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By E. S. Grumell

The correct sampling of coal and coke is becoming important to an ever increasing number of producers and consumers. This, therefore, may be an opportune moment to examine where we stand with regard to existing and proposed methods. In this paper an attempt will be made to review, as briefly as possible, some of the considerable amount of painstaking research which has been done in the last 10 years, and which has unquestionably led to a much better understanding of the subject. In so doing we shall try, on the one hand, to make clear where substantial progress has been made, and on the other hand, to draw attention to points that require further investigation. The paper is divided into sections in order to emphasize the major divisions of the subject; further, in order to retain this emphasis, detailed consideration of some of the points has been relegated to appendixes. No attempt has been made to give a complete survey of the literature, as this has already been done by L. A. Bushell.2 The following abbreviations will be used: A.E., average error; P.E., probable error; S.W.R., size-weight ratio; B.S.I., British Standards Institution. Section 1.—Cost It has often been stated that the cost of sampling should bear some relationship to the value of the product, and since coal is a relatively cheap raw material it would be a major error to specify, for ordinary commercial purposes, elaborate and costly methods of sampling which nobody would adopt. Specifications must be of such a nature that they will be used. We believe that one of the achievements of the last 10 years has been to reduce the cost of sampling. There is, however, another point to be considered; via., that the cost of sampling should bear some relationship to the value of the investigation. Correct sampling is a cog (possibly a small one) in the wheel

Jan 1, 1940

By E. S. Grumell

The correct sampling of coal and coke is becoming important to an ever increasing number of producers and consumers. This, therefore, may be an opportune moment to examine where we stand with regard to existing and proposed methods. In this paper an attempt will be made to review, as briefly as possible, some of the considerable amount of painstaking research which has been done in the last 10 years, and which has unquestionably led to a much better understanding of the subject. In so doing we shall try, on the one hand, to make clear where substantial progress has been made, and on the other hand, to draw attention to points that require further investigation. The paper is divided into sections in order to emphasize the major divisions of the subject; further, in order to retain this emphasis, detailed consideration of some of the points has been relegated to appendixes. No attempt has been made to give a complete survey of the literature, as this has already been done by L. A. Bushell.2 The following abbreviations will be used: A.E., average error; P.E., probable error; S.W.R., size-weight ratio; B.S.I., British Standards Institution. Section 1.—Cost It has often been stated that the cost of sampling should bear some relationship to the value of the product, and since coal is a relatively cheap raw material it would be a major error to specify, for ordinary commercial purposes, elaborate and costly methods of sampling which nobody would adopt. Specifications must be of such a nature that they will be used. We believe that one of the achievements of the last 10 years has been to reduce the cost of sampling. There is, however, another point to be considered; via., that the cost of sampling should bear some relationship to the value of the investigation. Correct sampling is a cog (possibly a small one) in the wheel

Jan 1, 1940

By J. Petit, P. Girard, A-M. Gilbert, D. Faucher

L&apos;annee 2000 a été une année stratégique et significative pour l&apos; Aluminerie de Bécancour dans la réduction des gaz a effets de serre. L&apos; équipe s&apos; etait donne comme defi un jour d&apos;arriver a l&apos;objectif de zero effet d&apos;anode dans les sa11es de cuve (720 cuves au total). Le cha11enge s&apos;averait tres ambitieux car la fa<;on d&apos;alimenter une cuve d&apos;electrolyse pour obtenir son maximum de rendement faraday est de travailler le plus pres possible de l&apos;effet d&apos;anode pour eviter l&apos;embourbement de ce11es-ci. Ceci signifie que techniquement nous a11ons en moyenne, quotidiennement, 12500 fois pres de l&apos;effet d&apos;anode sur !&apos;ensemble de nos 720 cuves qui produisent aujourd&apos;hui environ 405 000 tonnes d&apos;aluminium liquide annuellement. Des 695 kg equivalent en C02 par tonne d&apos;aluminium produite en moyenne annue11ement, de 1990 a 2000, nous en sommes rendus ace jour a environ 35 kg equivalent en C02 par tonne produite. Nous sommes pres d&apos; atteindre notre objectif mais les demiers gains sont plus difficiles a atteindre. Les strategies d&apos;automatisme et d&apos;informatique, incluant l&apos;aspect de !&apos;implication humaine sur les bonnes pratiques de travail sur les cuves d&apos;electrolyse, sont les cles de ces resultats.

Jan 1, 2006

By I Thomson, L D. Black, R G. Boutilier

The social license to operate (SLO) began as a metaphor for the ability of communities to stop mining projects. Soon people began to speak of levels of social license and the ability of communities to raise costs without completely stopping exploration or mining activities. Today the concept has spread to other industries and is evolving towards becoming a general management perspective on the socio-political rights and responsibilities of the corporation. The business case underpinning the SLO perspective on management is the need to assess and manage socio-political risk. For mining companies, socio-political risk has been steadily rising for several reasons, including the efforts of non-government organisations (NGOs) to pre-empt government policy options; the entry of mining companies into jurisdictions weak in the rule of law; and the weakness of global economic and financial regulation. In this perilous world beyond the comfort zone of effective government, mining companies must build their own socio-political stability by engaging directly with stakeholders to develop a strong SLO. In order to become a practical management perspective, the SLO needs to be measurable, which amounts to measuring the socio-political risks and opportunities presented by stakeholders. We present the results of efforts since 2009 to develop a globally reliable and valid measurement tool at the project and industry levels, validated in studies in Australia, Bolivia, Mexico and the USA. A factor analysis suggested four factors that mine management can address to raise their SLO level. The most chronically neglected factor involves the political right of the corporation to foster more equitable social contracts locally and regionally in both developed and developing countries. We conclude with recommended steps to earn and maintain the social license at the project and industry levels.CITATION:Boutilier, R G, Black, L D and Thomson, I, 2012. From metaphor to management tool - How the social license to operate can stabilise the socio-political environment for business, in Proceedings International Mine Management 2012, pp 227-238 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 20, 2012

By Frank Riddle

THE life of thermocouples is governed, to a large extent, by the protection they receive when in use; particularly when the temperatures being measured are high and the products of combustion are reducing in character. Several types of protection tubes are being used for this work; they are made of ordinary and aluminous porcelain, fused quartz, carborundum, fused alumina, fireclay, and some kinds of metal. All of them may render good service under certain conditions but they must be used for the purpose for which they are intended. For example, fused-quartz tubes are very good when excessive changes in temperature occur, but under certain conditions they will crystallize and lose their strength. Porcelain is used in several ways in connection with the use of thermocouples: for insulating the two wires of the couple and to. prevent short circuits; for protecting the entire couple in a refractory gas-tight tube, which is ordinarily fastened to and is considered part of a properly equipped thermocouple; and for protecting the complete thermocouple from the possibility of physical injury and to support it. This tube is made porous and strong and is ordinarily built right into the wall or crown of the furnace, as the case may be. The insulating tubes must be small and sufficiently refractory to withstand the temperatures at which they are to be used. The length of the tubes, porosity of the material from which they are made, etc. does not particularly matter. The tubes are sometimes made with two or even four holes, the latter being used where it is desirable to have two couples as close together as possible for calibration work. The thermocouple protection tube must be of proper size to permit the thermoelements and insulating tubes to be inserted or taken out easily. It must be porous enough to withstand sudden changes in temperature and yet gas-tight, as well as refractory enough to withstand the furnace temperature, even if the tube protrudes into the furnace in a horizontal position without support for a reasonable distance. If the

Jan 9, 1919

By Laura Prati

Gold-on-carbon catalysts resulted very active and selective in liquid phase oxidation [1, 2, 3] and during past studies we established that activity is ruled by two factors: particle dimension and gold exposure on the carbon surface [4]. These two parameters are also connected to the nature of activated carbon where the chemical nature of the surface oxygenated groups probably plays an important role [5]. Our studies, focused on the design of an active gold-on-carbon catalyst for liquid phase application, were aimed, operating a fine tuning of all the parameters involved in determining activity, to optimize gold-on-carbon preparation, keeping in mind that gold particle activity increases with decreasing particle size and increasing gold exposure. Gold sol immobilisation represents a versatile methodology in preparing gold catalysts [6]: throughout this work we used as sol stabilizing agent, the N-dodecil-N,N-dimethyl-3-amino-1-propan sulphonate (SB), that provides both sterical and electrostatic stabilisation, a condition we already demonstrated necessary to preserve particle dimension when carbon is used as the support [6]. We generated different gold sols by varying the relative amount of reagents (HAuCl4, SB, NaBH4), the type of alkali and the temperature of the gold precursor reduction. Moreover, as we had recently evidenced that there is also a carbon effect [5], we employed only one active carbon to immobilize the sols.

Oct 1, 2003

By H. Copur

Roadheaders have become widely used in underground mining and tunneling operations because they offer the flexibility to excavate soft- to medium-strength rock formations. A critical issue in successful roadheader application is the ability to develop accurate and reliable estimates of machine production capacity and the associated bit costs. Increased pressure to improve productivity and lower costs in the mining and civil construction industries has been the driving force behind the more wide- spread use of mechanical excavation systems. This has caused both industries to move away from the conventional drill and blast methods. Additional benefits of mechanical mining include improved safety, reduced ground support requirements and fewer personnel. These advantages, coupled with recent enhancements in machine performance and reliability, have resulted in mechanical miners taking a larger share of the rock excavation market. Roadheaders are the most widely used underground partial-face excavation machines for soft- to medium- strength rocks, particularly for sedimentary rocks. They are used for development and production in soft rock mining (main haulage drifts, roadways and crosscuts), particularly in coal, industrial minerals and evaporitic rocks. In civil construction, roadheaders are used to excavate railways, roadways, and sewer and diversion tunnels in soft ground conditions. They are also used to enlarge and rehabilitate underground structures. Their ability to excavate almost any profile opening also makes them attractive to those mining and civil construction projects where various opening sizes and profiles are needed. In addition to their high mobility and versatility roadheaders are generally low capital cost systems com- pared with most other mechanical excavators. Because of higher cutting-power density due to a smaller cutting drum, roadheaders can excavate rocks harder and more abrasive than continuous miners and borers.

Jan 1, 1998

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

1.1 INTRODUCTION When underground mining involves total extraction, it induces overburden strata movements. If not properly planned it causes surface subsidence and affects surface environmental conditions. Total extraction usually refers to longwall mining and room and pillar mining with pillar extraction. Surface subsidence has long been a subject of intensive research for scientists all over the world and considerable achievements have been obtained. However due to its difficulties and complicated nature, research into overburden movement has been thus far incomplete as compared to that into surface subsidence. Since surface subsidence is a manifestation of the results of overburden movement, the processes and mechanisms of overburden movement must be fully understood in order to establish the mathematical prediction models of surface subsidence. In this chapter the processes of overburden movement and its zones of movement in the overburden will be discussed. 1.2 PROCESSES OF OVERBURDEN MOVEMENT When total extraction is used, it produces a large void in the coal seam and disturbs the equilibrium conditions of the surrounding rock strata. The roof strata bend downward while the floor heaves. When the excavated area (or gob) expands to a sufficient size, the roof strata will cave. As a result, the overlying strata continue to bend and break until the piles of the fallen rock fragments are sufficiently high to support the overhanging strata. At this tie the overhanging strata no longer cave, but bend and rest on the underlying strata. Strata bending and subsidence develop upward until reaching the surface and forming a subsidence basin. The whole overburden strata and the surface subsidence basin will further go through a period of compaction and gradually become stabilized. Current knowledge regarding the process of overburden movement has been derived from several sources. One is the direct observation of the mined sections and their surfaces, another from field monitoring of strata movements in the overburden, and others from computer modeling and scale modeling in the laboratory. For example, Figs. 1.1 and 1.2 show the results of a borehole monitoring experiment (Borehole B-2) using the full profile borehole inclinometer and full profile borehole extensometer for lateral displacements and vertical subsidence, respectively (Conroy and Gyarmaty, 1983). The panel, 400 ft wide by 5000 ft long and 630 ft deep, was extracted from the Pittsburgh No. 8 seam which had an average thickness of 54 in. The face advanced from east to west. Curve A in Fig. 1.1 shows the inclination of the borehole when it was 45 ft ahead of the face at Position A (i.e., on the solid coal side). The total deviation of the borehole was 1/7000. The first shear movement occurred at 130 ft above the coal seam and appeared to have occurred along the bedding plane with large contrast of rock strata on both sides, i.e., sandstone vs. shale. When the face had passed the borehole 79 ft at Position B, the borehole deformed to assume a bow shape with a maximum deviation of 2.8 in. from the center line. Numerous shearing planes occurred and extended further upward. When the face had passed the borehole by 75 ft at position C, strata subsidence was measured. Fig. 1.2 shows that a strata separation as large as 4 ft occurred at 40 to 100 ft above the coal seam. Above this level, the strata subsided more uniformly with bedding separations within 1 to 2 in. Another example is the longwall mining with complete caving in the Soviet Union&apos;s Karaganda Coalfield (Kolebaeva, 1968). The coal seam was 6 ft thick and 154 ft deep. The immediate roof was the fine-grained sandstone interbedded with coarse-grained sandstone. Above this, there was sandy shale and shale. In order to monitor the process of overburden movement, 15 stations were established in a borehole (Fig. 1.3). The elevations of those stations were measured when the face was at various distances from the borehole. Fig. 1.3 shows the movement history of each station. For convenience of comparison, the movement history of each station was plotted on a common reference point as shown in Fig. 1.4. Clearly, the movements of strata from Station # 1 to #6 were nearly simultaneous, i.e., they behaved as a single unit. The strata from Station #7 to # 15 had differential separations, the maximum of which was 40 in. As the face passed by and moved away, strata separations in general reduced gradually. But strata separation between Stations #12 and #13 remained the same until the end of the monitoring period. The total bed separation between Station #7 and # 15 reached 11 in., approximately 18% of seam thickness, when the face was 132 ft beyond. Based on the above observations, the author developed a conceptual model of overburden movement due to underground longwall mining (Fig. 1.5). The above-mentioned two case studies illustrated that subsidence in the overburden strata propagates upward and subsidence velocity decreases from the bottom to the top. When the subsidence velocity at the surface reaches the maximum value, the subsidence velocity at the bottom portion of the overburden strata has decreased considerably and the strata have begun to compact. Bed separation occurs within a certain distance above the coal seam and reduces from the bottom to the top. When the face has moved away, bed separation reduces gradually, so that eventually some beds completely close and others partially close. Bed separation reaches the highest level when the subsidence velocity at the surface is at its maximum value. For instance in Fig. 1.4 surface subsidence velocity was maximum when the face was 38 to 44 ft past the point. At this time the total bed separation was maximum. Overall the strata directly above the opening are subjected to tension in the vertical direction. But above a certain level, all strata move nearly simultaneously. There is also significant shear movement along some bedding planes. 1.3 ZONES OF MOVEMENT INTHEOVERBURDEN After the extraction of a longwall panel or room and pillar section of sufficient width the strata in the overburden are subjected to various degrees of movement from the bottom to the top. According to the movement characteristics, the damaged overburden can be divided into four zones (Fig. 1.6). Caved Zone. After the extraction of coal, the immediate roof caves irregularly and fills up the void. The strata in this zone not only lose their continuity completely, they also lose their stratified

Jan 1, 1992

By Lucas Gonçalves Generoso, Larissa Rodrigues de Carvalho Rocha, Mauro Vivaldino Fernandes, Gustavo Serrano dos Santos, Gustavo Alexandre da Costa Manso

O Alto-Forno 3 da Usiminas iniciou sua atual campanha em 1999. Projetada para durar 15 anos de operação, previa-se uma produção de 41 Mt de gusa. A partir de 2008, iniciaram as avarias mais significativas no sistema de refrigeração, possibilitando a entrada de água e os consequentes distúrbios operacionais. Buscando estender a atual campanha até 2021, a Usiminas realizou, dentre outras atividades de preservação, duas trocas de staves, em 2015 e 2018. Neste trabalho, são apresentados o planejamento e os principais procedimentos relacionados à parada operacional do alto-forno para a segunda troca de staves, com foco nas atividades operacionais que permitiram a redução do nível da carga até 25,1 metros em relação ao stock line. Destacam-se os cuidados e o êxito no abaixamento de carga até o nível das ventaneiras e a metodologia de troca de staves adotada. Por fim, a previsão de extensão da campanha em 7 anos, possibilitará atingir uma produção acumulada de 53,5 Mt, superando em 12,5 Mt o valor projetado.

Oct 1, 2019

By Nicholas P. Gura

This paper presents the analysis, design, and load test program results for drilled shafts to support a 70-story hotel tower and associated 10 story podium structures located in Las Vegas, Nevada. A mat foundation for the structure, although feasible, would have produced undesirable differential settlements within the high rise tower footprint and at the interface with adjacent lower level podium structures. The resulting foundation system solution for the 70-story, approximately 800-foot long hotel tower was four-foot diameter drilled shafts, extending 100 below shaft cutoff. The design compression capacity for the four foot diameter drilled shaft was 5,000 kips. The load test program proved that the unique Las Vegas valley geology lends itself to efficient installation, precluding the need for engineered slurry or casing. The load testing program included the installation of two four-foot diameter test shafts to 100 feet below theoretical cutoff and the performance of two highly instrumented independent Osterberg load cell tests with maximum sustained test loads of 12,000 kips. Sonic caliper testing was also performed to document actual in-place shaft diameters and verticality. The load test program also provided information necessary for the foundation design of the podium structures. In addition to the tower?s drilled shaft load test program data, continuous flight auger (CFA) pile evaluation and testing was also performed. The CFA testing consisted of traditional ?top down? loading techniques and was performed as an alternative to the recommended shallow drilled shafts for the lower level podium structures. The results of the CFA test program provided supplemental, insightful soil friction data which was used in conjunction with the drilled shaft test program data in evaluating the most appropriate support for the lower level structures.

Jan 1, 2007

By D. H. Stonehouse

"ALTHOUGH final statistics are not yet available concerning Nova Scotia' s mineral production for 1958, it is estimated that the value will be slightly lower than that of 1957. In making this statement it is but fair to say that one year ago a reduction in mineral value was also forecast (for 1957 as compared with 1956) and final figures from the Dominion Bureau of Statistics ($68,058,743) indicated, instead, an increase of $2,000,000 over 1956. However, 1958 did bring problems to the industry which would indicate a drop in production value. The estimates figure for 1958 is $64,000,000. COAL Coal production came from 28 mines, 5 less than were operating since 1957. There were 5,269,896 tons of coal produced by 1,978,512 man-days' labour, for a tons per man-shift of 2.66 (0.05 higher than during 1957) or an increase of 1.9 per cent over the 1957 all-time high for this Province. This record has been established as a result of concentrated efforts to mechanize the coal mines of the Province, and it is hoped that it will be bettered during 1959. Collieries which showed some production for 1957 and none for 1958 were the McGregor at Stellarton, the Franklin at Sydney Mines, the MacLellan, the Rose bank No. 2 and the Rose bank No. 3 at Inverness."

Jan 1, 1959

By R. Loj

Katowice Coal Holding (KHW) is one of the largest coal companies in Europe. The company owns five coal mines with a daily output of around 60 000 tons of coal. In the Katowice city area, very good quality steam coal is extracted. The depth of mining varies between 400 and 1100 m and some of the mines rank amongst those most afflicted by methane in the Upper Silesian Coal Basin, with the methane hazard having a significant impact on production and safety. ?Myslowice-Wesola? mine, being a part of the Katowice Coal Holding is characterized by high methane and rockburst hazards. The mine is equipped with a modern surface drainage pumping station. At the mentioned mine, longwall 514 was selected for analysis of its methane emission. The longwall was extracting the thick seam 501 (over 4 m), at a depth of over 900 m, which was subject to high stress and high gas concentration conditions. At a distance of about 50 m above, seam 501 is covered by thin seam 416. The strata between these two seams are sandstone and mudstone layers. In the floor of seam 501, with a thin shale interlayer, is located seam 510, which is the thickest (over 10m) and richest coal seam in the Upper Silesian Coal Basin. In longwall 514 and it?s vicinity, methane was the dominant risk. For this longwall three methods of classical gas drainage were carried out. At first, drainage was undertaken using long boreholes from a cross-cut at level 665 m, located far from the production panel. Later, it was performed using typical boreholes drilled from the ventilation gallery (tail gate road). Finally, it was undertaken using a degassing drift executed in the upper coal seam 416. All of these headings, ie. cross-cut level 665 m, the ventilation road and conveyor road in seam 501 and the degassing drift in seam 416, were equipped with media for drilling and proper drainage pipelines. Due to the high rockburst hazard and some problems with caving, several kinds of de-stressing and torpedo blasts were carried out in both coal seams and roof rocks. The subsequent analysis and experiments allowed the establishment of some general rules for carrying out degassing operations effectively. The effectiveness of a degassing drift was demonstrated. In some instances, the system allowed more than 70% of the total methane released to be captured and fed into pipelines. The aforementioned blasts had a marginal impact on methane emissions and the efficiency of degassing methods. Cyclic drilling and regular switching of boreholes in phase with the mining progress and working face advance were shown to be essential.

Aug 1, 2013

By A. J. McEvily, R. C. Boettner, C. Laird

The processes leading to fatigue failure in the low-cycle range were studied to obtain an understanding of the basis of Coffin&apos;s law. Particular attention was paid to the manner of mack nucleation and to the determination of the portion of the total lifetime spent in crack propagation. It was observed that cracks are nucleated at surface notches meated during reversed plastic deformation, Principally at grain boundaries, and that crack growth occupies greater than 75 pct of the fatigue life. When inclusions are present, they are often the primary sites of crack initiation. From consideration of the incremental distances advanced by a crack in each cycle and of the lifetimes of different materials, it appears that the universality of Coffin&apos;s law arises from the similarity of crack-tip strain for a given applied strain amplitude, irrespecti7:e of material. STUDIES of low-cycle fatigue have been particularly rewarding in that a simple correlation between fatigue lifetime and plastic-strain range has been shown to exist. This empirical relationship takes the forml,2 Nner=C [1] where N is the number of cycles to failure, Er is the plastic-strain range defined as the total plastic strain per cycle (the sum of the tensile plastic strain plus the compressive plastic strain), n, the exponent, is a constant independent of the material to a first-order approximation and often has a value of about 0.5, and C is a constant which for a number of ductile and metallurgically stable materials is approximately 0.65 for fully reversed strain.3 This relationship has been shown to be valid for a wide variety of metals and alloys and is often referred to as Coffin&apos;s law. However, it is yet not clear why so many different metals and alloys exhibit the same over-all behavior in the low-cycle range. Less is known about the mechanisms of fatigue failure in the low-cycle range than in the high-cycle range, but it is expected that the processes would have much in common. For example, an important characteristic common to both high-and low-cycle ranges is that fatigue crack propagation can occupy a considerable portion of the total lifetime. This has been shown by Laird and smith4 by a count of the number of ripples present on the fracture surface of copper specimens, each ripple corresponding to one cycle of load. On the other hand, some marked differences have been noted. For example, in copper,5 nickel, and aluminum4 there is a greater tendency for grain boundary cracking to occur as the strain amplitude is increased. The purpose of the present work was to study further the processes leading to fatigue failure in the low-cycle range and to obtain an understanding of the basis for the validity of Coffin&apos;s law. Particular attention has been paid to the manner of crack initiation and to the determination of the portion of total lifetime spent in crack propagation. MATERIALS A number of metals and alloys were tested, although more attention was focused on a particular few, once general trends had become apparent. Alloys were selected so as to investigate the effects of crystal structure, number of phases, stacking-fault energy, and impurity content. The materials in this grouping were: high-purity copper OFHC copper 1100 aluminum (commercially pure) 70-30 brass (Cu-30 wt pct Zn)

Jan 1, 1965

By N. S. Dalal, B. Jafari, F. H. Y. Green, V. Vallyathan

"INTRODUCTIONThe mechanistic details of the biological events leading to coal workers pneumoconiosis (CWP) are not yet fully understood, despite several decades of extensive epidemiologic1.2 and laboratory studies.2-6 Epidemiologic studies1.2 have shown, for example, that the incidence and severity of CWP differ markedly in the different regions and mines at comparable exposures, but laboratory investigations2-6 have demonstrated only partial correlation of the epidemiologic data with the differences in the mineral composition and the rank of coal mined. In particular, while epidemiologic data indicate1.2 direct correlation of the· prevalence and severity of CWP with the rank (i.e., % carbon content) of coal, this correlation has not been established by laboratory4'6 and animal exposure3 studies. In order to explain these results, in 1980 Artemov and Reznik' suggested that perhaps they arise from the fundamental differences in the surface properties of the coal. dusts inhaled by the coal miners and those used in the laboratory studies: while the miners inhale freshly fractured coal particles, henceforth called the 'fresh' coal dust, the laboratory studies generally utilize 'aged' coal dusts (i.e., dusts that have been stored for many days or even longer). Reznik and Artemov' used electron spin resonance (ESR) spectroscopy to show that the mechanical crushing of some Soviet Union coals generated organic free radicals and that the concentration and the decay times (hence the reactivity) of these radicals were higher for the coals of higher ranks. Since some of the radicals decayed within a few minutes in air, the authors surmised that while these radical species could lead to certain specific pathogenic reactions at the sites of mining operations, this might not be the case in the laboratory studies, due to the conventional use of 'stored' dusts which would be expected to contain significantly smaller free radical concentrations. This is consistent with the recent findings that, standardized, aged coal dusts exhibit minimal cytotoxicity. 8While the above mentioned work of Artemov and Reznik7 did suggest a possible new clue to the pathogenesis of CWP (i.e., the role of the coal based free radical species), no direct biological/cytotoxicity data were provided to show that 'fresh' coal dusts were, indeed, more pathogenic than the 'stale' ones. Because of the rather significant implications of the Artcmov-Reznik hypothesis to the understanding of 1hc biochemical mechanism and, hence, the strategies for the eventual containment of CWP, we have initiated a comparative study of the free radical formation and cytotoxicity properties of freshly crushed coal ·particles. As done by Artemov and Reznik, 7 we have used the ESR technique as the direct method of measuring the concentration and decay kinetics of the coal based free radicals. Our preliminary ESR studies on two Pennsylvania coals, a bituminous (carbon content 72 % ) and an anthracite (95 % carbon) coal, have confirmed the Artemov-Reznik finding that the crushing-induced free radical sites are higher on the coals with higher carbon content (i.e., higher rank).9-11 In the present work we describe our more recent results of a parallel study of the time dependence of the decrease in the free radical content of a freshly made anthracite dust (as measured via ESR) and that of the dust's cytotoxicity potential as measured by the extent of hemolysis of (sheep) erythrocytes. We have also investigated the effects of free radical scavengers on the cytotoxicity potential of the dust and deduce that indeed the free radicals could play a significant role in the initial events in the mechanism of the cytotoxic effects due to the inhalation of coal dust inhalation."

Jan 1, 1990

By J. T. McCartney

Thin sections of coal have been prepared for many years either by cutting on a microtome (1,2)5 or by grinding and polishing (2,1). Specimens of most coals had to be demineralized in hydrofluoric acid and softened in caustic solutions before they could be successfully cut with a conventional microtome. Some lignites could be cut without pretreatment. Jeffrey (1) obtained sections from coals of all ranks, although long and tedious pretreatment was required for anthracite. His sections were less than 1 cm.2 in area (probably much less for high rank coals) and about 5 microns thick. Thiessen (2) also cut sections of softened coals, mostly of low rank. His sections were a few square millimeters in area and about 3 microns thick. However, Thiessen preferred ground and polished sections for detailed studies because they were larger and had not been chemically attacked by softening media. He prepared ground sections as large as 4 cm.2 in area and usually 5 to 10 microns thick. Thousands of such sections, covering complete column samples of many coal beds, were studied for a number of years (4). With the development of the electron microscope and increased interest in quantitative optical studies of coals, a need for much thinner sections arose. However, no suitable technique for preparing such sections was known until Fernández-Morán developed a diamond knife for cutting ultrathin sections of hard materials (5). Workers at the Centre d&apos;Etudes et Recherches de Charbonnages de France applied this new tool to the cutting of coal sections and have recently described their technique to the Burg, of Mines. Using their methods, with some modifications, researchers of the Bureau of Mines have succeeded in cutting ultrathin sections of coals ranging in rank from lignite to

Jan 1, 1961

By Byron Pickard

AT MIDNIGHT of Sunday, Aug. 27, 1922, a fire was discovered in the main shaft of the Argonaut mine, Jackson, Cal. With the excep-tion of a shift boss and two skip tenders who escaped, the entire night shift of 47 men was trapped in the mine and entombed for 22 days, when the bodies were recovered. The entombed miners made a valiant but fruitless effort to save their lives. Rescue crews toiled strenu-ously day and night for 21 days to save the entombed men, only to find they had arrived too late. From the experiences of the entombed miners and of the rescue crews, valuable lessons may be learned to prevent future catastrophes.

Jan 11, 1922

By T D. Jolly, L A. Balzan, A R. Harris, Z Bauk

"The availability of real-time data in any process control system is critical for effective management of that system. At MMG LXML Sepon Copper Operation in Lao PDR, controlling the copper grade of the feed to the leaching circuit is paramount for maximising both productivity and copper recovery. Since 2008, real-time analysis data has been available to the process engineers from the Geoscan on-belt analyser installed on the Copper Mill feed conveyor, and highly accurate results on copper grade have allowed the process engineers to strictly control blending to overcome the variability in plant feed grade. The availability of real-time analysis data is crucial, as effective sampling of this stream is both difficult and error-prone, being only a snapshot in time of highly variable ore stockpiles, as well as having significant laboratory turnaround delays. Periodic sampling is used to verify the Geoscan data as accurate. Since its installation, the copper result has been the most critical output of the Geoscan. In recent times, the process engineers have sought to utilise other results from the Geoscan to further improve and control the process. One measurement of particular interest involves the real-time measurement of calcium and magnesium as a gauge of ore feed gangue acid consumption. To satisfy this end, accurate measurement of calcium and magnesium is becoming increasingly more important and has been the subject of ongoing collaboration between Sepon and Scantech, the manufacturer of the Geoscan. This paper discusses this work, describing the calibration process and results obtained, while also providing a basis for Geoscan use in operations to enable enhanced process control for all potential Geoscan users.Citation:Balzan, L A, Jolly, T D, Harris, A R and Bauk, Z, 2016. Greater Use of Geoscan On-belt Analysis for Process Control at Sepon Copper Operation, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 231&ndash;238 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By W. H. Emmons

THIS paper is the third of a series treating of the relations of ores of the metals to igneous rocks. The first1 outlined the general problem; the second2 proposed a classification of lode ores, and showed how the distribution of the deposits is in a certain measure controlled by the contour of the roof of the parent batholith. A number of mining districts were discussed, the examples being selected chiefly from group 2 of the classification. The proposed classification of lode deposits, briefly stated, is as follows: 1. Cryptobatholithic: near hidden batholiths which have not yet been exhumed by erosion. 2. Acrobatholithic: in and near cupolas or domes of batholiths-the high points of larger underlying masses. 3. Epibatholithic: on batholiths, near their rims, but below eroded cupolas, or below the highest points of the roofs. 4. Embatholithic: among closely-spaced outcrops which probably are parts of a single batholith. The outcrops of invaded rocks predominate. 5. Endobatholithic: in and near roof pendants of large batholiths. The invading rocks predominate. 6. Hypobatholithic: in deeply-eroded batholiths, where even roof pendants, have been -eroded. Deposits of certain metals on the roofs of the batholiths are much more numerous in some positions than deposits of other metals, and the horizontal zonal arrangements of lodes of different metals are shown characteristically only in certain groups. In a large measure, also, the character of the fracturing and metallization is determined by, the-position on the roof, of the batholith.

Jan 1, 1927

By Shiro Ban-ya, John Chipman

Using the same experimental method previously described, the activity of sulfur in a number of quaternary and more complex liquid iron alloys at 1550oC is determined. A "cross-product term zj.zk is determined for each of the quaternary alloys to take account of the mutual effects of the combinalion on the indi-vidual contributions to In 5 The interactiotl coefjicienls delermined for ternary solutions and cross-pvoducl coefficierlts from the quaternary systems are used to estimate the activity coefficient of sulfur in more complex alloys. OUR recent papers 1,2 have described new determinations of the activity of sulfur in Fe-S binary and Fe-S-j ternary solutions, but the practical applications of these data require the estimation of activity of sulfur in more complex systems. Previously, Sherman and chipman3 devised an approximate method for calculating the activity of sulfur in complex systems by summing the contributions to the logarithm of the activity coefficient of each element in the alloy. This procedure is more successful, as one of us has pointed out,4 when the concentration variable is taken as zs = ns/(nFe — nS t CLJ~ nj) rather than the conventional atom fraction x~ = nS/(nF, + nS + Enj). In the former definition, which is based on analogy to interstitial solutions, the value assigned to uj is normally +1 for metallic and -1 for nonmetallic elements." In addition, however, the effects of mutual interaction between alloying elements must be considered for the accurate estimation of the activity of sulfur in more complex systems. The purpose of this investigation has been to devise a more reliable method for estimating the activity coefficient of sulfur in nmIticomponent iron alloys. The measurements were carried out on some quaternary systems Fe-S-j-k for combinations of silicon, chromium, nickel, cobalt, molybdenum, and tungsten. Manganese was not included since the experimental method has been found unsatisfactory for volatile, sulfide-forming elements. The experimental method was the same in every detail as described previously. 1,2 The concentrations of sulfur, silicon, chromium, and copper in the quenched metals were determined by chemical analysis but those of nickel, cobalt, molybdenum, and tungsten were calculated from charged amounts. Iron sulfide was included in the charges in amounts estimated to furnish approximately the final sulfur content. CALCULATION W2 have shown2 that when zs and zj are the concentration variables in a ternary Fe-S-j solution the logarithm of the activity coefficient S = a~/zg is a linear function of the variables as and aj defined as follows:

Jan 1, 1970

By E. T. Turkdogan, J. H. Swisher

bi the fivst part of the paper results ave given on the rate of decarburization of Fe-C melts ln CO2-CO atmospheres at 1580°C. The rate -controlling step is believed to he that irvlloluing dissociation of curbotz dioxide on the suvfuce of the melt. 4 genevral reaction mechanistm is poslnlated jor gels-t11eta1 veactions oc-curit~g on the surface of iron coutcotamncited with chemi-sovbed osygesL. Oxygen the present work on decavbuvization of liquid iron and previous studies on the kinetics of nitrogen absorption and desorplion are discussed in terms of the postulated mechanism, ManY of the early studies of rate of decarburization of liquid steel were of an exploratory nature and laboratory exppriments carried out pertained to open-hearth or oxygen steelmaking processes. References to previous work on this subject may be found in a literature survey made by Ward. Using more sophisticated experimental techniques, several investigators have recently studied the kinetics of decarburization of molten Fe-C alloys in oxygen-bearing gases. For example, Baker et al2.&apos; reported their findings on the rate of decarburization of liquid iron, levitated by an electromagnetic field, in carbon dioxide-carbon monoxide-helium atmospheres. In these levitation experiments the samples used were small in size, e.g., -0.6-cm-diam spheres weighing -0.7 g, and the rates were measured for decarburization from about 5 to 1 pct C at 1660°C. The rates obtained under their experimental conditions were considered to be controlled primarily by gaseous diffusion through the boundary layer at the surface of the levitated melt. Parlee and coworkers3 measured the rate of absorption of carbon monoxide in liquid iron. The rates were found to follow first-order reaction kinetics, yielding a reaction velocity or a mass transfer coefficient in the range 0.2 to 0.4 cm per min. The coefficient was found to decrease with increasing carbon content of the melt. These investigators attributed the observed rates to the transfer of carbon or oxygen through the diffusion boundary layer adjacent to the surface of the melt. In the work to be reported in this paper, an attempt has been made to study the kinetics of gas-metal surface reactions involved in the decarburization of liquid iron. EXPERIMENTAL The experiments consisted of melting 80-g samples from an Fe-1 pct C master alloy in an induction furnace and decarburizing in controlled CO2-CO mixtures at 1 atm pressure and 1580°C. The master alloy was prepared by adding graphite to electrolytic "Plastiron" melted in racuo. None of the impurities in the master alloy exceeded 0.005 pct. The reacting gases were dried by passage through columns of anhydrone; in addition, CO2 impurity in carbon monoxide was removed by passage through a column of ascarite. A schematic diagram of the apparatus is shown in Fig. 1. A 1.25-in.-diam recrys-tallized alumina crucible containing the sample was placed inside a 3-in.-diam quartz reaction tube, all of which was surrounded by an induction coil. A 450-kcps induction generator was used as the power source. Water-cooled brass flanges, which contained the gas inlet, gas exit, and sight port, were sealed to the top of the reaction tube with epoxy resin. The reacting gases were metered with capillary flowmeters and passed through a platinum wire-wound alumina preheating tube, 0.25 in. ID and 11 in. long. The gases were preheated to about 1300°C. A disappearing-filament optical pyrometer was used to measure the melt temperature. The pyrometer was initially calibrated against a Pt-6 pct Rh/Pt-30 pct Rh thermocouple. The temperature was controlled to within +10°C by manually adjusting the power input to the induction coil. In a typical experiment, an 80-g sample of the master alloy was melted in a CO2-CO atmosphere having pcO2/pco = 0.02 and flowing at 1 liter per min. A negligible amount of carbon was lost and no significant reduction of alumina from the crucible occurred during melting, e.g., 0.005 pct Al in the metal. After reaching the experimental temperature of 1580°C, the gas composition was changed to that desired for a particular series of decarburization experiments. The duration of the transient period for obtaining the desired gas composition at the surface of the melt was about 20 sec . The flow rate of the reacting gas was maintained at 1 liter per min. After a predetermined reaction time, the power to the furnace was turned off. During freezing, which took about 10 sec, the amount of gas evolution was not sufficient to result in a significant loss of carbon. The samples were analyzed for carbon by combustion and in a few cases they were analyzed for oxygen by the vacuum-fusion method. RESULTS A marked increase in the rate of decarburization of iron with increasing pcO2/pco ratio in the gas stream is demonstrated by the experimental results given in Figs. 2 and 3 for pco2/pco ratios from 0.033 to 4.0. In one series of experiments, denoted by filled triangles in Fig. 2, the reacting gas was diluted with argon (48 vol pct) resulting in a slower rate of decarburization. Samples from two series of experiments with pco2/pco = 0.033 and pco2/pco = 0.10 (with argon dilufion) were analyzed for oxygen. In these Samples the oxygen content increased with reaction time

Jan 1, 1968