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By Willian dos Reis Lima, Breno Totti Maia, Guilherme Silva Machado, Adeilson Roberto Ribeiro, Carlos Alberto de Souza, Douglas Augusto Pena, Fabrício da Silveira Garajau, Marcelo de Souza Lima Guerra

Oct 1, 2019

By R. W. Raymond

THE sudden death of Dr. Drown, on Nov. 17, 1904, brought to multitudes the pang of personal loss. Of all those who, as students at Lafayette College, the Massachusetts Institute of Technology and Lehigh University, were brought under his influence, not one can have forgotten him, and all will now acknowledge, with an unfading gratitude, what they owe in professional equipment to his wise, clear, skillful guidance, and in character to his sympathetic, stimulating personality. Outside of this wide circle, the experts of both hemispheres will lament the departure of one who, in his department of scientific inves¬tigation, both knew what work was needed, and how to do it so that it needed not to be done again. The older members of the American Institute of Mining Engineers will remember with gratitude the part he took, as Secretary from 1873 to 1884, in promoting the interest of that organization, and in shaping the plan and policy which it has followed with such noteworthy success. And, in addition to all these (though partly included among them), there are a host of friends; who can scarcely think of the public loss by reason of their own. Thomas Messenger Drown was born March 19, 1842, at Phil¬adelphia, Pa. His parents were natives of New England; and. he was the youngest of their three sons. He received his preliminary education in the public schools and the High School of Philadelphia, from which he was graduated in 1859. Be-

Jul 1, 1905

By Stephen Krickovic

THERE is today no proven continuous mining machine that can be used under all the varying conditions found in most bituminous coal mines. During the last five years, however, both the machines and methods of using them have been improved. Coal operators can now exercise a fair degree of selectivity, except those who must struggle with thick and hard partings and operate in seams less than 40 in. thick. It must be realized, then, that more than one type of continuous miner may be used advantageously in the same coal mine. Although this appears to violate the standardization generally sought after, the gain can more than offset added maintenance and, to a lesser degree, operating problems. It is equally important to realize that most coal mines will require one or more conventional loader setups in the plan of total mechanization. Seam variations, adverse physical conditions, and special designs of main headings for haulage and ventilation, to name only three reasons, may render 100 pct use of continuous miners uneconomical. Of course, this situation may change. In one mine, prior to 1954, machines that ripped coal from the face achieved only average results, as compared to the impressive performance of boring-type units in another coal mine. A boring-type unit was therefore selected for the first mine in the belief that the chance of success would be greater if it were used in a combined cutting and wedging operation, cutting a full face in its forward movement. The miner was located in a part of the mine where a special projection could be readily developed, a solid area scheduled for mining according to the overall progress plan. All openings surrounding the particular area were carefully investigated for the presence of sulfur balls. Fig. 1 (opposite) shows the representative seam section, containing the several pyritic bands 1/4 to 1/2 in. thick; a characteristic Pittsburgh seam parting averaging 5 to 6 in. of slate, coal, and slate; and the occasional py-rite lenses 2 to 3 in. thick. Under normal conventional mining methods, the top 10 or 12 in. and the bottom 4 in. are left in place to offer roof protection

Jan 1, 1958

By R. J. Brison, R. D. MacDonald

This chapter is written primarily from the standpoint of development of flotation processes for treatment of specific ores. However, most of the principles and techniques discussed are equally applicable to other forms of applied flotation research such as improvement of procedures in existing plants and development of new reagents or methods without reference to specific ores. Because it is not possible to cover this subject in detail here, much of the material presented is in the form of a guide to the literature, with emphasis on readily available, English language publications. Details are included only if considered to be widely applicable or not readily accessible in the literature. Related subjects which are covered elsewhere in this volume, such as flowsheet design and statistical techniques, are not emphasized in this chapter. The material presented is intended primarily for those familiar with the theory and practice of flotation, but with limited experience in applied flotation research. It may be of value to the experienced investigator as well, by calling to his attention some useful viewpoints or techniques that he may have overlooked. It has often been said that flotation is more of an art than a science. In- deed much of the knowledge of flotation behavior is empirical. However, the unsystematic, artisan approach to flotation problems results only in an ac- cumulation of unrelated data. Like any other physical or chemical process, flotation is amenable to investigation by systematic experimental methods. The experiments should be based, as far as possible, on an understanding of the natural laws that govern the process and on good experimental procedures. SAMPLING OF A DEPOSIT In working out a flotation process for a mineral deposit, no step is more important than that of obtaining a suitable sample or samples. Experimental procedures cannot compensate for the lack of a representative sample. Methods of sampling a deposit have been published elsewhere,l-4 but a few points are of special importance from the standpoint of flotation. First, the sample should be representative of the deposit, not only in respect to chemical composition, but also in respect to mineralogical composition and degree of dissemination of the valuable minerals. To help ensure

Jan 1, 1962

By J. J. Shields

The Bureau of Mines observed two different types of longwall mining; the advancing longwall method and the retreating longwall method at seven bituminous coal mines located in five of the coal producing states. In two mines using the advancing face method, the coal faces were abandoned because of roof failures and poor production. In the other five mines using the retreating method, roof control was good as was production. The success of longwall mining depends on good roof control and good production. The use of self-advancing hydraulic roof supports, such as Westfalia, Dowty, and Wanheim props along with Westfalia coal planers and the Anderton and Eickhoff shearer-loaders have been a factor in recent successes.

Jan 1, 1967

By H. G. Moulton

THE INCREASING scale of mining operations over the past decade, particularly in connection with the exploitation of large bodies of comparatively low-grade copper ores, has made necessary the study of the behavior of rock under failure. When extractions of as high as 40,000 tons a day are reached in steam-shovel operations and as high as 20,000 tons per day in underground mining operations, it is evident that excavations soon become of such size that the behavior of the rock will often be dependent on fundamental principles applicable to more or less homogeneous materials under stress, rather than on weathering or local planes of weakness. In the ordinary rock excavations encountered in civil-engineering practice, such as excavations for subways, highways, railroads, etc., rock faces are not carried to depths sufficient to develop any pressures, or consequently any failures, except those due to loose masses resulting from local planes of weakness. When, however, orebodies whose tonnage is measured by hundreds of thousands are to be extracted by steam-shovel operations in open pits, it is necessary to determine if possible in advance the slopes that the sides of the pits will assume, for these slopes influence the shape of the excavations and thereby the amount of ore that may be extracted by the open-pit method. When large orebodies are mined underground at depths of 300 to 500 ft. (91 to 152 m.) by caving methods, the surface subsiding as the ore is withdrawn, it is necessary to determine in advance the outermost limits of fracture in the subsiding surface, for on the delineation of these limits depends the location of surface improvements, mills, smelters, etc. and also the location of development work, such as hoisting shafts, main haulage levels, and all other underground openings that must remain undisturbed during the period of ore extraction. Heretofore, in civil-engineering practice, it has been assumed that earth and rock behave differently in failure. For the calculation of earth

Jan 2, 1920

By N. H. Coates

The Bureau of Mines designed, constructed, and operated an 18-inch- diameter, fluidized-bed combustor for highly caking coals to evaluate the method for possible application to power generation. In initial tests, combustion was carried on successfully, with carbon losses as low as 0.8 percent. Coal ash was not satisfactory as a bed material, but mullite proved to be acceptable. Overall heat-transfer coefficients from the bed (-20 + 50 mesh mullite) to a water-cooled tube were about 75 Btu/hr-ft2- °F.

Jan 1, 1970

By Edilson Alves Maranhão, Willi Anton Buehler, Andrea Pezza

O processo de vácuo é a solução típica aplicada quando é hora de atualizar o nível de qualidade do aço de uma aciaria existente. A instalação do novo RH de 150t na planta integrada no México e do novo Tanque de Desgaseificação à Vácuo (VD) de 135t no Canadá para a produção de tubos de alta qualidade na América do Norte, é um exemplo típico disso. A nova instalação inclui o mais recente conceito de design para este tipo de equipamento: operações totalmente automatizadas e menor impacto ambiental são os principais aspectos do projeto. A solução de bomba mecânica a seco continua sendo a melhor abordagem de investimento e custo operacional para a tecnologia de vácuo. Os principais aspectos do design da planta são descritos neste documento. O desgaseificador a vácuo RH é responsável pela melhoria das performances metalúrgicas, especialmente em termos de nitrogênio, remoção de hidrogênio e nível de pureza do aço. O processo a jusante e os resultados estão estritamente ligados aos desempenhos do desgaseificador a vácuo. O monitoramento de processos através da linha completa é uma prática padrão para obter uma melhoria efetiva da qualidade do produto. Os desempenhos metalúrgicos do desgaseificador a vácuo RH são analisados e comentados neste trabalho.

Oct 1, 2019

By N. Saeidi

"For a mine section comprising 100 different stoping activities, there are ~10158 different possible stoping sequences to explore in the mine schedule optimization search and this figure is indicative of the challenge posed in maximising profitability through mine schedule optimization. Any viable excavation timetable must satisfy all conditions that will constrain feasible sequences. In general, consideration of additional constraint will tend to reduce profitability. Geomechanical constraints on mine schedules, introduced with the aim of avoiding excavation instabilities, can be established for underground mine schedule optimization processes but, currently, this is done only rather crudely; precedence relations between excavation activities (stoping or development) are encoded to produce particular stope sequence patterns (e.g. ‘chevron’ or pyramidal sequences).The work reported aimed to investigate automated processes for formulation of such geomechanical constraints. The timetable optimization approach to the problem precipitated a fresh look at time dependent, and excavation sequence dependent, deformation. Abundant evidence emerged that supported the notion that any methodology developed that failed to account for such time and sequence dependencies was itself bound to failure. Yet, it was also recognised that any deviation at all from an assumption of linear elasticity for geomechanical numerical modeling, adopted to inform geomechanical constraint formulation, would rapidly lead to impractical levels of computational effort.Thus the approach presented here represents a compromise: the elastic stress fields induced by individual mining activities are computed for each activity, independently considered within a computational domain identical to that of all other activities. These are stored in a database. Instability risk for a specific sequence of mine activities can be assessed through superimposition of stress fields drawn from the pre-computed database. Time stamps allocated to events corresponding to the excavation processes permit the conversion of the sequence of elastic stress fields into a time table of stress fields. For specific locations within the computational domain, the aggregate effect of the sequence of mine activities leads to stress-time histories having a castellated appearance, a result of the near instantaneous response of an elastic medium to the excavation perturbations. Time dependent deformation responses are crudely allowed for through the re-profiling these castellated histories through graphical construction of curved responses, where the curve gradients are related to a viscous parameter. Such curves form the basis for a geomechanical constraint formulation to be adopted for mine schedule optimization."

Jan 1, 2015

By Norbert L. Piret

From the copper smelter operation a various outputs are to exert negative impacts on the environment. Particularly, emissions like SO2 and dusts, solid wastes like contaminated gypsum and arsenic sludge, effluents and heated cooling water are generated. Most countries have sitipulated an environmental legislation to provide the framework for an efficient environmental management. As part of the planning activities for x major greenfield copper smelter, the Following elaborations are required and were drawn up. -The Environmental Impact Assessment which is the first step of environmental management and mandatory for every new project. - The Environmental Monitoring Plan, which describes the monitoring procedures for the main impacts identified and assessed. -The Environmental Management Plan, which deals with the minimization of environmental impacts, and the methods of its implementation.

Jan 1, 1996

By S. Gqebe, A. Lewis, M. Rodriguez-Pascual

"Gravitational sedimentation of suspensions in various precipitation processes is hindered by colloidal stability. This is due to the high surface charge of the suspension and results in strong attraction/interaction between the ions on the particle surface and counter-ions in solution. Moreover, this strong interaction results in a charge build-up that renders the suspension stable. In order to induce gravitational sedimentation of these particles, a redistribution of ions close to the particle surface is required. We therefore sought to redistribute ions close to the particle surface by applying a magnetic field. This results in the reduction of interparticle electrostatic repulsive forces and a subsequent increase in the zeta potential of a suspension. For the purpose of this study, a copper sulphide suspension was used. Copper sulphide particles were exposed to a range of field strengths for set exposure times and their zeta potential was measured before and after exposure. All particles had an initial zeta potential value equal to or less than –40 mV prior to magnetic field exposure. A significant increase in zeta potential was observed, with values reaching a maximum of –16.5 mV when exposed to a 2 T field strength for 40 minutes. This is due to Lorentz ion shifts resulting from the Lorentz force exerted by the magnetic field on the particle surface. IntroductionMetal sulphide precipitation is a process for treating industrial waste streams with high residual concentration of mixed metal ions and acid mine drainage. This process has been reported to achieve higher metal removal compared with the industrially prevalent lime precipitation, due to the sparingly soluble nature of the metal sulphide precipitates formed. However, precipitation is governed by high levels of supersaturation which favour the formation of fine particles (Lewis, 2010). Consequently, some of the metal sulphide fines formed exhibit slow settling dynamics and thus tend to remain suspended in solution."

Jan 1, 2016

By Robert H. Richards

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 pssible, 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. The writer is himself convinced and hopes to convince the reader that, for obtaining the above result, not only is better classification needed, but also more classification; that is, there should be more spigots or pockets to the classifier. His attention was recently called to a small mill which had a three-pocket cone classifier fed with from 2 to 0 mm. stuff, supplying feed to four Wilfley tables. The first spigot had about all the

Jan 1, 1914

By Lewis E. Dr. Young

MINING engineers in the United States understand that mining conditions in the British coalfields are much more difficult than in most of the mines now being operated in the United States. We realize also that in the near future we must face some of these same difficult problems. Early Mechanization The progress made in mechanical loading in the United States is the result of a long struggle in many coalfields to mine with power tools safely and to increase output per man-shift. Attempts to use power to loosen coal and to transport broken coal in the U. S. may be said to date from the Stanly Header, or Entry-Driver brought from England in 1888. The principle of this boring machine was used in the McKinlay Entry-Driver, almost continuously used in the U. S. since 1920. Since 1888 experimental loaders, scrapers, and conveyors were installed with more or less success. Beginning in 1918, considerable progress was made with mobile loaders, and in 1920 the first wage agreement for the operation of mechanical loaders was made in Indiana. In 1923 the Pocahontas Fuel Co. loaded nearly 1 million tons of coal using 23 Coloders. Labor Policy The United Mine Workers of America have never officially opposed the Mechanization movement. On December 10, 1945, in a statement before the House of Representatives Labor Committee, John L. Lewis said, regarding the policy of the United Mine Workers of America: "We have welcomed progress; we have welcomed machines. We have told our people that they had to accept that condition; that it was the process of progress, and that they would have to take their chances." Recent Development In 1951 about 71 pct of the tonnage mined underground was mechanically loaded. Over 4000 shuttle cars are in service and it is estimated that much more than half of the tonnage loaded underground is produced by trackless mining. In 1947 roof-bolting was introduced extensively and it is estimated that more than 2 ½ million bolts are now being used per month in about 600 mines. The use of roof bolts has permitted the more effective and safer use of loaders and shuttle-cars. Continuous Mining The McKinlay Entry-Driver could have been used for continuous mining, but for many years it was used only .for entry driving. In 1946 the Silver machine was developed in Colorado, and in 1947 this was acquired by the Joy Mfg. Co. and called the Joy Continuous Miner. Other types of combination mining-loading machines which eliminate drilling and blasting operations are the Marietta Miner, the Colmol, the Lee-Norse Miner, the Junior Miner, the Goodman and the Konnerth machines. There are several other types in the process of development. Probably by January 1953 there will be about 250 continuous miners in operation in the U. S. Payment of Mine Labor One of the most important problems in mechanizing was the establishing of rates of pay that would be attractive to the best men. Prior to the installation of mobile loaders the hand loaders and cutters have been paid by the ton. It was felt that a system of a day's pay should be established and that piece and tonnage rates should be abolished completely. Without exception, all coal loaded in mines equipped with mobile loaders is prepared and loaded by men paid an hourly rate. Trends in Mining Research There are two diverse approaches in coal production research; to try to design a mining machine to fit current methods, or to adapt mining practice to take advantage of proven machines. A great deal of credit must go to the mine operators who have been enterprising and dynamic enough to use available equipment intensively and to discard it as soon as an improved or new machine is available. Effect of Changing Markets Formerly there was an important demand for lump and prepared coal in the larger sizes for domestic use. The use of bituminous coal for house heating has decreased so that not more than 19 pct of the annual tonnage goes to retail trade, only 12 pct is used by the railroads, and much of the retail and most of the railroad coal is in stoker size. As a result of this decline in the market for coarse coal most of the large mining operations have crushers installed in the tipples or preparation plants. Mass production at the face requires increased preparation facilities. Another important trend is in connection with complete seam mining with mechanical loading. In many mines where the immediate roof is apt to fall with the coal no effort is made to separate roof material from coal prior to

Jan 1, 1952

By C. N. Garman

Homestake-New Mexico Partners operate a 750-tpd carbonate leach uranium concentrate mill near Grants, N.M. The highly mineralized water available as process water leaves much to be desired. The 628 ppm as CaCO 3 makes the use of raw water very troublesome in pipes and on filter cloths. However, the residual sodium carbonate in the final filter cake going to tails makes an ideal softening agent. To take advantage of this fact, all makeup water used in the mill is first used as tailing slurry dilution water and comes to the mill from the tailings pond. The 5-acre tailings pond serves as a thickener and 100 to 150 gpm of nearly clear solution is decanted to a pump to be returned to the mill. Since this tailings water has small quantities of uranium in the solution an ion exchange scavenger unit was installed to remove as much uranium as possible. The ion exchange raffinate is then used as final filter wash ahead of the tailings slurrying step. In spite of the large settling area this return water is not clean enough for ion exchange feed. The solids present are very fine and composed of approximately 15 pct (by weight) burnable carbonaceous material common to the sandstone uranium ores in the area, 40 pct SiOz plus 45 pct CaC03. Laboratory work showed that this material responds very well to flotation. Before deciding to use flotation, various clarifying systems such as pressure leaf filters, sand filters, and continuous vacuum pre-coat filters, were considered. Each of these could have solved the problem but with much more operating labor, more reagents and greater installation costs than the flotation step. About 100 to 150 gpm of fouled water is fed to two 66-in. Fagergren cells, in series. Reagents used at the beginning were Arquad 2HT75 and Arquad C50, at the rate of about 1% lb per 8-hr shift, or about 0.0053 lb each per ton of ore. This did not completely remove the solids but does an acceptable job. Approximately 75 pct of the slimes are a size that can be caught on a 41-Whatman Paper are removed. Removal of these slimes also allows much better settling of the coarse nonfloatable material. Advantage is also taken of this fact in a small settling tank ahead of precipitation. Removal of this amount of the slimes makes the ion t:xchange feasible. PREGNANT SOLUTION CIRCUIT The carbonate? leach-caustic precipitation method of uranium concertration does not provide for any process purification step ahead of precipitation. Therefore, any fine solids getting into the pregnant solution through the filter cloth show up in the final concentrate. This, of course, lowers the grade, and, at times, the slimy nature of these very fine solids rendered final filtration of the concentrate difficult if not impossible At Homestake-New Mexico Partners a 75-ft thickener was available for gravity clarification of 100 to 120 gpm of this pregnant solution. However this did not sufficiently remove the slimes. Laboratory investigation of the whole range of flocculants that were suggested by literature, salesmen, and friends failed to turn up anything of consequence. A continuous vacuum pre-coat filter would do the job and was investigated. The capital cost and the operating labor and materials made this a last chance choice. Following work done in the metallurgical laboratory on the tailings return water, it was found that some changes in the reagent strengths and combinations made a very definite decrease in the solids in the pregnant solution. Concentrate grade improved about 5 pct anti the final product after drying had an appreciably greater bulk density. Compared to a cost of about 2.2e per ton for pre-coat filter opelation for cleaning just one circuit, flotation costs less than 1.0 per ton of ore for cleaning two circuits. While a pre-coat filter would do a more thorough job, the flotation does all that is required for either circuit. Gravity causes the froth produced to run back into the leach circuit. This does not appear to result in a build-up of objectionable slime. No extra manpower is required; the operators in the separate areas can observe the operation of the cells and mix the small quantities of reagents as needed. Normally the 66-in. Fagergren cell requires 15 hp per cell, but this very dilute slurry needs only 10 hp for both cells. Originally, a combination of the two Arquads mentioned previously served as frothers and promoters. As further testing

Jan 1, 1962

By R. Wild, F. A. Wright

When a blast furnace has a certain defined burden and is operated under fixed conditions of blast temperature, etc., the fuel efficiency is determined by the extent to which the reducing gases can remove oxygen from the burden in the furnace stack. This is determined by two distinct factors: 1) The uniformity of gas-solid contact, and 2) The ease with which oxygen can be removed from individual pieces of burden. This latter is often called burden reducibility. When burdens were poorly prepared the first factor was by far the most important and a study of the reducibility of individual lumps was of rather academic interest. In recent years good burden preparation with emphasis on uniformly sized material has led to greatly improved gas distribution in the stack, and thus the second factor has become much more important and there has been a marked increase in interest in methods of measuring reducibility. This paper explores the Linder method of measuring such reducibility. The measurement of reducibility of burden materials must be carried out under conditions duplicating, as nearly as possible, those of the blast furnace stack. This is very difficult since the blast furnace process is a counter-current one, and thus the initial conditions encountered by the solid (gas temperature, composition, etc.) are the result of heat and mass transfer occurring lower down the stack. Any method of burden testing which does not take this into account is, at least to some extent, based on arbitrary assumptions. In an attempt to study blast furnace reactions under non-arbitrary conditions BISRA adopted the SCICE technique as a method of investigation. This technique has been used with a measure of success.' The SCICE technique, however, was found to be too slow for use as a routine test for burden materials and it was decided to construct additional equipment for burden testing. A test was required which would: 1) Be as realistic as possible. 2) Be quick and easy to operate. 3) Give some indication of the breakdown likely to take place during reduction in addition to a reducibility index. After a critical assessment of the reducibility tests which have been proposed it was decided to adopt the Linder test equipment and procedure as a basis for burden testing. THE LINDER TEST APPARATUS AND PROCEDURE The apparatus which was constructed (Fig. 1) was the same as that described by Linder2 except for minor changes in design. Linder also laid down a test procedure which he had derived from the results of investigations on Swedish blast furnaces. The variations of temperature and gas composition during the test were defined; these are shown diagram-matically in Fig. 2. BISRA's intention was to use the standard temperature and gas composition programmes for testing a variety of burden materials and also to investigate the influence of different programmes on standard burden materials, making use of information from the SCICE apparatus wherever this is possible. Up to the present, effort has been concentrated on the first part of the programme, and work on the second part has only just commenced. For each test 200 g of coke and 500 g of burden material were used. Linder had recommended that the coke and burden material should be between 1 and 1 1/2 in. and this was adhered to in early experiments on ores and sinters. Since the eventual aim of this work was to relate the test results to blast furnace operation, it was decided to carry out subsequent experiments using burden material in the size range used in the blast furnace, as far as this was possible. If the main interest was, for example, a comparison of the products resulting from different methods of agglomeration, then there would be advantages in using burden materials as close as possible to a standard size. After the charge had been placed in the reaction tube and this had been connected to the gas supply, rotation of the reaction tube at 30 rpm was started and the reduction programme was commenced, the gas temperature and composition being manually controlled according to the programme shown in Fig. 2. After the reduction test the charge was cooled in a nitrogen atmosphere. It was then removed from the reaction tube, the coke and burden separated, and the extent of burden breakdown assessed by screening it at 10 and 30 mesh. The extent of reduction was then

Jan 1, 1964

By K R. Brentley

"Tshepong Mine, Harmony’s largest gold producing mine, is situated in the Free State portion of the Witwatersrand Basin, east of Odendaalsrus and north of Welkom, South Africa. The Basal reef is a narrow tabular conglomerate and is the main gold bearing ore body. The strata strikes mainly in a northsouth direction with an average dip of 22 degrees. Currently mining takes place at depths of between 1 690m and 2 150m below surface. The Basal reef unit consists of a conglomerate ranging between a few and 30cm in thickness. Some 80-90% of the gold occurs in the lower portion of the conglomerate. Directly above the conglomerate is a quartzite placer unit, the Basal quartzite, which is a clean, hard competent rock and varies in thickness between 50cm and 200cm. The Basal unit is overlain by the Khaki shale which can vary between 30cm and 500cm in width. It is weak and incompetent khaki coloured chloritic shale.A method of stoping, termed “undercut”, is practiced in order to remove the lower 15-20cm of the gold rich strata. The method entails removing this portion along with some 70-80cm of footwall host rock which is a strong competent quartzite. The hard competent quartzite overlying the conglomerate is left in-situ to form a competent hangingwall which can stabilize and prevent exposure of the weak Khakishale. This method of extraction was developed through trial and error and has been used reasonably successfully since its inception in the early 1960’s where the method was developed to extract the Basal reef at the Western Holdings Mines. Recently prospect drilling and a number of localised hangingwall failures indicated that the residual quartzite beam could be less than 60cm and in some instances as thin as 30cm in certain areas of the mine. The result of this is a less stable hangingwall which not only poses a safety risk to workers in these areas but also had the potential of locking up some 840 000 Tonnes of gold bearing ore. At the present gold price this amounts to revenue of approximately ZAR1.5 billion (U.S. $143 million).This paper discusses the method of “undercut stoping” and existing practices in place at Tshepong Mine. It also explores new initiatives such as the use of a simple freely-supported elastic beam solution to perform back analysis, on-site data capturing of beam thickness and support geometries in order to develop empirical design charts and numerical modelling to assist in the verification of the empirical design method."

Jan 1, 2015

By Éder Domingos Oliveira, Leandro Viana Teixeira, Carlos Antonio Morais, Ruberlan Gomes Silva

The purity of rare earth products depends on the ability of the reagents used to selectively promote the precipitation of the rare earth elements (REE) from impurities present in sulfuric liquors. The present study aims at investigating the conditions required to precipitate REE, either as rare earth oxalates or carbonates, from purified rare earth sulfuric liquor with a low Al3+ and UO2 2+ and a very high Ca2+, Mg2+, Mn2+, and SO4 2− impurity content. The rare earth sulfuric liquor was obtained in a previous study by the reaction of a rare earth ore with sulfuric acid, pyrolysis at 700 °C/2 h, cooling at 20 °C, and water leaching. The impurities such as Fe3+, Th4+, and PO4 3− were completely removed from the liquor in two consecutive neutralization steps. In order to recover the REE, the purified rare earth liquor was then treated with oxalic acid and sodium carbonate. Rare earth oxides (REO) precipitated with oxalic acid displayed a very high purity (99.2% w/w) and low impurity content. In contrast, the rare earth carbonates precipitated with sodium carbonate presented a comparatively lower REO content (68.5% w/w) and 94.2% of purity, but with a higher content of impurities. Better precipitating conditions were achieved at 60 °C, compared with 20 °C, and under stoichiometric reagent consumption, with REE precipitation efficiencies higher than 96%. The dosage above the stoichiometric condition caused the precipitation of calcium oxalates and co-precipitation of rare earth sulfates when oxalic acid is used, and precipitation of calcium, manganese, and uranium carbonates and also a co-precipitation of rare earth sulfate when sodium carbonate were used. An integrated flowsheet to produce a high-quality rare earth product from rare earth sulfuric liquor containing some impurities is proposed.

By F. R. Archibald, C. M. Nicholson

The present paper may be considered complementary to an earlier contribution on the same subject by F. R. Archibald and C. F. Jackson.1 It is particularly concerned with engineering and technological aspects of the operations at the Harleyville Alumina Plant during its period of operation from January 1945 to July 1946. The plant was built for and operated by Ancor Corporation, at Harleyville, South Carolina, as part of the government sponsored wartime program related to alumina production from domestic raw materials. The main purpose of the program was to secure technical and engineering data, based on actual plant experience, that might serve as a foundation for large scale production of alumina from raw materials other than bauxite. The Harleyville plant was directed specifically to treatment of clays found in very extensive deposits in the Carolinas and Georgia using the lime-sinter process developed by Ancor Corporation. It is considered that the purpose stated above has been served and this paper is to record and discuss some of the more important findings. The main factor in location of the plant at Harleyville was the presence in that locality of a rather remarkable deposit of high calcium limestone. The clay mining operation was centered near Aiken, S. C., and although a rail haul of ninety miles was involved there were, at the time, factors favorable to the Harleyville location offsetting, in part, the cost of transportation of clay, There are, no doubt, locations in the general area where clay and limestone deposits may be found in closer proximity hut they had not been proven up when the project was undertaken. Availability of power from the Santee-Copper development whose high tension line skirted the Property; ready access to rail and highway transportation; the promise, realized in actuality, of ample supply of soft water; and availability of labor supply were other factors of importance. Moreover, the general area has the decided advantage, for this type of operation, of a mild climate. Construction was begun in August 1943, and although stopped for a time priding review of the general situation that fall, it was resumed, albeit with less urgency, and substantially completed by the end of 1944. The total capital cost was approximately $3,460,000 The first alumina was turned out in April 1945 and from then until the plant was closed down in July 1946, the operations followed a succession of periods of production alternating with periods in which minor alterations were made in conformity with experience. Steady improvement was made and at the time Operations were discontinued it was considered that experience and the necessary engineering were in shape for removal of the remaining bottlenecks and a somewhat major revision in one section of the plant. Unfortunately this could not be carried out and it was necessary to discontinue the

Jan 1, 1949

By W. W. Schmittroth, A. Y. Bethune

THE commercial analysis of Special High-Grade zinc usually involves the determination of lead, iron, copper and cadmium as impurities in the base metal. Over the past 20 years, as the result of metallurgical improvements in the art of producing Special High-Grade zinc metal there has taken place a gradual exclusion of almost all extraneous elements, and the few that do remain in the final product have been virtually reduced to the vanishing point. Thousands of tons of zinc .are now commercially produced annually and made available to the trade, in which impurities are present only in the order of 0.0010 per cent or less and the sum total of all impurities does not exceed 0.0025 per cent. To such zinc the term "superpurity" might well be applied without undue exaggeration. The techniques used for production control and final analysis of Special High-Grade zinc by means of emission spectra in the laboratories of the electrolytic zinc plant of the Sullivan Mining Co. are based on the premise that self-electrodes are to be preferred over all other forms for the examination of the zinc in question. Furthermore, after a survey of the literature and considerable experimental work,1 it was decided that the spectrographic method of internal control, the principle of which was first announced by Gerlach,2 using a modification of the high-voltage alternating-current arc source introduced by Duffendack and Thompson,3 with the calibration of each photographic plate as recommended by Wolfe,4 would yield the most accurate and reproducible results. Self-electrodes were chosen in order to make use of one of the unique advantages inherent in spectrographic analysis; namely, the rapid and almost effortless detection of elements other than those being quantified by the analyst; or conversely, the proof that the zinc is free from such impurities if no spectral line of the suspected element appears. This fact was verified for certain elements, and will be discussed later. Carbon electrodes impregnated or packed with zinc salts chemically prepared from the zinc to be assayed were avoided for a number of reasons. First, the chemical manipulation was another time-consuming step, which increased the chances for contamination. Second, none of the common zinc salts, such as the chloride, nitrate or sulphate, yielded reproducible results under our conditions, nor did the spectral line of the impurity appear as readily in the low concentrations as when using zinc metal. Third, no commercial carbon, except the very expensive, was pure enough to use with this grade of zinc and the continued time, effort and expense incurred in the purification of carbon was not desirable. The use of Special High-Grade zinc self-electrodes introduces certain complications with regard to obtaining satisfactory excitation conditions, because it seems necessary to raise the temperature

Jan 1, 1946

By J Wu

Swirl Flow Technology (SFT), a novel approach for agitation in large industrial slurry tanks, was jointly developed and patented by CSIRO and Queensland Alumina Ltd (QAL). Currently, SFT is used in 16 alumina precipitation tanks at QAL, with the first installation in January 1997. This paper provides insight to the underlying fluid dynamics, which were investigated through pilot-scale experiments. A key finding was that SFT produces tank wall velocities approximately four times that of conventional draft tube agitator systems for the same power input, whilst for solids suspension, much less power is needed. SFT is energy efficient and practical in scaling suppression as demonstrated by the full-scale experience at QAL.

Aug 8, 2011