

Jan 1, 1980

By J. P. Mogan

As diesel exhaust contains a mixture of several materials of varying degrees of toxicity which interact to a currently undefined extent in their effects on health, the prescription of a quantity of ventilating air to dilute the exhaust to a safe level is necessarily, to some extent, arbitrary. Some jurisdictions rely on the application of a safety factor to a single "governing" component (usually nitric oxide for most modern diesel engines). Others specify that the total of the "adjusted" concentrations (actual concentration divided by maximum permissible concentration) not exceed "one" for related constituents (irritant gases for example). Neither of these approaches, however, takes account of the interaction of the gaseous constituents with diesel soot (although many studies verify such interactions), nor do they provide any economic incentive for reducing pollutants which are not included in the criteria for ventilation. Because of these factors, the Canadian Department of Energy, Mines and Resources awarded a contract for a study of the current literature bearing on the "Health Implications of Exposure of Underground Mine Workers to Diesel Exhaust Emissions" (1). One goal of this study was the formulation of a new ventilation criterion. The contractors, Ian W. French and Associates, studied more than 1500 literature citations which they interpreted in a 350 page report and summarized in a companion 1600 page annotated bibliography. In the recommendations section of the report, they defined a new 'health effects index' or ventilation criterion "V" as follows: [ ] The contractors consider that a "V" between 3.0 and 4.0 poses a moderate threat to health which could be alleviated by the use of protective equipment such-as filter helmets. If V exceeds 4.0, an increase in ventilation quantity to reduce the pollutant levels is indicated. In the case of mines with a quartz content in excess of 20% in the ore, the contractor recommends that a term. [Quartz]/ [Threshold Limit Value Quartz] be added to the equation, and [2.0 = ?v? = 30] be the range justifying protective equipment, with added ventilation if "v" exceeds 3. A degree of pragmatism was inherent in the development of the equation, as the contractors recognized that the components of the index, if it is to be useful in the near term, should be within current routine underground environmental assay capability.

Jan 1, 1980

By Huayang Luo, Hui Li, Wei Zhou, Gang Chen, Scott Huang

Abundant beach and offshore resources of gold deposits are demonstrated along the coast of Alaska, and Nome is one of the richest places. Considerable amounts of mineral resources data were collected and stored in geodatabase of ArcGIS. In searching for the best model for the interpretation of the Nome offshore gold placer deposit and accurate estimation of the reserve, the data were processed and visualized using ArcGIS. The gold concentration data collected from the Nome offshore deposit are of characteristics such that: (1) near lognormal distribution of data, (2) high positive skewed distribution, (3) a small number of extreme values dominating the reserve estimation, and (4) apparent anisotropy of the variogram. These characteristics make the development of an interpolation model for the gold deposit a challenging task. In this study, it is determined that five geostatistical methods provided by ArcGIS with visualization capability, i.e., Inverse Distance Weighted (IDW), Ordinary Kriging (OK), Ordinary Kriging with lognormal transformation (OK-log), Simple Kriging (SK) and Indicator Kriging (IK) are suitable candidates for the development of an interpolation model. The five algorithms are evaluated and compared with each other in the study. Cross-validation technique is employed to examine the accuracy of the five approaches. It is determined that the OK and SK algorithms are not suitable for this dataset because of the difficulty of fitting the semivariogram model despite their lower RMSE (Root-mean-square error) and the higher RMSS (Root-mean-square Standardized). The OK-log algorithm does not provide satisfactory prediction either, because it places too much weight on the role of extreme values resulting in overestimation and extremely high RMSE. Reserve estimation of each blocks is conducted using each of the interpolation methods and compared with the reserve calculated using the conventional polygon method. The IWD and IK algorithm seem to provide estimations more agreeable with the conventional polygon method.

Jan 1, 2005

A chartered bus with about 30 attendees of the just completed 1998 SME Annual Meeting headed toward Green Cove Springs, FL, which the author had been told was located somewhere south of Jacksonville, FL. In conjunction with the meeting in Orlando, RGC (USA) Mineral Sands provided a tour of its facility. The author had worked with heavy minerals on a laboratory scale, performing the usual gravity and magnetic separations, but had never been to a heavy mineral mine and processing facility. Years ago as a metallurgist, the author once applied for a job in a heavy mineral operation in Florida -it may have been the same operation. After turning north on to Highway 95, the endless billboards promoting beach-front retirement communities proved the Atlantic ocean was just a few miles away (although the author never caught a glimpse of it). After another hour, the bus suddenly turned west and continued inland for about 65 km (40 miles). The thought occurred, "We are talking about beach sands, what are we doing heading away from the ocean?" As the bus continued inland, the author was struck by the flatness of the land and by the lakes, rivers and wetlands. He was also impressed by the many fine stands of pine and hardwoods. The office of RGC (USA) Mineral Sands is located about 16 km (10 miles) south of Green Cove Springs. Upon leaving the bus, members of the tour were greeted by Graeme Sloan, general manager, and Elliot Mallard, manager of mining. They provided some general information about the operation, including a slide show. RGC's Florida operation produces four concentrates: zircon (ZrSiO,), rutile (TiO,), ilmenite (FeTiO,) and leucoxene (an alteration product of ilmenite containing 67% to 70% TiO,). With a work force of about 170, the annual production of the above products is about 23,27, 45 and 14 kt (25,000 30,000 50,000 and 15,000 st), respectively. RGC also has other mineral sand operations, in addition to gold and base metal mines, mostly in Australia.

Jan 1, 1998

By Drew F. Holbrook

Most of the siliceous rocks of the Arkansas Novaculite are novaculite, but they also include some chert, siliceous shale and, rarely, sandstone. Novaculite is defined as a homogeneous mostly white or light-colored rock, translucent on thin edges, with a dull to waxy Juster, and comprised almost entirely of microcrystalline quartz. The most plausible theory of origin for novaculite is that primary silica from submarine volcanism was picked up by organisms and deposited as amorphous silica, which was subsequently converted to microcrystalline quartz (novaculite) through diagenesis. The Arkansas Novaculite crops out over an east-west distance of 200 miles in the Ouachita Mountains of Arkansas and Oklahoma. Three divisions of the formation have been defined: a Lower Division, almost entirely massive light-colored novaculite; a Middle Division of interbedded dark chert and shale; and an Upper Division of massive calcareous novaculite. The Lower Division and most of the Middle Division are Devonian in age, and the remainder of the Middle Division and all of the Upper Division are Mississippian. The massive novaculites of the Upper and Lower Division attain their maximum development along the southern and central outcrop belts of the formation in Arkansas, and it is these rocks that are the source of current silica production. The calcareous novaculite of the Upper Division weathers readily to a light-colored, porous, less brittle rock, and in places to tripoli deposits. Currently tripoli is mined from a single deposit near. Hot Springs, Arkansas, and the rock is processed for highly specialized filler and abrasive applications. In the vicinity of Hot Springs, Arkansas, certain novaculite layers from the Lower Division with unique textural characteristics are quarried and processed into whetstones. Several small quarries produce this rock for two local and one out-of-state whetstone-fabricating plants. Chemical analyses of composite samples of novaculite from both the Upper and Lower Divisions taken at widely separated localities average 99 percent SiOz indicating a substantial resource of high-silica rock.

Jan 1, 1978

By Stephen G. Hardcastle

CANMET has developed a simple yet very effective tracer gas method to test the efficiency of stench gas emergency warning systems which can easily be performed by mine personnel. This test method uses sulphur hexafluoride (SF,) as a surrogate for such stench gases as ethyl mercaptan. The test has many advantages: it is non-disruptive to production because SF, is odourless, colourless; it is safe to use as it is inert and only a small volume, insufficient to reduce oxygen levels, is necessary to dope the air with a measurable concentration; and most importantly it provides definitive results to assess the adequacy of the warning system. Specific findings of these tests have included, over doping sections of the mine and subsequent production loss due to sickness, the warning being ignored, the injection period being too short, excessive dispersion, and inadequate notification times. The simple test protocol involves releasing a known volume of the SF, under similar circumstances to those of the stench gas, and then sampling for its presence downstream. These samples are collected in air-tight syringes which are then analyzed in a laboratory with a gas chromatograph. By direct proportion, between the size of the stench and tracer gas releases, it is possible to convert the SF, concentration into equivalent stench gas concentrations profiles. From these, it is possible to define, the initial arrival time of the gas, the time it takes to reach a "target" notification concentration, whether the gas peak will exceed TWAEV, STEV and CEV exposure limits, and the duration of the peak above these levels. Since 1993, this test protocol has been employed a total of eight times on four different Canadian mine properties and the results are being used to optimize the stench gas release system with respect to volume of gas used, duration of injection, and number of injection locations.

Jan 1, 1997

By S. G. Mougiakakou

A computerized system (SCAPEVIEWER) based on landscape indices and artificial intelligence techniques is proposed in order to classify landscape images into various categories according to the quality of the pictured scenery. Photographs (totally 108) from various landscapes have been clustered, from seven experts in accordance to the mean estimation of landscape quality, into distinctive (C1), typical or common (C2), and indistinctive (C3). For each of the landscape photographs, ten (10) environmental and operational indices have been selected and determined, corresponding to visibility (V), naturalness (N), relief (R), flora quantity (FC), flora quality (F), water (W), skyline (SL), visual entities (G), season (S), and viewer position (P). These landscape indices are fed to a classifier. The classifier is a feed forward Neural Network (NN) trained by a novel hybrid method combining genetic algorithms and the back-propagation algorithm with adaptive learning rate and momentum. A total classification rate of 93.5% has been achieved, using only 8 input neurons. The discarded input neurons correspond to the indices S, and P. The pilot system SCAPEVIEVER has shown the feasibility of landscape photograph classification into three categories (C1, C2, and C3) according to the landscape quality, selecting simultaneously the most robust indices, by means of a computerized system combining the knowledge of an expert with a NN based classifier.

Jan 1, 2003

By R. Norman Holme

Sulfonate promoters and collectors were originally developed as relatively inexpensive anionic collectors to selectively concentrate the iron-bearing minerals in low grade iron ores and to remove iron and other mineral contaminants by flotation in the processing of glass sands, feldspar and other ceramic raw materials. The first reagents produced were effective collectors for such iron oxide minerals as goethite, hematite, limonite, magnetite and siderite. Further development has resulted in additional reagents with a broadened application to the selective flotation of a variety of nonsulfide and nonmetallic minerals. The products in this group are anionic sulfonate-type promoters, natural petroleum sulfonates, and synthetic modified sulfonates. Beyond their applications to iron-bearing minerals, the sulfonate collectors have gained commercial use in the flotation treatment of barite, celestite, fluorite, apatite, chromite, industrial garnet, kyanite, mica, rhodochrosite and cassiterite. in addition, extensive laboratory studies indicate their potential suitability for flotation of the following minerals: Anatase Anhydrite Bauxite Beryl Calcite Chlorite Some Copper Carbonates Corundum Dolomite Hornblende Huebnerite Ilmenite Lepidolite Magnesite Phosphate Minerals Rutile Scheelite Spodumene Sillimanite Talc Tourmaline Wolframite Zircon Sulfonate promoters are water soluble or dispersible liquids, syrups or pastes of varying viscosity with colors ranging from dark blackish-brown to clear. Customarily, they are fed to the conditioner or flotation as 5% to 30% water solutions or dispersions. In addition to their normal collector activity, certain of these products display notable frothing characteristics. Improved collector efficiency and adequate froth control can frequently be obtained by employing mixtures of sulfonate reagents, of ten with results quite markedly superior to those achieved with equivalent dosages of the individual promoter alone. Reagent requirements will vary with the type of mineral being treated; however, a general overall range of 0.13 to 2.0 kg/metric ton (0.25 to 4.0 lb/ton) of feed should cover most applications.

Jan 1, 1986

By Robert A. Jankowski, Charles A. Babbitt, Natesa I. Jayaraman

Today, approximately 31% of underground coal mining operations have designated occupations or areas with more stringent (below 2.0 mg/m3) respirable dust standards in place. Many U.S. coal mines have been able to maintain compliance with very low dust levels, and the Bureau of Mines, in an effort to identify those factors contributing to low dust levels, completed a survey of twelve continuous miner sections which were able to maintain their dust levels at or below 0.5 mg/m3. The quantity and quality of face ventilation appeared to be the most important factor in controlling dust exposure. The results directed Bureau research towards identifying improved face ventilation techniques for reducing quartz dust levels on continuous mining operations. This research has involved the evaluation of a half-curtain face ventilation technique which reduces the cross-sectional area of the entry, thus increasing the air velocity over the operator. The Bureau has also conducted extensive laboratory studies to compare tubing versus brattice, including an evaluation of both systems in a crosscut turn. In an effort to improve ventilation and reduce dust exposures of roof bolter operators working downwind of the continuous mining machine, the Bureau has evaluated two novel face ventilation systems. The first system consists of a fan discharge tubing technique which routes dusty air away from the return entry being bolted. In a technique similar to the fan/tubing concept, exhaust air from the upwind continuous miner face can be scrubbed using a dust collector to capture a portion of the dusty air, clean it, and supply it to the bolter operators as a split of clean intake air. A better understanding of face ventilation techniques should lead to solutions to the problem of more stringent dust standards due to quartz.

Jan 1, 1989

Stope width was controlled around 3.5 feet for the narrow high grade vein system at Hollister Gold Mine. Without top access, backfill material was pumped from bottom access up into stope through a series of pipes after ore had been slushed out in each mining cycle. Backfill material was changed from one layer of low strength concrete plus one layer of high strength concrete cap to one layer of low strength cellular concrete plus one layer of high strength concrete cap. Backfill cost was cut down to 30 percent during this process. To reduce the steel concrete pipe set up and tear down time which was about 2 to 3 shifts in each mining cycle, steel pipes were left in man-way between mining cycles. Cellular concrete was firstly introduced into backfilling stopes at Hollister mine in Nevada Mining. This process including concrete pump selection, pumping pipe selection, material mixing, material mechanics and quality control were fully covered in this paper. CUT AND FILL MINING CYCLE The most flexible mining method, cut and fill (Figure 1), was used to mine the high grade, narrow steep vein system on this Northern Nevada property which is located eighty miles north east of Winnemucca. Without top access, stopes were accessed through two man-ways on each end, in which people travel ladders and material transport slides were set up (Figure 2). Each mining cycle was called a cut, from 6 to 8 feet high. Overhand sequence was used. Mining cycle started with drill and blast the middle portion of the stope, then expanded towards both ends. Broken ore was slushed into mill hole located at middle of the stope and then mucked out through bottom access into main haulage. Stope length was about 300 feet, which was restricted by slushing distance. Timber man-ways and mill hole would be raised and established for backfill and ready for the next mining cycle.

Jan 1, 2011

By Brian Flintoff

SCOTT COULTER: Our second speaker today is Brian Flintoff. Brian graduated from the University of Alberta in 1973 with a master's of science degree and worked as a project metallurgist at Brenda Mines until 1977. He returned to the University of Alberta in late 1977 to work on his Ph.D. and later joined the faculty of engineering in 1980, leaving in 1988 to rejoin Brenda in the newly formed process technology division. In 1998 he left Svedala-Brenda to take a position with Simons International's newly formed enterprise development services group. His topic today is "People: Reduce, Reuse, and Recycle." INTRODUCTION BRIAN FLINTOFF: Historically, the mining industry has relied on the people within a corporation to get things done. The culture of most companies was one of a slow but sure development of people, processes, equipment, and technology. Many of the larger companies supported in-house engineering and research organizations. However, the metal market pressures of the 1980s and the emergence of the global economy in the late '80s and early '90s put enormous financial pressures on the industry. As the old adage says, "necessity is the mother of invention," and the mining industry, like many others, has undergone a transformation. A cornerstone element of the transformation was the downsizing (or right-sizing) that was calculated to "trim the fat" from the capital and operating budgets. The goal was and is to find the highest sustainable operating profit.' While I'm sure one can find examples of losers (e.g., those who can be classified as suffering from corporate anorexia or institutional Alzheimer's) in such business reengineering efforts, there is little doubt that, on the whole, the transformation has been successful. The reality that now faces mill management is "How to do more with less?" At the risk of oversimplifying, one can look at this on two levels:

Jan 1, 1998

By W. H. Andrews

DEFINITION OF BAUXITE The term "bauxite" (1) dates back to 1821, when it was assigned to a deposit of red earthy material outcropping near the village of Les Baux in Southern France. Chemical analysis indicated the presence of aluminium oxide, iron oxide and water. Deposits of bauxite are widespread throughout the world although the major developments occur within 30 degrees of the Equator. Essentially bauxite is the product of weathering and leaching of a variety of sediments or rocks, generally in a high-rainfall tropical or sub-tropical climate, resulting in the enrichment of the aluminium oxide content. It is characterised by the presence of one or more of the minerals, gibbsite, boehmite, and diaspore. The other constituents are usually clay, (typically kaolin), hematite and/or goethite (iron-oxides), anatase (titanium oxide), and possibly silica as quartz. A variety of minor elements may be present, such as calcium, magnesium, phosphorus, vanadium, gallium, manganese, and many others in very small quantities. Organic compounds are normally present from development in association with a soil horizon where vegetation may be encouraged by the high rainfall necessary for bauxitisition. The physical form depends upon the geological history of the deposit and correlation also exists with the mineralogy. Older deposits which have been subject to burial are invariably hard and compact with the alumina predominantly occurring as boehmite or diaspore, while the more recent, surface deposits are frequently less compacted and gibbsite is the dominant alumina mineral. In these later deposits, the physical form can vary from the fine red soil (terra rosa) of Jamaica to the Weipa pisolites to the cemented Arkansas type. The presence of concretionary structures is an important aid in identification in the field.

Jan 1, 1984

By S. Bhagwat

Coals capable of emitting 2 lbs SO2/106 Btu heat input must be defined as high sulfur coals. This is because electric utility plants built after 1971 may not emit more than 1.2 lbs SO2, and those built before 1971, and regulated under the State Implementation Plans (SIP), may exceed the 2 lbs S0 limit only under closely defined restrictions concerning ambient air quality and prevention of air quality deterioration. Coals produced in Kentucky, Illinois, West Virginia, Ohio, Pennsylvania, and Indiana account for 77% of SO2 pollution potential and are, therefore, the subject of this investigation. These coals have suffered market share losses in the 10 years from 1974 through 1984. This is due to their sulfur content as well as their lack of price competitiveness with low sulfur western and imported coals. Kentucky and West Virginia have been able to increase their compliance quality coal production substantially, but other states have largely been unable to do so because of nonavailability of low sulfur coal deposits. Assuming that the current relative productivity and cost competitiveness of coal producing states will not change significantly by 1994, this paper makes two projections of coal production in the selected states. In scenario I, the current clean air regulations are assumed to apply until 1994, and coal production in the selected states is projected to in¬crease by about 6% in 1994 compared to 1984. In scenario II, it is assumed that acid rain legislation will be implemented, which will adversely affect the markets for high sulfur coals. Coal production in the selected states under this scenario is projected to decline about 10% between 1984 and 1994, jeopardizing up to 13,000 mining jobs. The author concludes that better coal quality and greater cost competitiveness are both essential to stem the market losses and to regain strength.

Jan 1, 1988