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By Mark Colwell, Russell Frith

This paper summarizes the results of a research project whose goal was to provide the Australian coal industry with a chain pillar design methodology readily usable by colliery staff. The project was primarily funded by the Australian Coal Association Research Program and further supported by several Australian longwall operations. The starting point or basis of the project was the Analysis of Longwall Pillar Stability (ALPS) methodology. ALPS was chosen because of its operational focus; it uses tailgate performance as the determining chain pillar design criterion rather than simply pillar stability. Furthermore, ALPS recognizes that several geotechnical and design factors, including (but not limited to) chain pillar stability, affect that performance. There are some geotechnical and mine layout differences between United States and Australian coal fields that required investigation and, therefore, calibration before the full benefits offered by the ALPS methodology could be realized in Australia. Ultimately, case history data were collected from 19 longwall mines representing approximately 60% of all Australian longwall operations. In addition, six monitoring sites incorporated an array of hydraulic stress cells to measure the change in vertical stress throughout the various phases of the longwall extraction cycle. The sites also incorporated extensometers to monitor roof and rib performance in response to the retreating longwallface. The study found strong relationships between the tailgate stability factor, the Coal Mine Roof Rating, and the installed level of primary support. The final outcome of the project is a chain pillar design methodology called Analysis of Longwall Tailgate Serviceability (ALTS). Guidelines for using ALTS are provided.

Jan 5, 1999

By David G. Hulbert, Daniël J. Oosthuizen, Bernard Muller, Daniël de V. van der Spuy

"OVERVIEWExpert Online is a software system that aims to mimic an informed expert on a minerals processing plant. The primary objective of Expert Online is to provide operators with effective training and on-the-job assistance. Training will mostly be done off-line (e.g. using simulators of the plant), while on-the-job assistance will help the operator during normal operation to run the plant in an effective manner. Although Expert Online focuses on the operator, it is expected to also be of value to other plant personnel, including supervisors and plant managers.Expert Online will make use of novel artificial intelligence technology (developed by MINTEK) known as the Interpreting Expert System (IES). Previously the IES was only able to explain the actions taken by the controller. Expert Online will extend this functionality to also take the process dynamics into account.Expert Online is being developed by a South African consortium comprising MINTEK (focusing on the research and development of the system as a whole), Lonmin (globally one of the biggest platinum producers, with plants suitable to test Expert Online on) and the University of Natal (academic partner, providing consultation and focusing on complex process modelling). The initial application of Expert Online will be on milling- and flotation plants.This paper discusses the structural design of Expert Online for implementation on the latest version of MINTEK’s open control platform, PlantStar, as well as several advances and additions since the previous implementations of the Interpreting Expert System."

Jan 1, 2003

By Percy Coulter

"GeneralThe Malartic Gold Fields mill is presently treating ore from four different companies. The ore is delivered by truck, by the shipper, and dumped into the Malartic Gold Fields ore bin. This ore varies in size from 'Run of mine' to minus 8 11, All ore is crushed to pass a 3/ 8 11 x 4 11 cross slotted screen.Weighing and SamplingAfter crushing the ore passes over an adequate-weigher (250 T .P.H. capacity) on its way to the fine ore bins. This weigher has a remote counter at the sampling point.A 36” Vezin sampler cuts 15% from the ore stream as it enters the mill bins. This portion is elevated, by means of a short belt conveyor, to feed the No. 2 Vezin sampler. (This elevation is necessary to give headroom for the following machines). The ore on the conveyor is levelled to give a uniform stream to the second sample cutter. The second, third and fourth cutters are 2411 •The No. 2 Vezin sampler cuts 19% from its feed and this portion falls directly into a 1611 x 1011 Denver rolls crusher. This crusher reduces the ore to -1/ 8 11 , approximately. The discharge from the rolls crusher falls on a revolving circular plate with scraper, which gives a uniform stream to the No. 3 Vezin sampler.The No. 3 Vezin cuts 10% from its feed. This portion falls on another revolving plate to give a uniform stream to feed No. 4 Vezin."

Jan 1, 1966

By Roswell Johnson

THE Manual of the Oil and Gas Industry, under the Revenue Act of 1918, published by the Treasury Department for the guidance of oil companies in preparing their estimates of future recoverable oil for the purpose of calculating depletion, gives the first large public collection of comparative decline curves for the whole country. It is a matter of both scientific and practical interest to so arrange these data that the pools can be readily compared. There are certain difficulties in such a comparison, however: 1. The economic limit varies from 50 to 2000 bbl. a year in different pools taken. (2) Because of this variation in economic limit, the range of data shown makes comparison possible only for wells of intermediate size. In order to be as inclusive as possible, I have taken as an expression of the rate of decline the amount of oil produced in the period during which a well drops from 3000 to 2000 bbl. a year.. No period of smaller production could be used because of the high economic limit in California; and no period of larger production and yet include the small well areas of the Appalachian. As it is, the Lima-Indiana wells are excluded. In a few instances, curves were extrapolated to obtain the reading, where the curve seemed regular enough to warrant it. In general, the amount of oil, in barrels, produced while a well declined from 3000 bbl. to 2000 bbl. a year is shown in Table 1.

Jan 1, 1920

By Naoki Hiroyoshi, Hajime Miki, Tsuyoshi Hirajima, Masami Tsunekawa

"Acidic solutions for chalcopyrite leaching contain cupric, ferrous and ferric ions, and the oxidizing power of the solutions is indicated by a redox potential, determined by the concentration ratio of ferric to ferrous ions. This paper reviews our previous studies on the redox potential dependence of chalcopyrite leaching in sulfuric acid solutions. Leaching experiments indicated that copper extraction rate was accelerated by cupric and ferrous ions , and was higher at potentials below a critical value (critical potential) in the presence of these ions. This was in good agreement with the results of electrochemical measurements, i.e. in the anodic polarization curves for a chalcopyrite electrode, active-passive behavior was observed when cupric and ferrous ions co-exist, but the activation at low potential did not occur without these ions (Figure 1). The analysis of the AC impedance spectra for chalcopyrite electrode indicated that a high-resistance passive layer grows on the chalcopyrite surface without cupric and/or ferrous ions, and that coexistence of these ions causes a formation of another product layer and inhibits the passive layer growth at low potentials.These results are explained by a reaction model assuming the formation of intermediate Cu2S due to chalcopyrite reduction and the subsequent oxidation of the Cu2S. The critical potential corresponds to the redox potential of the Cu2S formation and the predicted values matched with the experimental results. Catalytic effect of silver ions on chalcopyrite leaching was also discussed based on experimental results and the above model."

Jan 1, 2003

By Thomas E. Krakowski

INTRODUCTION This paper presents a method of calculation of internal moments and thrusts in tunnel liners of any material. The method is intended for hand calculation, but as it is numerical in nature it can be programmed for use in computers. The information required for design or analysis is presumed active internal and external pressures, the effective moment of inertia and cross-sectional area of the tunnel liner, the liner's modulus of elasticity, and the coefficient of subgrade reaction of the material surrounding the tunnel. Starting with active external and internal pressures and following the procedure described herein using the structural properties of the liner and surrounding medium, external passive reactions and internal moments and thrusts are calculated. The method is intended for use by design engineers. GENERAL Experience has shown that flexible tunnel liners have less internal bending stresses than those calculated using classical structural theory applied to free-standing arches. In fact, liners which are very flexible in relation to the surrounding soil or rock have virtually no bending moments when compared to free-standing arches supporting equal active loads. In order for this phenomenon to occur, the original active pressure diagram must be "modified" by additional external reactions (passive pressures) from the surrounding material, so that the thrust line within the lining follows the centroid of the tunnel arch more closely than that of a similar free-standing arch. These passive pressures are reactions developed through outward deflection of the tunnel liner. Thus, the outward moving portions of a

Jan 1, 1979

By E. H. Dix

The purpose of this paper is to introduce the subject of Stress-Corrosion Cracking at the Joint Symposium of the American Society for Testing Materials and the American Institute of Mining and Metallurgical Engineers. In this Symposium we are concerned with spontaneous cracking which results from the combined effect of high, prolonged stress, and corrosive attack. Stress-corrosion cracking may be considered a specific phase of the more general subject of the effect of high constant stresses in accelerating the rate of corrosion of metal products. The kindred subject of corrosion fatigue in which the life of a metal part under repeated stresses is reduced by corrosive attack is not a part of this Symposium. Likewise "creep," the failure of a metal part by high constantly applied stress, in the absence of a corrosive environment, is not to be included. How- ever, corrosive influences may have more effect on creep failures than is sometimes realized. The penetration by molten metals such as often occurs during tinning or soldering operations, or by mercury, frequently causes cracking when the part is stressed. Strictly speaking this is not stress-corrosion cracking. However, the committee has decided that this phenomenon should be included in the Symposium. The penetration by molten metals such as often occurs during tinning or soldering operations, or by mercury, frequently causes cracking when the part is stressed. Strictly speaking this is not stress-corrosion cracking. However, the committee has decided that this phenomenon should be included in the Symposium.

Jan 1, 1945

By Charles H. Kline

Industrial minerals are the Cinderella of the mining I industry. Often considered as just dirt by traditional hard-rock miners and oil drillers, these products nonetheless comprise the second largest segment of the industry and the fastest-growing one. With annual sales of $5.7 billion, industrial minerals make up 22% of the dollar value of U.S. mineral consumption. By contrast, metals account for only 11%. Fuels, of course, are dominant, representing 67%. The dollar values of production and consumption for the three segments of the industry in 1968 are shown in Table I. The growth rate of industrial minerals is shown in Table II. From 1953-1968, the current-dollar value of all U.S. mineral production increased at the average annual rate of 3.8% while that of the gross national product increased at 5.9%. Industrial minerals are the only segment of the mining industry that approached the growth of the overall economy in this period, with production values increasing 5.1% a year and apparent-consumption values 5.6% These data, expressed in current dollars, include the effects of inflation. The growth in U.S. industrial minerals production is even more striking when expressed in indexes of physical production. As shown in Table III, production of all U.S. industries increased by 4.1% a year from 1953 to 1968. The growth in production of stone and earth minerals just matched this figure. By contrast, the average annual increase for oil and gas was 2.2%: for metals, 1.6% : and for coal, only 0.6 %.

Jan 1, 1970

By M Maulen, K van Olden, T Elkington

A large open pit is planned at the site of an existing underground mine. Strategic planning work has been undertaken to investigate the project value associated with an expansion of underground operations in conjunction with the proposed pit. Previous analysis determined that any depletion of the available pit resource that changes the optimum pushback sequence has a significant negative impact on overall project value. Therefore, to investigate the viability of an underground expansion requires a system to identify the stope resource that would add value to a combined operation if mined ahead of the pit. The key solution to this problem was to generate a resource model variable that defines the blocks, which have greater discounted value if mined by stoping rather than by the open pit. A key assumption was that the open pit is the main driver of overall project value. From the identified potential underground resource, series of mineable stope shells were generated at a range of cut-offs. This stope reserve data formed the input to a mixed integer program based software tool that was developed and used to optimise the underground mining plan, at conceptual level, for various project configurations. A dynamic programming tool was then applied to evaluate the optimum value of a combined open pit and underground operation. The overall objective of this project was to determine the strategic plan that yields the optimum value for a combined undergroundûopen pit operation. The results provided a clear focus and direction for more detailed design and planning.

Jan 1, 2009

THE Institute of Metals Division of the American Institute of Mining and Metallurgical Engineers appointed a Research Committee in 1925 to de-termine what could be done by the Division to promote research work in the United States on non-ferrous metals. It was recognized that a considerable volume of use-ful research work in this field was being carried on by individual industrial companies as well as by the uni-versities and other research institutions, such as the Bureau of Standards. One need only refer to the recent research literature on non-ferrous metals in the TRANS-ACTIONS of the A.I.M.E., and elsewhere, to be con-vinced of healthful research activity in this field. The appointment of the Committee, however, voiced the impression that what was being done was, however excellent, inadequate for the needs of the industry. It was felt that the Division could usefully devote some of its time and effort in stimulating activity in this field. The deliberations of the Research Committee have not failed to develop wide differences of opinion even on the simple questions: (1) whether the Division can aid in any practical manner in this cause and (2) what the features of such aid might be. However, it was clear that there was only one type of research activity with which it could be very usefully concerned, viz., that on subjects of fundamental interest and general application. It was obvious that research work having special objects and more immediate commercial appli-cation should best be left to the industry itself.

Jan 1, 1927

By Clemens Jones

A FORM of apparatus for volumetric determinations which shall possess the features requisite for the constant use of potassium permanganate, is a very desirable fixture in a well-appointed laboratory. Expedients to economize time and insure accuracy are generously employed for various other reagents, while the permanganate seems to have been consigned to the use of an ordinary burette, and a black bottle of the standardized solution. But the introduction of new, methods dependent on the permanganate, and the urgency of technical requirements where it has long been indispensable, require a revision of practice more scientific and convenient. The apparatus designed by the writer for the laboratory of the Thomas Iron Company at Hokendauqua, Pa., is so simply contrived and works with such entire satisfaction that a description of it may be of use. It may be prefaced that the characteristics of a suitable device offer some mechanical difficulties when combined in a continuous appliance. The burette, B (50 C.C. to 1/10 c.c.), shown in Fig. 1, consists of the graduated tube proper and an arm, E, fused to it below the 50 c.c. mark. At its top are rubber connections with the blast-aspirator, shown in the cut (Fig. 2) of the front view. By means of an extra stopcock, D, connection through E is established with the reservoir F. The burette is clamped Securely to a slide, C, which is counterpoised and moves freely on guides between two parallel sides, L, suitably

Jan 1, 1887

By Francois E. Heuze

INTRODUCTION Rock slope stability typically is governed by the geological discontinuities. This stability can be improved by drainage, unloading, adjustment in slope orientation, adjustment in slope angle, and also by reinforcement such as bolts or tendons. This paper addresses the modeling of bolt reinforcement in rock slopes. Most geological discontinuities such as joints have rough surfaces; they dilate during shear displacement. Their shear strength is characterized by two envelopes (Fig. 1); the peak envelope is distinct from the residual envelope when the joints can dilate. If the dilation is restrained transversely, such as with bolts, the normal stress will increase and the joint strength also will increase. It has been demonstrated that this increase may be considerable (Obert et al., 1976). The new normal stress at peak strength is a com- posite function of the properties of the rock mass and of the steel (Heuze, 1979). Hence, the dilation and its resultant effects on joint shear strength must be included in the analysis and design of the reinforcement. We present a model which can be used to perform such an analysis. Early studies of rock reinforcement in mining and civil applications (Lang, 1957 - Panek, 1962 - Osen and Parsons, 1966 - Lang, 19721, did not consider the specific effects of bolts on joint shear strength. In 1973, Heuze and Goodman reported direct shear tests on rough sandstone joints with bolts perpendicular to the joint plane (Fig. 2). The peak shear strength of such joints was greatly in excess of the shear strength of the unbolted joints. The excess

Jan 1, 1983

By B. H. Bergstrom, C. L. Stevenson

In any process wherein materials must be pelletized, a crucial period exists during the interval between the manufacture of the green pellets and their subsequent heat treatment. The pellets must possess sufficient strength and plasticity to withstand conveying and sizing. Also, they must be neither excessively brittle nor excessively plastic when placed in a bed on a traveling grate for heat treatment. An acceptable pellet, in the event sufficient forces are applied to cause its fracture, will merely split into two or three large fragments which would not drastically alter the permeability of the pellet bed. However, when brittle pellets are broken, they shatter so that the resultant fine particles either decrease the perimeability of the bed or become entrained in the gas stream and thereby cause dust problems. On the other hand, excessively plastic pellets may not break at all but may merely flatten. This would seriously hamper bed permeability. In the past, pellet evaluation has usually been left to the creative and artistic talents of the balling operators. A multitude of techniques have evolved, of which the most common is the simple compression test. For this test a hand-operated device loads single pellets to fracture, but reliability is strictly operator- dependent. Probably the next most common test is the drop test, which is conducted according to several different procedures. This is an important quantitative test for estimating the amount of handling a pellet will withstand. There are many other tests, such as pinching a pellet between the thumb and index finger, but none are capable of producing reproducible results.

Jan 4, 1962

By K. V. S. Sastry

INTRODUCTION Flotation is by far the most important unit operation of mineral processing. Most of the common base metals are produced from flotation concentrates of metal sulfides. Flotation is proving to be a valuable process for producing non-metallic minerals such as ceramic oxides, silicates, clays and phosphate fertilizers. All these fltation separations are carried out on materials which are predominantly in the 10 to 100 micronssize range. The present practice of flotation circuits for the production of large tonnages of metallic and non-metallic mineral concentrates is fairly well understood, developed, and standardized. Two kinds of problems are encountered (1, 2) with fine particles smaller than 10 microns. First is the interference of the fine particles with the flotation of the coarser particles and the second is the poor flotation behavior of fine particles. In general, these problems are handled either by eliminating, minimizing, or ignoring the fines in the flotation circuit feed. In recent years, however, a considerable number of experimental and theoretical studies have been carried out to understand the flotation behavior of fine particulates. This literature is summarized in four review papers (2-5). Application of the conventional flotation techniques for the separation of mineral fines is known to be uneconomical, and to present several scientific and technological problems (1-5). The objective of this presentation is to examine the flotation behavior of fine particles in terms of their physical-chemical properties and to review the potential techniques of fine particle flotation. Specific reference to the problem areas and research needs will also be included in the discussion.

Jan 1, 1979

By W. Scott Dunbar

INTRODUCTION Methods of linking boundary integral (BI) solutions with finite element (FE) solutions have been well described theoretically in other publications (e.g., Zienkiewicz, et al., 1977). The purpose of this paper is to point out some practical aspects of implementing these linkage algorithms for rock mechanics applications and to describe an alternative linkage algorithm. A not so obvious motivation for using FE-BI linkages is illustrated in Figure 1. The cost of modeling the excavation of the crusher station by the BI method can equal or exceed the cost of an equivalent FE model. This is due to the large surface area which must be modeled by the BI method. In such cases, a more economical approach might be to model the region immediately adjacent to the crusher station by finite elements and link the solution at the periphery of the FE model with a solution for the exterior region. Further economies can be realized if some flexibility in the type of exterior solution is possible. For example, it would be desirable to use a coarse discretization of the FE-BI interface for a BI solution, which would be less demanding of computer storage and time, or to use an analytical solution for the exterior region. The alternative linkage algorithm to be discussed allows such flexibility. BOUNDARY INTEGRAL METHODS There are basically two formulations of BI methods: the direct and the indirect. FE-BI linkage algorithms depend on which formulation is used. The purpose of this section is to briefly describe each method.

Jan 1, 1982

By Robert Richards

Introduction. WALTER RENTON INGALLS recently gave a very interesting talk before the student mining society of the Massachusetts Institute of Technology. In it he showed the present status of mining as he sees it. In the talk he explained that, while discoveries of great mining districts are still possible, they are being made less and less often; that the great mining companies, although there are large numbers of mines offered for sale, are finding it less and less possible to find new mines that are up to the standard of richness that they wish for purchase. The natural conclusion is that better, cheaper, and more efficient methods of mining, concentrating, and smelting are needed if the lower-grade mines are to be worked at a profit. This is true also if the high-grade mines are to yield their greatest profit. This paper deals only with the concentrating or ore-dressing problem, and gives some thoughts which point towards greater profit. In a paper on The Development of Hindered-Settling Apparatus the writer showed that there were two main lines of work that would lead to greater saving, lower tailing assay, and, therefore, greater profit. The first is the recrushing of middling or tailing products for the saving of the contained values. This is well understood and largely carried out in existing mills. The second is the saving of fine free mineral at the first opportunity. This is a matter less well understood in the mills, and to dwell upon this the present paper is written.

Jan 9, 1913

By Niels Hansen

Dispersion-strengthened sintered aluminum products are normally manufactured from surface-oxidized aluminum powder, SAP,' or from atomized aluminum powder.' The powder-blending technique consisting in mechanical mixing of the constituents has been used in the manufacture of numerous dispersion-strengthening alloys, but has never been successfully applied to aluminum alloys. This technique may, however, be applicable if it is possible to produce an unagglomerated mixture of the constituents in the form of fine powders, and the present study was initiated to test this assumption in the aluminum/aluminum oxide system.3 Table I lists the two grades of aluminum powder and the aluminum oxide powder used in the manufacture of the alloys. MD 105 A is an aluminum powder with an average grain size less than 5 IJ., which is isolated from the MD 105 powder by centrifugal separation in cyclohexanol. The aluminum powders and the aluminum oxide were blended in a high-speed blender and then subjected to cold compaction and hot compaction at 550" to 600°C. Subsequently the compacts were hot-extruded with an extrusion ratio of 15 to 1. The alloys were evaluated by tensile testing at room temperature and at 400°C, microscopic examination, and heat treatment at 500°C followed by tensile testing at room temperature. For the purpose of comparison, a commercial SAP quality supplied by Swiss Aluminium Ltd. was tested in the same manner. The tensile data at room temperature and at 400°C are plotted in Figs. 1, 2, and 3 as a function of the weight percent of oxide. The oxide content

Jan 1, 1964

By Jafar Safarian

A new kinetic modelling technique has been applied to study the rate of MnO reduction from slags by dissolved carbon in liquid iron. Based on a consideration of the fundamental reaction kinetics, a general differential equation for the slag reduction process has been formulated, which then is solved analytically for certain special cases. By re-writing the solutions in a dimensionless form the poorly known kinetic constants are eliminated, which makes the model both flexible and applicable to a wide range of experimental conditions. For example, changes in process parameters such as temperature, metal composition, slag composition, gas composition can readily be captured along with variations in the crucible geometry. Based on a comparison with a series of experimental reduction curves being reported in the literature, it is concluded that the predictive power of the model is surprisingly good, yielding outputs that are both reliable and physically reasonable. An attractive feature of the model is that it provides an alternative way of calculating the activation energies from experimental reductions curves. For example, it suggests that the activation energy for the reduction of MnO by dissolved carbon in liquid iron is about 90 kJ/mol, which is different from the previously reported activation energies. The model also suggests that the crucible geometry, which affects both slag the volume and the resulting slag/metal interfacial area, has a significant influence on the overall reduction kinetics. This design parameter is a key to further optimisation of the process on an industrial scale.

Jan 1, 2006

By M. R. Stokes, A. J. Ross

As a result of higher levels of production and greater depths the underground mining industry in the UK is encountering hotter climates in some workplaces. It is widely known that high heat and humidity in an underground workplace will result in reduced performance and may also seriously impair the health and safety of the workforce. It is therefore of paramount importance that the underground environment can be effectively predicted in order that sound decisions can be made to reduce the effect of climate on working personnel. Over the years numerous computer programs have been developed to predict the climate in underground mines. One such important program, developed by McPherson, is CLlMSlM TM which is a prediction model of heat and mass transfer along a single roadway. This paper describes the development of a computer model, called DRIVE, to predict the climate in single entry drivages. This program incorporates the theory established by McPherson for the development of CLlMSlM TM. This involves the algorithm for unsteady state heat transfer. Throughout the development of DRIVE underground measurements have been taken to correlate the program. DRIVE will predict the changes in psychrometric and thermo- dynamic conditions along the drivage depending on a number of parameters. Firstly, the type of auxiliary ventilation system employed such as forcing, exhausting, overlap. Secondly, the addition of heat and moisture due to autocompression, machinery, strata, fissure water, oxidation, broken rock, fallen rock, refrigeration equipment, lighting, personnel, service water. Thirdly, due to the velocity of the air.

Jan 1, 1997

By R. H. Johnson

The Manual of the Oil and Gas Industry, under the Revenue Act of 1918, published by the Treasury Department for the guidance of oil companies in preparing their estimates of future recoverable oil for the purpose of calculating depletion, gives the first large public collection of comparative decline curves for the whole country. It is a matter of both scientific and practical interest to so arrange these data that the pools can be readily compared. There are certain difficulties in such a comparison, however: (1) The economic limit varies from 50 to 2000 bbl. a year in different pools taken. (2) Because of this variation in economic limit, the range of data shown makes comparison possible only for wells of intermediate size. In order to be as inclusive as possible, I have taken as an expression of the rate of decline the amount of oil produced in the period during which a well drops from 3000 to 2000 bbl. a year. No period of smaller production could be used because of the high economic limit in California; and no period of larger production and yet include the small well areas of the Appalachian. As it is, the Lima-Indiana wells are excluded. In a few instances, curves were extrapolated to obtain the reading, where the curve seemed regular enough to warrant it. In general, the amount of oil, in barrels, produced while a well declined from 3000 bbl. to 2000 bbl. a year is shown in Table 1.

Jan 1, 1921