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By J. E. Pahlman

Spontaneous oxidation and combustion of sulfides and resultant fires are serious problems in underground mines. The Bureau of Mines has con-ducted fundamental thermal gravimetric analysis (TGA) research to deter-mine the mechanisms and factors governing the rate of low-temperature pyrite oxidation, and to identify possible inhibitors to spontaneous oxidation and combustion. The mechanism of low-temperature pyrite oxidation was found to be topochemical, meaning that the oxidation process is controlled by the chemical reaction at the interface between pyrite (FeS2) and iron sulfate (FeS04)' An activation energy of 8.9 kcal/mol FeS2 was obtained for oxidation of minus 500-mesh pyrite in an atmosphere of 60 vol pct water vapor and 40 vol pct . Water-vapor content of the atmosphere was found to be the critical factor in the rate of spontaneous oxidation of pyrite, with a water-vapor content greater than 20 vol pct needed for accelerated oxidation. An atmosphere of pure oxygen resulted in negligible oxidation. The rate of pyrite oxidation is also greatly affected by the particle size of pyrite. Pyrite-oxidation rate increased drastically for particles less than 40 ~m. Room-temperature pyrophoric behavior was observed for freshly ground unoxidized pyrite (55 pct minus 5 ~m) upon exposure to air. Addition of water-adsorbing materials such as activated alumina (Alz03) decreased the rate of pyrite oxidation.

Jan 1, 1986

By Christopher J. Bise

INTRODUCTION Perhaps no other area reflects the art as well as the science of mining-engineering design quite as obviously as strata control. This is due to the fact that the design of mining structures is difficult because rock rarely behaves as a homogeneous medium. Further, the original state of stress and the mechanical proper- ties of the in situ rock are somewhat difficult to deter- mine. However, those aspects that are hard to treat mathematically can and have been successfully approached by model studies and in situ measurements. This chapter will review the fundamentals of strata- control design and will provide examples of applications to typical problems. STRESSES AROUND MINE OPENINGS A close examination of Fig. 1 reveals that, prior to mining, stresses exerted in the overburden can be in both the horizontal and vertical directions. In the example problems of Chapter 1, it was shown that the maximum compressive stress (uv) at any depth (D) below the surface can be approximated as follows: [ ] Further, the lateral stress (oh) is related to the maxi- mum compressive stress by utilizing Poisson's ratio (µ): [ ] As soon as a heading is driven, such as the one in the coal seam shown in Fig. 2, the stresses shift to the sides, or abutments, of the opening. Clearly, these compressive stresses around the opening exceed the pre-mining conditions. The problem of designing an underground opening, therefore, is to determine the maximum stresses and to note if they exceed the ultimate strength of the rock. Table 1 lists several rocks with their corresponding strengths. Figs. 3 to 5 depict the boundary stress concentrations around circular, ovaloidal, and rectangular openings. M is the ratio of the horizontal stress to the vertical stress, or as defined above, p/(l-p). This text will concentrate on the state of stress represented by M = 0.33, which corresponds to the condition of no lateral deformation in a rock having a Poisson's ratio of 0.25. For an ovaloid with an opening width-to-height ratio of 2.0 (Fig. 4), the boundary stress concentrations are at their maximum values at the comers of the opening, with values in excess of 3.0. If the depth of overburden is 500 ft, then the maximum boundary tangential stress ([ ] ) of the ovaloid is: [ ]

Jan 1, 1986

By E. Goodwin

"Maclaren Forest Products Inc.Lime is commonly used for the neutralization of acid, and the removal of heavy metals in mine waste waters. However, the precipitation of heavy metals as hydroxides must take place under conditions of controlled pH, to produce minimum concentrations of dissolved metals. An account is presented explaining the concepts applied to lime-treatment processes at Bell Mine during the period of a supernatant discharge to Babine Lake; October 18, 1982to January 25, 1983. IntroductionBell Mine of Maclaren Forest Products Inc., Babine Mining Division, an open-pit copper mine, is located on Newman Peninsula, Babine Lake approximately 65 kilometres northeast of the town of Smithers, British Columbia (Fig. 1).A depressed market for copper metal in 1982 forced a management decision to suspend operations at the mine for an indefinite period from October 3 I, 1982. For some time prior to the mine closing date, mine management was concerned about the increasing level of supernatant in the tailings pond , and its possible effects on the stability of the dam when cycloning of mill tailings ceased. Consideration of these potential hazards led to an application being filed with the Provincial Waste Management for a permit to discharge the excess water to Babine Lake. This large lake, an important link with several significant salmon spawning sites, is a major contributor to the British Columbia commercial fishing industry. Jurisdiction for the environmental affairs in the area is shared by the Department of Fisheries and Oceans, the Environmental Protection Service, and the Provincial Waste Management Branch. These agencies were directly concerned in the formulation of discharge permit following joint consultation, and exchange of information with the company on environmental concerns. An approval was finally granted to discharge approximately 1,000,000 cubic metres of supernatant from the tailings impoundment at Bell Mine into Hagan Arm, Babine Lake from October 18, 1982 with a daily discharge not to exceed 9000 cubic metres. Provisions were made that a particular discharge procedure be followed, and that the company would be responsible for setting up a sampling and monitoring program to safeguard the quality of the effluent."

Jan 1, 1986

By G. Paquet, E. Godin

"Since the mid fifty's, worlwide pellet production capacity has constantly risen (Fig. 1, Ref. 1 l. There was however at that time no real incentive to decrease production costs but only to produce as many pellets as possible. State of things changed with the iron and steel market recession, starting in 1975, and is currently leading producers to improve pellet quality and to reduce fuel and power consumption.At the Port-Cartier pellet plant, the energy needed for the induration of pellets ranks first in overall operating costs. An important step was made when starting admixing coke breeze in 1980. Oil prices justified the use of coke breeze as a direct replacement in terms of dollar/ton, however, while gaining expertise in using solid fuel a significant decrease in energy consumption was observed (Fig. 2, Ref. 2).In 1985, the ""Centre de Recherches minérales"" was commissioned to undertake research work to investigate the effects of further increasing coke addition to the pellet mix. The positive results of that work led to a 1,5% coke addition plant trial during the fall of 1985. Both laboratory and industrial results are discussed in this paper."

Jan 1, 1986

Blasts should be designed so as to maximise the percentage of charges which shoot to a free face (Le, to a pronounced rock/air interface).Free faces should be rough, parallel and reasonably near to blastholes and, very importantly, concavely·biplanar relative to the charges (see Fig. 6b). The presence of biplanar free faces requires that each charge fragments and detaches its quota of the burden rock before the detonation of:1. later-firing adjacent charges in single-row blasts, and2. dependent charges in multi-row blasts.In multi-row blasts, biplanar faces are best achieved:1. by drilling blastholes on equilateral triangular grids,2. by selecting a VI, V2 or V3 initiation sequence, and3. by using down-the-hole delays.Biplanar faces start to be created only when the conventional value of effective blasthole spacing/effective burden distance (Se/Be) increases beyond about 2.4. Values of Se/Be less than about 2.4 result in planar free faces. Where biplanar faces are created, conventional values of Se/Be are invalid and appreciably greater than real values of Se/Be.

Jan 1, 1986

By P. G. Roden, G. A. Maki, P. J. Fullmann

"In October 1984, Rocanville Division began using a stacking system to handle potash mill tailings. Previous to this date, mill tailings were slurried to the top of the tailings pile and allowed to debrine while flowing mobile stackers. The stackers consist of cyclone packs to debrine the tailings and drag conveyors with screw discharges to deposit the tailings onto the existing pile. The paper will deal with the various aspects of the project from conception to its present state. Key Words: stacking, mobile stackers, Linatex Separators, potash mill tailings, reclaim brine ponds, settling ponds, tailings pile. IntroductionPrevious to 1984, when the stacking project was first implemented, potash mill tailings were handled as follows. Mill tailings were first debrined to approximately 65% solids using a cyclone cluster. The cyclone underflows were then introduced onto a vacuum scroll filter table. Here the tailings were further debrined and discharged into the disposal tank where the tailings were reslurried with reclaim brine and pumped to the tailings disposal site. At the disposal site the slurry was discharged onto the tailings pile and by a natural settling process the tailings were debrined with the brine ultimately collected in the reclaim brine ponds. The shape and profile of the tailings pile was controlled by building salt dikes to control the flow of the discharged tailings slurry. As the tailings pile grew in height, booster pumps were added to the disposal system."

Jan 1, 1986

By Roger G. Haga, Richard P. Campbell

Introduction Project scheduling . . . . • What is it? • Why is it done? • Who does it? • How is it done? • When is it done? These are some of the questions that will be addressed in this paper. It is intended that by exploring these questions, an understanding of project scheduling, its purpose and benefits can be achieved. The purpose of this article is not to describe in detail "how to schedule", but rather to give an overview of the scheduling function on a project. A successful scheduling effort is dependent upon the people responsible for developing, implementing, and monitoring the schedule, and the tools and procedures that are used. These three elements are integral components of project scheduling. What is it? A schedule is the integration of network relationships, resource constraints and the calendar. The schedule identifies when the activities "will" be done, while the network states when they "must" be done. Networking techniques are the most widely accepted scheduling tool for complex projects. The benefit of the network is the simplicity in illustrating, presenting and analyzing the dependency relationships of a project. The logic network is the foundation of the schedule. When calendar dates are applied to the logic network, it becomes a schedule which provides the basis for evaluation of project status, forecasting the occurrence of events and controlling the project. The sequence of activities that requires the longest time to complete determines the overall duration of the schedule. This longest sequence of activities is called the "critical path". The "critical path method" (CPM) is used to describe this particular schedule network methodology. The two common conventions of CPM networks are the "arrow diagram met hod" (ADM) and the "precedence diagram method" (PIM). There are arguments in favor of each of these two methods, and usually the one that is used is a matter of personal preference, project directive or corporate procedure.

Jan 1, 1986

By S. S. Peng

Hydraulic powered supports have been employed in the U. S. longwall mining for the last two' decades. Yet the sciences of roof control at the powered supported faces remain largely unknown. Many questions remain unanswered. To address the problems, field investigations were conducted from 1980 to 1984 in seven longwall faces (1-10)(Table 1). Some of the findings have been reported by Peng and Chiang (11). The basic geological conditions of these panels are listed in Table 2. The mining equipment used in these longwall panels are listed in Table 3. Instrumentation and Observation Programs Two events were generally instrumented: one was the hydraulic leg pressures of the powered supports, and the other was the canopy and base plate load distributions.

Jan 1, 1986

Historically, the selection of mining sequence for multiple seam conditions has been based on many factors other than ground control considerations and minimization of interaction effects. Today, numerous operations are finding themselves mining above existing or abandoned operations, some of which exhibit large scale caving, pressure arches and subsidence troughs. Development of design criteria for the upper seam that would alleviate damage due to lower seam interaction effects can contribute substantially to the safety and productivity of many mining operations. The primary objective of this re search has been to identify methods of improving ground control conditions in the upper seam when mining through either passive or active subsidence waves. These methods concentrate on improved ground control practices for the upper seam, rather than reducing the amount of subsidence in the upper seams by improved practices in the lower seam. Design factors such as pillar size, pillar and opening orientation relative to the subsidence trough, offsetting intersections, and different support methods have been studied for the upper seam in a variety of typical geologic environments. Analysis has demonstrated criteria for ameliorating negative interaction effects and facillitating design in the superincumbent seam.

Jan 1, 1986

By W. M. Jr. Aubrey, R. L. Stone

INTRODUCTION Most metallic minerals and some industrial minerals have a relatively high specific gravity and most have at one time or another been concentrated by gravity methods. The principles of gravity concentration are straightforward, well documented and easy to understand, but in practice there are many pitfalls that await the engineer in the development of an integrated flowsheet. Gravity concentration equipment has a long history of use and there is a great diversity available. The equipment is capable of treating material over a broad size range, only falling off in efficiency in the very fine sizes. Gravity concentration is also inexpensive, generally free of environmental problems, and commercial equipment can process large tonnages of ore with a minimum of capital cost. As might be expected of an efficient and low cost process, the bulk of the tonnage quickly and inexpensively. Another obvious application is treating mineral deposits such as hematite and manganese ores that are sufficiently liberated after a minimum of crushing to allow production of a coarse concentrate that will meet market specifications. Except for hand picking and mechanical sorting, there is no competing process on coarse size material. Conversely, it is sometimes obvious that gravity concentration does not warrant consideration. Examples are ores that require very fine grinding to attain liberation, for ores where the minerals of interest have a gravity similar to the gangue, and ores that are strongly magnetic. Actually a mineral processing engineer may not know if gravity is the best way to concentrate an ore, but he should have little difficulty recognizing when gravity concentration deserves consideration based upon specific gravity differences and liberation gravity concentration has wide use. size. His challenge then is to At times its selection as the identify and test the applicable primary concentration process is processes, to develop a obvious. Low grade placer deposits metallurgical balance and flowsheet of tin, gold and beach sands usually for the best process, to estimate are only profitable because gravity production rate over the life of concentration is able to throw away the deposit and to make equipment

Jan 1, 1986

By L. A. Morley

Electricity is a major cause of lost-time accidents and fatalities in the mining industry. Major hazards and unsafe practices that contribute to electrical hazards are identified in this paper. Recommendations for reducing electrical accidents in surface mining are given and are directed toward the following three areas: preventing equipment contact with overhead lines, preventing maintenance related accidents, and defining system protection techniques.

Jan 1, 1986

By F. Q. Barnes

"Uranium, thorium and yttrium have been recovered, since the 1950s, from pyritic quartz-pebble conglomerates occurring in the basal sediments of the Quirke syncline at Elliot Lake. Gold, an element of interest in similar type deposits occurring worldwide, is a very minor constituent of the Elliot Lake conglomerates because the provenance area of the sediments had few gold-bearing rocks.Uranium, thorium and rare elements are found in resistate and diagenetic minerals occurring in the matrix of the conglomerates. The valuable elements are recovered from the acid soluble minerals during the milling process.Large uranium reserves occur in mines found on both limbs of the Quirke syncline. Major reserves are being developed on the South Limb of the syncline, where the Stanleigh Mine, among four past-producers, has been reactivated.IntroductionUranium and thorium commonly occur in anomalous amounts in coarse basal sediments. Radioactivity in basal conglomeratic sediments is generally attributable to resistate minerals, that rarely occur in concentrations high enough to be of economic significance. One type of basal conglomerate found worldwide, in Precambrian cratonic basins, has uranium or uranium and gold concentrations large enough to be of economic interest. Principal among the occurrences are the pyritic quartz-pebble conglomerates of the Huronian in Canada, of the Witwatersrand in South Africa and at Jacobina in Brazil.The uranium and gold mineralization of these conglomerates was derived from provenance areas where both elements occurred and were later concentrated through source degradation and deposition in placers in a reducing environment. Only the Huronian conglomerates, generally much richer in uranium but with negligible gold, have been exploited solely for their uranium content.The uraniferous conglomerates of the Huronian at Elliot Lake, 500 km north of Toronto, continue to represent a major world source and supply of uranium concentrate."

Jan 1, 1986

By J. S. Hall

The productivity and fuel efficiency of iron-ore blast furnace operations has greatly increased in recent times. These improvements, particularly to the Japanese steel industry, have primarily arisen from process engineering studies of the blast furnace and associated operations such as sintering. The development of fluxed sinter based on iron-ore fines reacted to produce highly reducible, low slag sinters has contributed further to a reduction in fuel requirements. The Pilbara province of Western Australia is a significant source of these iron-ore fines supplied to the international market. Increasingly, this market is being offered high quality ores from other foreign sources. To technically assist companies in improving the processing efficiency of ores, and to improve the technical marketing of ores by modelling the processing characteristics of Australian/international ore blends, the Commonwealth Scientific and Industrial Research Organization (CSIRO) is actively assisting the industry by researching the general theme of 'iron ore from mine to metal in this paper we outline progress towards a numerical description of: a. ore deposit classification. b. Degradation during transport. c. Direct reduction of iron ore. d. Liquid bath ore reduction/smelting. e. Industrial sintering.

Jan 1, 1986

By Rustu S. Kalyoncu

Nonoxide ceramics with improved high-temperature properties could substitute for high-temperature alloys and reduce toe Nation's dependence on imports of Cr, Co, Ni, and Mn. To meet this objective, the Bureau of Mines conducted research to synthesize u1trafine reactive boron nitride (BN) powders BN powders were prepared at temperatures ranging from -75? to 750? C. Low-temperature reactions (-75? to 200? C) between boron halides and N compounds led to formation of e1emento-organic compounds that were thermally decomposed to u1trafine (approximately 100-to 150-A particle size) reactive BN powders. BN powders were also prepared through the reaction of a 10w-me1tinginorganic B compound (boric acid, borax) with an organic N compound (carbimide and thiocarbimide) in N2 and/or ammonia (NH3) atmospheres at temperatures between 500? and 750? C.

Jan 1, 1986

By W. W. Simpson

During cyanidation of mercury-bearing gold-silver ores, significant amounts of mercury are extracted. The presence of mercury decreases. gold loading and increases mercury stripping time on activated carbon, complicates fire refining of the gold cathodes, and creates a potential health hazard. Addition of calcium sulfide (CaS) to the cyanide leach slurry precipitates the solubilized mercury but also some silver. The Bureau of Mines investigated silver loss during CaS precipitation of mercury. Addition of 0.1 lb CaS per short ton ore to a mercury-bearing gold-silver ore precipitated both mercury and silver. However, if copper was added with CaS, or if the ore contained copper, much of the sil-ver loss was prevented. Addition of 235 ppm Cu with 0.1 lb/st CaS, prior to carbon-in-pulp adsorption, resulted in 89-to 97-pct, 80-to 91-pct, and O-pct adsorption of the solubilized gold, silver, and mer-cury, respectively, on the activated carbon.

Jan 1, 1986

By R. Raymond

The occurrence and distribution of organic sulfur in coals may be investigated using electron probe microanalysis, scanning electron microscopy, and optical petrography. A general relationship exists between macerals and organic sulfur contents in individual coals: sporinite, resinite = micrinite, vitrinite > pseudovitrinite = semifusinite = macrinite > fusinite. With increasing organic sulfur content of a coal, the total range in organic sulfur content between macerals containing high and low organic sulfur contents likewise increases. Furthermore, the organic sulfur content of the vitrinite in most coals equals the organic sulfur content of the coal. For any individual coal, organic sulfur contents of macerals correlate inversely with reflectance-in-oil measurements, but the amount the organic sulfur contents vary between macerals cannot be calculated based solely on the coal's rank. Formation of pyrite following peatification has no effect on organic sulfur contents of surrounding macerals.

Jan 1, 1986

By A. V. Castelli

Since the record year of 1981 through 1985 the tonnage of weighting materials (barite and other heavy minerals) used in drilling fluids has decreased an estimated 49%. As in the past, barite still makes up over 90% of the minerals used for weighting materials in drilling fluid. Domestic production of barite has decreased 69% since 1981 and imported barite has made up over 50% of the total barite consumed since 1982. See Figure 1 for usage of barite. The number of operating grinding plants has decreased from 69 in 1982 to 39 at the end of 1985. With the decrease in demand of weighting materials for drilling fluids and an over capacity of grinding existing, prices for weighting materials have decreased dramatically. Although the American Petroleum Institute has not changed any of the specifications for barite, some users have initiated their own new specifications and have proposed to have them included in the API specifications. The exploration for and development of oil and gas dictates the level of weighting material used in the drilling fluids industry. Most forecasters do not foresee (or predict) increases in the level of drilling until the 90's. The primary or unground barite produced in the U.S. mines has a difficult time competing with imported barite due to the location of the mines. These mines are located in North Central Nevada, with small tonnages being produced in California, Missouri, Illinois, Montana and Washington. The major reason, is the difference in transportation costs from the mines versus ocean freight rates from the foreign ports to the major area of use, the U.S. Gulf. Since 1981, the production of primary barite has decreased from 2,849,000/ST (2,584,596/MT) in 1981 to an estimated 890,000/ ST (807,403/MT) in 1985(1). If it were not for the use of unit trains by some producers the production of primary barite in the U.S. would have probably declined another 20% in 1985. During the same period, of the total amount of barite consumed in the U.S., imported barite ore increased from 417, in 1981 to 65% in 1985. The major exporting countries to the U.S. during this period were China 46.27, Morocco 15.27., Chile 10%, Peru 8.4%, India 7.9%, Mexico 5.4%, Thailand 3.9% and Ireland 2.3%.

Jan 1, 1986

By W. E. Webb

This report describes a method the Bureau of Mines devised for measuring the dielectric constant and loss tangent of low-loss minerals at microwave frequencies. Determina-tion of these properties makes it possible to predict the effects of using microwave heating in potential mineral processing applications. The method uses a test cell and radio-frequency (RF) impedance analyzer with automated data acquisition and analysis. In this initial phase of microwave heating research, the Bureau measured dielectric constants and loss tangents at frequencies between 100 MHz and 1 GHz. The objective was to establish a reliable data base for use in predicting the effects of microwave heating on minerals. Comparison with previously measured data showed the Bureau's measurement method to be accurate to within ± 3 pct. This report describes the method and discusses the preliminary results of the Bureau's investigations. Typical comparative data are included.

Jan 1, 1986

By F. C. Lendrum

"This paper had its origin in a panel discussion organized by the Canadian Uranium Producers Metallurgical Committee (CUPMC) at their Annual Meeting in Blind River in May, 1964. The reason for the panel was the formation of the National Uranium Tailings Research Program funded in the amount of ten million dollars by the Federal government. According to the bureaucrats in charge the program would seek to solve the perceived problems as well as any actual problems they believed existed with uranium tailings. Dr. Allan Ashbrook of Eldorado Nuclear chaired this panel and his summation of the panel's purpose was set out as ""Regulation and the Regulatory Process in the Nuclear Industry, Where Have We Been, Where Are We Now and Where Are We Going""?"" This paper will try to relate these questions to the whole mineral industry.The Canadian Mining Association has been telling us our natural resources are today required only if they are competitively priced with those produced in so called ""third world"" countries where labour rates are lower, environmental restrictions may be less and every effort is made by the local government to sell their product. It is in this hostile economic climate that the Canadian mining industry now finds itself. Before going further I want to stress these comments apply only to: a) Mines financed by private capital in a free enterprise society.b) Mining companies want to operate in an environmentally safe manner. However, they must earn a profit because no company financed with private capital can borrow money ad infinitum.WHERE HAVE WE BEENNone of us today would want to return to the days when the mill waste management system consisted of pumping the tailing to the nearest slope and letting nature's law of gravity do the rest. This was the way it was in early gold mines in Nova Scotia, tin mines in Cornwall, England or silver mines in Cobalt, Ontario and survives today in countries where climate, customs and methods of doing business are far different from Canada. As mill tonnage rates increased, more sophisticated systems evolved often designed for mining areas with different climates. Systems imported from arid climates such as south-west United States retained the solids but allowed bacterial oxidation of the sulphides promoting acid drainage. Environmental regulations were introduced and Canadian mines learned to handle tailings in our climates."

Jan 1, 1986

By Donald Christofferson

This report documents the program conducted to demonstrate the application of a flywheel-powered propulsion system to underground mining shuttle cars. Such application can replace the present electrical trailing cable common to most shuttle cars and thereby eliminate those safety hazards, maintenance problems, and haulage route limitations inherent in the trailing cable. Described are the activities undertaken to specify and design a flywheel of optimum size, shape, and material; to identify the best concepts of input and output systems; and to achieve installation of the power package into a shuttle car. The resultant Flywheel-Powered Shuttle Car is an FMC Model 6L shuttle car, 6-ton capacity, 20-ton gross vehicle weight, powered by a flywheel system with 4.5-kilowatt-hour usable energy. The completed vehicle was delivered to the Bureau of Mines in Bruceton, Pennsylvania, for testing and evaluation.

Jan 1, 1986