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By D. Turnbull, J. C. Fisher

ACCORDING to the "reaction-path" modell,2 of martensite nucleation, the shear angle of the embryonic martensite plate must be treated as a variable, and included in any calculation of nucleus critical size. Also, as can be deduced from this model, the interfacial free energy between austenite and martensite does not reach its final value until the shear is completed. It is zero for zero shear angle. However, in order to account for the kinetics of the martensite transformation, some sort of interfacial energy barrier appears to be necessary even with the reaction-path model, for otherwise the volume and the energy of formation of the critical size nucleus both collapse to zero.3 Cohen independently suggested that surface energy could be incorporated into the reaction-path model, with the overall free energy of a martensite embryo being a function of its volume and shear angle.' It is possible to estimate the energy associated with the formation of a critical-size martensite nucleus starting with the reaction-path model and including a surface free-energy barrier. As the dependence of interfacial free energy upon shear angle is unknown, a simple type of dependence will be assumed, with the belief that the true dependence would not lead to appreciably different results. Consider the work required to form a lenticular martensite plate with radius r, thickness t, and shear angle 8. There are three contributions; one being the interfacial free energy, one being the free energy change in the martensite plate, and one being the free energy increase in the surrounding austenite. The interfacial free energy u is assumed to depend upon the shear angle 0 according to the relationship s=s0(?/?0)n [1] where 8, is the equilibrium shear angle and n is an exponent that may lie in the range 0 n 2. The work required to form the interfaces of a martensite plate then is W. = 2pr² s0(?/?0)n [2] The free energy change per unit volume of martensite is composed of two parts, one the ordinary volume free energy ?f1. which is negative, and the other the elastic strain energy G?m²/2, where G is the shear modulus and 7, the shear strain relative to the martensite structure. This expression for the strain energy is valid only when the shear strain ym, is sufficiently small that the martensite is within its linear elastic range. There is no doubt that ym, lies beyond the linear elastic range for embryos that are considerably subcritical. However, for critical nuclei it will be shown that ym, is 1.5 pct or less, within the linear elastic range of martensite. For embryos of nearly critical size, then, the strain energy of the martensite is correctly given by G?m²/2. The shear strain in the martensite is ym, = 8, — 8, and the work required to form the strained martensite is Wm --= (pr²t/2) [?fv + G(?O - ?)²/2] [3] The free energy change in the austenite is entirely that due to elastic distortion. The elastic strain is not uniformly distributed in the austenite, being large near the martensite plate and small elsewhere. Approximately, however, the energy corresponds to a uniform shear strain ya= (?t/2)/r [4] throughout the volume 4pr³/3 surrounding the plate. The work required to strain the surrounding austenite then is Wa = (4pr³/3) (G?a²/2) = (G?²/6) prt² [51 For simplicity, the same shear modulus G is assumed for each structure. The total free energy for forming a plate then is W = W3 + Wm + Wa. = 2pr² s0 (?/p?0)n + (pr²t/2) [?fr+G(?0-?)²/2] + (G?²6) prt2 [6] This expression is correct for nuclei and for embryos of nearly critical size, where, as will be shown, the strain energy in the martensite is correctly given by the expression G (? — ?)². Having W as a function of r, t, and 8, as in Eq. 6, there is a saddle-point where W has a stationary value, W subsequently decreasing indefinitely as the nucleus volume increases along the reaction path. The stationary value of W is the energy of the critical nucleus. The critical nucleus has radius, thickness, and shear angle such that ?W/?r - awlat: = ?W/?p? = 0. Performing these differentiations and calculating the critical nucleus energy, W* = [8192p(G?/6)²;s/27 ?fv4] [7] where a= (?/?0)3n+1[l +G(8"-8)'/2af.]' [7a] and where 8 is to be determined from the equation (1 + 3n/4) + G8(6O - (9)/[Af. +G(6>o-6>)72] = 0 [8] For ?f, near —200 cal per mol or —10" ergs per cc, and 8, near 1/6, as for iron-base alloys, Eq. 8 gives ?0 - ? ~ - (4 + 3n) ?f1./4G0O [9] as the difference between the equilibrium shear angle and the actual shear angle for a critical nu-

Jan 1, 1954

By Glenn A. Shirey

The objective of this contract as to prepare a detailed, readily usable handbook on dust control techniques for longwall mining operations. Information was gathered from available literature, and through contacts with industry and government personnel. The handbook describes those methods of dust control applicable to today's longwall mining systems, and contains sufficient design elements to assist operators with problems in testing or implementing new control procedures. Complete reference lists are presented at the end of each chapter to direct the reader to additional information.

Jan 1, 1985

By Weng Fu, James Vaughan, Stefan Lakemond

Uranium bearing copper sulphide flotation concentrates produced from iron oxide copper gold uranium (IOCG–U) ore contains radionuclides such as 238U, 230Th, 210Pb, 226Ra, 210Po, which have the potential to reduce their marketability, thus posing a challenge to the industry. Thus, a two–stage leach process has been proposed in this study. In first stage, a sulphuric acid leach process effectively dissolve most of the 238U and 230Th into solution, but other radionulclides (mainly 210Pb, 226Ra and 210Po) remain associated with copper sulphide minerals owing to their low solubilities in sulphuric acid solution. In the second leaching stage, the use of methanesulfonic acid (MSA, CH3SO3H) is investigated to selectively dissolve the 210Pb, 226Ra and 210Po from the residue of the first stage. The results indicate that 96% of 238U, 87% of 230Th, 90% of 210Pb, 86% of 226Ra and 17% of 210Po are dissolved using 3 mol/L MSA solution at 90°C for 3 hours, which represents a dramatic decrease in the radioactivity of the copper concentrate. INTRODUCTION Australia is host to one of the largest iron oxide copper–gold–uranium (IOCG–U) deposits in the world. Most of the value in these deposits is associated with copper. In mineral processing, the gangue minerals are separated from copper sulphide minerals that report to a copper flotation concentrate. However, due to the strong association of uranium with copper minerals and hematite, even clean copper concentrates can contain significant amounts of uranium (238U) and its radioactive decay elements, of which thorium (230Th), radium (226Ra), lead (210Pb), and polonium (210Po) are the long life radionuclides. These hazardous radionuclides negatively affect the marketability of the copper concentrates. Customer copper smelters in China, Japan and Europe may consider the radionuclides penalty elements or may not be able to process the feed. Therefore, there is a strong motivation to remove radionuclides from copper concentrates at the mine site before shipping to custom smelters in global market. Hydrometallurgical leaching has been widely employed to remove radionuclides from different kinds of minerals such as copper concentrates (Ragozzini & Sparrow, 1987), zircon concentrates (Aral, Sparrow, McDonald, & Norgate, 2007), and coal ash (Lei, Qi, Sun, Xu, & Wang, 2014). Leaching using sulphuric acid (H2SO4) solutions is the most widely used process for the extraction of uranium from uranium ores due to the relatively low cost and wide availability of the acid (Bhargava et al., 2015; Gilligan & Nikoloski, 2015). The leach process involves treating ground ore slurry with sulphuric acid at elevated temperatures (typically 60–90 °C) under oxidising conditions in order to convert the uranium in the minerals 2+ from the relatively insoluble tetravalent form (UO2) into water–soluble hexavalent state (UO2), which then complexes with SO42- and forms UO2(SO4)22- in solution. The uranium oxidation is generally carried out in the presence of ferric ion (Fe3+), regardless of the reagent and oxidant used. The most commonly used oxidants for uranium include sodium chlorate, pyrolusite and oxygen (Ring, 1980). These oxidants act indirectly by converting ferrous sulphate, formed by the oxidation reaction of UO2 in slurries, into ferric sulphate. The presence of Fe–containing gangue minerals such as hematite in the ore usually provides sufficient Fe required for the oxidation of UO2. After sulphuric acid leaching, the locked thorium (230Th), radium (226Ra), lead (210Pb), and polonium (210Po) elements are chemically liberated from hematite and uranium mineral grains (Ciobanu, Wade, Cook, Schmidt Mumm, & Giles, 2013). However, the extraction rate of radium (226Ra), lead (210Pb), and polonium (210Po) in sulphate solution is very low due to the limited solubilities of (Ra, Pb, Po)SO4 (Bagnall & Freeman, 1956; Liu & Papangelakis, 2006; Nikitin & Tolmachev, 1933).

Jan 1, 2019

By R. W. Schultz

Mineral exploration carried out in Ireland since the mid-1950s has resulted in the discovery of ore deposits that are now in production atTynagh (Pb-Zn-Cu-Ag), Silvermines (Zn-Pb) and nearby Ballynoe (barite), and Gortdrum (Cu-Ag-Hg); a major Zn-Pb deposit at Navan that is now being drilled prior to development; potentially productive deposits at Keel (Zn-Cd-Pb), Ballinalack (Zn-Pb) and Aherlow (Cu-Ag); and a number of interesting prospects, including one of spodumene. Underground exploration of a formerly extensively mined copper deposit at Allihies has outlined new, albeit currently uneconomic, reserves at depth. The copper and pyrite deposits at Avoca have in recent years been re-explored and are now again being worked. Most Irish ore deposits are in the lower half of the Lower Carboniferous succession of limestones and subordinate shales and sandstones, and this stratigraphic zone is being most intensively explored at present. Other geological units under active exploration are Ordovician sediments-volcanics belts, Caledonian granites, and parts of the Old Red Sandstone and Dalradian series. Surface prospecting is made difficult in many places by glacial and other superficial deposits, including peat and alluvium. Geochemical prospecting methods have been outstandingly successful and primarily responsible for most new discoveries. Boulder hunting and outcrop prospecting have proved effective in detailed investigations of geochemically anomalous areas. Some parts of the country have been covered by three separate airborne geophysical surveys, albeit without success so far. The induced polarisation technique, the most widely used ground geophysical method, has contributed to several discoveries. The scope of application of electromagnetic surveys is more restricted than that of induced polarisation, but the method has been effective in a few cases and has contributed to the Tynagh discovery. To date, the resistivity and self-potential methods have found little application. Magnetic surveys are not applicable directly in the search for ore deposits of the various known types, but can be helpful in outlining geological structures in some areas. The gravity method has been used successfully in a survey over the Zn—Pb deposit at Navan ; it has also proved effective in trials over a blind sulphide deposit at Tynagh that cannot be detected with geoelectrical methods. An appendix lists summary descriptions of 15 Irish deposits and occurrences

Dec 1, 1971

By Bruno Gomes Silva

O Distribuidor é um equipamento indispensável e de grande importância no processo de solidificação do aço via Lingotamento Contínuo. Os distribuidores são revestidos por um material refratário de trabalho que entrará em contato com o aço líquido. Atualmente é utilizado um revestimento que possui a necessidade de adição de água durante a aplicação, que poderá ser manual ou automática. Com o distribuidor disponível para lingotamento, o mesmo é direcionado para as máquinas de lingotamento onde serão pré-aquecidos visando retirar o excesso de umidade e promover a sinterização do refratário. O presente trabalho apresenta uma alternativa de revestimento, onde não é utilizada água na aplicação do produto, ou seja, um refratário com características de auto-endurecimento que permite iniciar o lingotamento sem haver necessidade de pré-aquecimento, o que significa partir a frio com o distribuidor.

Aug 16, 2017

By I. R. Kramer

The stress associated with the high-dislocation layer at the surface of deformed aluminum crystals was measured by progressively polishing the specimen and determining the change in the initial flow stress when the specimen was reloaded. This surface stress increased linearly with strain in the Stage II and Stage III region. The depth to which the high-dislocation layer extends was found to be 0.0025 in. The influence of the surface layer on Stage I is discussed and it is shown that Stage I ends when, on the secondary slip system, the difference between the applied stress and the opposing stress due to the surface layer is equal to the critical resolved shear stress. In previous papers'-3 it was shown that the surface exerts a pronounced effect on the plastic-flow characteristics of metals. For fcc metals it was possible to increase the extent and decrease the slope of Stages I and II and decrease the work-hardening coefficient of Stage II by removing the surface of the specimen continuously during the deformation process.2,3 More recently,L it was shown that the activation energy for plastic deformation and the activated volume were also influenced by the surface. These changes were related to the existence of a high concentration of dislocations in the region near the surface of the deformed specimen and it was concluded' that the effective stress, T, acting on a dislocation was T = Ta-Ti-Ts [1] where 7, is the applied shear stress, Tj is the internal stress, and TS is the stress associated with the surface region containing the high concentration of dislocations. From Eq. [I] it is seen that the effective stress, 7, acting on a dislocation is a function of 7,. Therefore 7, becomes important in any description of the work-hardened state. In this paper the value of 7, as a function of strain and its variation with depth are reported for aluminum single crystals. In particular, it will be shown that the end of Stage I occurs when the net stress, 7, on the secondary slip system is equal to the critical resolved shear stress. To measure Ts, single crystals of aluminum were deformed in tension to various strains, whereupon the load was removed and a known amount of metal was removed by electrolytic polishing. The differ- ence between the final flow stress before the specimen was unloaded and the initial flow stress upon reloading is designated as ?Tp and is the decrease in the stress due to the removal of the surface layer in whole or in part. The value for Tp is taken equal to -Ts. EXPERIMENT PROCEDURE The details of the tensile apparatus, method of electrolytic polishing, and crystal preparation were the same as those used previously,1"3 The single crystals were 1/8 by 1/8 in. in cross section with a 3-in. gage length. The initial purity of the aluminum was 99.997 pet. The specimens were held in vacuo at temperatures 50°C below their melting points for 16 hr and furnace-cooled. Just prior to the application of the tensile stress the specimens were electrolytically polished to remove 0.004 in. from the thickness. This latter operation was conducted after the specimen was placed in the tensile apparatus. The tensile tests were conducted at a strain rate, i, of 10"5 sec-' and a temperature of 3°C. The aluminum crystals used for the determination of 7, are designated A1-3-12 and A1-116. The orientation of the various crystals used throughout the investigation is shown in Fig. 1. The ?Tp values for specimens A1-3-12 as a function of Ax were determined mainly at four strain levels, i.e., ? = 0.0217, 0.0381, 0.0593, and 0.075, where Ax is one half of the total reduction of the thickness after polishing and ? is the re-

Jan 1, 1965

By Frederick Laist

The ores received at the Washoe Smelter come almost entirely from the mines in Butte and contain the following minerals : Chalcocite, Cu2S; covellite, CuS; chalcopyrite, CuFeS2, (trace); bornite, Cu3FeS3; enargite, Cu3AsS4; malachite, Cu2CO4 + H2O (trace); pyrite, FeS2; sphalerite, ZnS; rhodo-chrosite, MnCO3, (less than 1 per cent.); rhodonite, MnSiO3, (less than I per cent.). The country-rock of Butte is of the granite family and is a quartz-monzonite containing a little less silica than typical granite. It is made up of the following rock-minerals in approximately the proportions indicated: Minerals. Composition. Per Cent. Quartz SiO2 20 Andesine 1/2 NaAlSi3O8; 4 CaAI2Si2O8 42 Orthoclase KA1Si308 23 Hornblende 1/2 Ca(MgFe)3Si4O12; & (MgFe)2(AlFe)4Si2O12 6 Biotite (HK)2(MgFe)2A12Si3O12 9 Titanite CaTiSiO5 0.1 Apatite (CaF)Ca4(PO4)3 0.05 Magnetite (FeO)(Fe2O3 0.15 Pyrite FeS2 0.02 99.32 The main country-rock (quartz-monzonite) is cut by dikes of aplite, a granite having the following composition: Quartz, 30; andesine, 20; orthoclase, 45; ferro-rnagnesian minerals, 3; pyrite, 0.1; total, 98.1 per cent. Closely associated with the copper-deposits are dikes and bodies of " Modoe" porphyry, technically known as quartz-

Jan 1, 1913

By D. Grobbelaar, A. Cichowicz, D. J. Birch

The Q-coda method for estimating the quality factor Q(f)= Qo(f)n was used to characterize seismic wave attenuation in a region of the Kaapvaal Craton that includes the mining areas of the Bushveld Complex and Witwatersrand Basin. Seismic waveform data, collected by locally distant stations of the South African National Seismograph Network (SANSN), consisted of mining-related events with magnitudes ranging from ML 1.8 to ML 4. Q was calculated for nine different source-receiver pairs spanning the study region. A weighted average Q based on the number of available data gave an estimated attenuation relation for the study region of Q(f) = 327 f 0.81.

Jan 1, 2015

By W. Scheetz

The Brewer Gold Mine is an open-pit, heap leach operation that processes about 122 kt/m (135,000 stpm). The mine is operated by Brewer Gold Inc., a subsidiary of Westmont Mining Inc., and is located near Jefferson, SC, about 80 km (50 miles) southeast of Charlotte, NC and 80 km (50 miles) northeast of Columbia, SC (Fig. 1). All of the ore to date has been produced from a breccia body of hydrothermal origin and a related smaller body of fault-controlled ore, known as the B-6. Reserves at the time the mine commenced production were 4.6 Mt (5.1 million st) grading 1.4 g/t (0.042 oz per st) gold. History Although some reports of a pre-Revolution War shaft at the Brewer are in the literature (Lieber, 1856, Nitze and Wilkens, 1897), gold was first produced in earnest from the Brewer Mine in 1828. Production occurred primarily from placer deposits south of the present mine workings. Lieber noted in 1856 that “every branch, gulch, ravine or even declivity of the main hill has been washed and rewashed.”

Jan 1, 1991

The Framework Convention on Climate Change (FCCC) places serious obligations on New Zealand as a signatory nation. This paper examines the nature of the FCCC, New Zealand 's proposed policy response, and the.implications for the minerals industry.

Jan 1, 1994

By J. E. Gould, H. J. Hand

Many independent studies are being made on slag-metal relationships in the open-hearth furnace, and these studies cannot help but result in an ultimate improvement in the quality of open-hearth steel of both the open and killed types. In maintaining quality control at the National Tube Company's plant, emphasis has been placed on an evaluation of the iron oxide content of the steel in as quantitative a manner as possible, in order that the most desirable level of oxidation might be obtained for a given type of steel, and to secure a more precise basis for final de-oxidation. In the past, various methods were used for estimating FeO content of the metal at tap to ensure uniform and thorough deoxidation of killed steel, avoiding overdeoxidation to minimize the amount of inclusions present and to make as sound a steel as possible. For the open-type steels, such control involves obtaining the necessary level of oxidation for the most suitable action of the steel in the molds to provide good surface quality and cross section for the specific grade involved. While rapid analytical methods are available for determination of FeO in the liquid metal, such methods are still not sufficiently rapid for routine control on the open-hearth floor, nor do such analyses provide information as to what causes high or low FeO values. The method of sampling can still be considered a problem in view of the variability encountered in taking simultaneous duplicate tests. With the accurate rapid methods now available for determining carbon in the metal, estimating slag basicity and obtaining temperatures of liquid metal, it was believed possible to determine the influence of the various furnace reactions in establishiug the state of oxidation at a given time. Simultaneous slag and metal tests were used in this analysis, aided by statistical methods, to determine the factors that affect the metal FeO, and also to make a quantitative evaluation of these factors. Data obtained are from actual iurnace operations, the metal test being of the I spoon, Herty type, and taken in duplicate. Metal temperatures were taken with the open-tube bath pyrometer developed by the Research Laboratory of the United States I Steel Corporation. This method of taking open-hearth metal temperature was invented by Collins and Oseland,l and was developed into its present form by L. 0. Sordahl,2 of the Research Laboratory. The instrument and its mode of operation have been fully described by Sordahl and Sosman in three publications.3,4,5 lag components were obtained by chemical analysis. Data were taken from four 165-net-ton furnaces during the working period of the heat. Recognizing that additions of ore affect the metal FeO. no tests were included within 30 min. after ore additions. In all, 65 tests (48 heats) were used in this study.

Jan 1, 1942

By J. E. Gould, H. J. Hand

Many independent studies are being made on slag-metal relationships in the open-hearth furnace, and these studies cannot help but result in an ultimate improvement in the quality of open-hearth steel of both the open and killed types. In maintaining quality control at the National Tube Company's plant, emphasis has been placed on an evaluation of the iron oxide content of the steel in as quantitative a manner as possible, in order that the most desirable level of oxidation might be obtained for a given type of steel, and to secure a more precise basis for final de-oxidation. In the past, various methods were used for estimating FeO content of the metal at tap to ensure uniform and thorough deoxidation of killed steel, avoiding overdeoxidation to minimize the amount of inclusions present and to make as sound a steel as possible. For the open-type steels, such control involves obtaining the necessary level of oxidation for the most suitable action of the steel in the molds to provide good surface quality and cross section for the specific grade involved. While rapid analytical methods are available for determination of FeO in the liquid metal, such methods are still not sufficiently rapid for routine control on the open-hearth floor, nor do such analyses provide information as to what causes high or low FeO values. The method of sampling can still be considered a problem in view of the variability encountered in taking simultaneous duplicate tests. With the accurate rapid methods now available for determining carbon in the metal, estimating slag basicity and obtaining temperatures of liquid metal, it was believed possible to determine the influence of the various furnace reactions in establishiug the state of oxidation at a given time. Simultaneous slag and metal tests were used in this analysis, aided by statistical methods, to determine the factors that affect the metal FeO, and also to make a quantitative evaluation of these factors. Data obtained are from actual iurnace operations, the metal test being of the I spoon, Herty type, and taken in duplicate. Metal temperatures were taken with the open-tube bath pyrometer developed by the Research Laboratory of the United States I Steel Corporation. This method of taking open-hearth metal temperature was invented by Collins and Oseland,l and was developed into its present form by L. 0. Sordahl,2 of the Research Laboratory. The instrument and its mode of operation have been fully described by Sordahl and Sosman in three publications.3,4,5 lag components were obtained by chemical analysis. Data were taken from four 165-net-ton furnaces during the working period of the heat. Recognizing that additions of ore affect the metal FeO. no tests were included within 30 min. after ore additions. In all, 65 tests (48 heats) were used in this study.

Jan 1, 1942

By S M. Aminossadati, R Bagherpour, F Hasheminasab

Mine ventilation is an essential part of underground mining operations used to provide fresh and cool air, as well as to dilute and remove contaminants from the work environment. This paper presents an overview of the ventilation and gas management systems in one of the underground coalmines in Iran: Parvadeh_1 mine. The mine is located in the Tabas coalfield in the South Khorasan province. It is the only fully mechanised mine in Iran utilising the longwall retreat technique. The mine produces 1.1 to 1.3 Mt of coking coal annually. The mine ventilation system currently utilises two 630 kW exhaust surface fans. The gas management system includes surface to in seam (SIS) predrainage boreholes. The underground mine is equipped with an underground gas reticulation pipeline that is connected to a surface plant for the production of methane gas. This study shows that the ventilation and gas management systems of the Parvadeh_1 mine are in line with the standards of Australian underground mines.CITATION:Hasheminasab, F, Aminossadati, S M and Bagherpour, R, 2017. An overview of ventilation and gas management systems in an underground coalmine in Iran, in Proceedings Australian Mine Vent Conference 2017, pp 41–46 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

By Tassos Grammatikopoulos, Gary Pearse, Massoud Aghamirian, Charlotte Forstner, Stephen Mackie

"SGS Minerals Services (Lakefield, Ontario site) has conducted a number of developmental and pilot programs that examined methods for the beneficiation, extraction and recovery of lithium from a variety of spodumene containing hard rock sources. An overview of the beneficiation programs is presented separately by Aghamirian, Mohns, Grammatikopoulos, Imeson & Pearse (2011). The focus of this discussion will be the hydrometallurgical steps involved in producing a high grade lithium carbonate (Li2CO3) sample from the spodumene concentrate. SGS has applied the standard Li2CO3 hydrometallurgical flow sheet to a number of spodumene concentrates and has determined that all of these samples respond in a similar manner to the process and produce a final product having a Li2CO3 grade of >99.5%. The standard process is sufficiently robust to accommodate the observed variability in the concentrates that have been tested. Advantage of one spodumene deposit over another will thus be gained through process optimization and innovation. Optimization at the beneficiation stage and at the hydrometallurgical stage must be considered. Some aspects of the process optimization of the hydrometallurgical flow sheet will be presented.INTRODUCTION The current activity in the exploration, development, and mining of lithium based minerals is due in large part to the key role that this metal plays in the rechargeable battery industry (U.S. Geological Survey, 2011). Lithium battery technologies are widely applied in a range of portable electronic consumer products. The expanding development and production of hybrid and fully electric vehicles has contributed to the increased interest in finding and developing new lithium sources in order to meet the expected demand from the lithium battery industry. Lithium may be sourced from continental brines (sub-surface salars), hard rock minerals (pegmatite, spodumene), hectorite clay deposits, geothermal brines and oilfield brines. Currently the majority of the lithium products are produced from Li brines. The four major lithium producing countries are Chile, Australia, China, and Argentina (U.S. Geological Survey, 2009; Roskill, 2009). These countries accounted for 97.6% of the total lithium produced in 2010."

Jan 1, 2012

By J. G. Bruwer

Paper written on project work carried out in partial fulfilment of B Eng Met Surface coatings persisting on zircon particles, after being leached in an acidic solution, are believed to be the cause of zircon misplacement during the electrostatic separation of heavy mineral beach sands. Electrostatic separation is employed to separate non-conductive zircon particles from more conductive rutile particles. Surface coatings on zircon particles are believed to render the non-conductive zircon particles more conductive, resulting in the contamination of the more conductive rutile stream with zircon particles. Resistivity measurements, X-ray photoelectron spectroscopy (XPS) and scanning electron microscopy (SEM) were performed to compare the surface characteristics of zircon particles found in the different product streams following electrostatic separation. XPS analysis showed that no significant difference exists in the outermost layer of the zircon particles found in the different streams. SEM analysis showed the existing surface coatings on zircon particles from different product streams to have similar chemical compositions. Resistivity measurements confirmed the results obtained from both the XPS and SEM analysis; no significant differences in the resitivities were observed. From the results obtained it suggests that surface coatings present on zircon particles are unlikely the primary cause of zircon misplacement. Zircon losses can possibly be a result of machine related aspects. Key words: XPS; SEM; resistivity measurements; electrostatic separation; heavy mineral sands.

Jan 1, 2007

By Leonard Obert

The ultimate object of this investigation is to develop a method for determining the pressure, in- situ, on rock mine pillars, arches, end other mire structures. The preliminary investigations of Bureau of Mines engineers, as well as those of other investigations4/ have shown that, the velocity of sound depends on the pressure if it is dependent to such an extent that a measurement of the velocity might be used as a means of determining the pressure, a sonic method of measuring the pressure on rock mine columns is possible. Briefly the method would require, first, laboratory determination of the variation of the velocity of sound with pressure in samples of rocks taken from mine pillars to establish a. velocity-pressure relationship for the type of rock and, next, measurements of the velocity of sound in the mine pillars from which the laboratory samples were taken. The pressure on the pillars could be determined by correlating the velocity in the pillar with the velocity-pressure relationship established in the laboratory. Obviously the success of the method depends on the extent to which the velocity of sound varies with pressure in rock.

Jan 1, 1939

By Miller F. S

The famous Golden Mile of Kalgoorlie in Western Australia was, between 1888 and 19ft, one of the richest and most important gold camps in the world. Annual production, which reached a peak of nearly 1,300,000 oz. in 1903, however dropped to below 400,000 oz. after 1920.Depression seized the camp, and many mines closed down. The output from this old field is now steadily mounting under the stimulus. of a higher price for gold. In 1935 300,000 oz. of precious metal were recovered and the future is bright. Total production from the field is 21,400,000 oz., whereas something like £23,600,000 have been paid in dividends.The Golden Mile has been studied by many geologists. Their somewhat diverse findings were published for the most part in the Australian technical press which, unfortunately, is not sufficiently represented in foreign libraries.The following article is based on an eighteen months' study in the district for Gold Mines of Kalgoorlie Ltd., preceded by four years' experience with almost identical rock types in the, Porcupine district of Ontario. The excellent record of observed fact left by previous invesigators, augmented by talks with observant mining men...

Jan 1, 1937

By L. T. Isaacs, W. H. Traves

A numerical model for analysing the drainage of fill in mine stopes has been described. The model can be used to analyse the process of fill drainage in existing and proposed mine stopes. The validity and potential value of the model have been established by experience with a simpler, 2-dimensional numerical model and by testing the new model against the earlier model. The new model is more useful than the 2-dimensional model because it allows for heterogeneity of fill material within the stope - a factor that has been demonstrated to affect drainage characteristics significantly. The 3-dimensional model is also capable of predicting pore pressures and drainage rates at specific locations in the stope without the need to resort to 2-dimensional generalisations. Such information, integrated with relevant field data as they become available, will provide valuable mine planning information for the scheduling of filling operations and design of fill mixes

May 1, 1991

By T King

This paper reviews the changes in the coal industry over the past 30 years and attempts to predict the trends over the next ten years. The last 30 years have seen a steady concentration of production into fewer, larger mines, which operate on 3: more continuous basis, and a major increase in productivity per employee. Coal exports have increased from zero to almost half the country's total production. As well as major changes in exploration and mining equipment and the increasing use of computer tools, the industry now operates in a relatively new regulatory and commercial environment.   Safety legislation now requires the writing and enforcement of policies and procedures by the mining companies, rather than adherence to inflexible regulations enforced by an Inspector. Accident rates have decreased dramatically. The transition from coal mining licences to permits, requiring landowner consent and resource consents, is well under way. Environmental controls are now an essential aspect of the industry.   Equipment suppliers now offer maintenance services, production guarantees and flexible financing.  Contract miners offer mine design and restoration as part of the contract. In a similar approach, Solid Energy now offer an ash disposal service and assistance with resource consents to their customers as well as investment in technology to ensure a consistent product to demanding specifications. Solid Energy has proposals for energy centres which install and operate the coal burning appliances and simply sell energy to the customers. The next 10 years should see an intensification of these trends of bigger mines, more exports and greater technology and a 'package' approach to customer service.

Jan 1, 1997

By Dong Hoon Han

The sulfation rates of magnesium oxide and calcium oxide (molar ratio 7:3) contained in calcined CMA ( calcium magnesium acetate) by SO2 were measured. The experiments were carried out in the temperature range 700 to 1100 °C and under the SO2 concentrations of 0.3 to I mol % (5 % 02). In the low-temperature range below 830 °C the reaction produced CaSO4.3MgSO4 and unreacted MgO, but no MgSO4 phase. The formation of CaSO4.3MgSO4 phase has a significant implication in that it allows the utilization of MgO for SO2 removal whereas the sulfation of MgO from conventional sources is so slow that it is of no practical interest. Thus, the calcined CMA can remove SO2 at lower temperatures than conventional SO2 absorbents and at a greater capacity. Above 950 °C only CaO reacts with SO2, but at a much greater rate than the calcined Ca(OH)2, which is one of the most reactive SO2 absorbents.

Jan 1, 1998