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By Donald A. Dahlstrom

If you have ever driven on the Autobahns of Western Germany, you are well acquainted with the "fast lane." There is no speed limit; there are only two lanes in the direction you are going, and if you choose to drive in the fast lane, you had better be prepared to drive faster than anyone or someone will tailgate you with a constant screeching horn until you move to the "slow lane." The prize for this mastery of the fast lane is some kind of a "macho" feeling and getting to your destination sooner than your competitor. This has some comparisons with today Is liquid-sol id separation. We've certainly had our moments in the fast lane but we've also had other countries forcing us into the slow lane at times. It can be very embarassing when someone almost runs over your back and then shakes a fist at you as you shakily scatter to the much slower and overcrowded lane. Because of necessity, there will be many new improvements and added requirements in the next decades into the 21st century and beyond. Those countries and industries that can maintain the pace in the fast lane will be successful while the others may well be relegated to the non-competitive lane. Before considering what will be the challenges in the future for competitive liquid-sol id separation performance, it would be wise to review progress over the past forty years in this area. The author has to admit his ancient age and believes he can act as an evaluator of these many advances. To place the beginning of the 40 years (the author predates his initiation into liquid-sol id separation by one more year), the United States was still building smelters and operating them with little or no sulphur dioxide removal , practically

Jan 1, 1989

By K. Lillkung, O. Forsen, J. Aromaa

"The dissolution rates of Pt, Pd and Rh in 250 g/L NaCl solution were calculated from the current densities determined by polarization curves and potentiostatic measurements measured with wire electrodes at atmospheric pressure. The results achieved with the two methods were similar. The dissolution rates at the corrosion potential were in the order of few nanometers per hour. The polarization curves indicated that to achieve dissolution rates required for economical operation on an industrial scale, the samples need to be polarized to higher potentials. When polarized to higher potentials, the metal dissolution rates were in the order of micrometers per hour. As the Tafel slopes of the metal dissolution region are fairly steep, the potential affects the dissolution rate greatly. Based on the activation energies, the dissolution reactions were controlled by a charge transfer step.IntroductionPlatinum group metals (PGM) have some very unique properties such as their inertness, giving excellent corrosion resistance, and a high melting point. At the same time, these factors make the refining of these metals very challenging. The main applications where PGMs are used are different catalytic converters, mostly in automotive catalysts. The used catalysts are also an attractive secondary source of PGMs. The exact concentrations of PGMs in catalysts vary widely depending on the manufacturer, the age and the capacity of the catalyst. The concentration of Pt in a new catalytic converter range from 300 to 1000 ìg/g, Pd from 200 to 800 ìg/g and Rh from 50 to 100 ìg/g [1]. Precious metal catalysts usually consist of a coat of the metal supported on ã- alumina. The recovery of metals from automotive catalysts has received much attention during the last few years [1-3]. In the near future, the use of PGMs in fuel cells is predicted to increase the amount of secondary PGMs available [4]."

Jan 1, 2012

A study of the effect of nickel on the cupellation process has shown that a very low "nickel limit" must be imposed for cupellation (0·08 grams nickel or 0·05 grams when copper is present). In the presence of slightly greater amounts of nickel, lead tends to spreadon the cupel with detachment of drops from the main lead pool. In crucible fusions nickel contaminates the lead button to a small extent, and this contamination is increased.in the presence of either iron or copper. Nickel also causes "shotting" of the lead, this effect being more apparent in the presence of iron.Both the spreading during cupellation and the shotting during crucible fusion are considered to be caused by adsorption of insoluble NiO at the Pb-PbO interface, lowering the interfacial tension.Excess litharge crucible charges are' fused at 116boC. to eliminate shotting. A brief· soaking in molten basic litharge slag effectively reduces the nickel content of lead buttons. The solubility of silver in litharge slags is considerably decreased in the presence of nickel. Corrections can be applied for silver losses in slags but these corrections are not as accurate as with copper ores.INTRODUCTIONThis paper on the effects of nickel on fire assays for silver is the second of an intended series describing investigations into the fire assaying method.In the first paper, discussing the effect of copper on fire assaying processes (Guerin, 1957), it was shown that certain elements, because of their relatively low affinities for oxygen, will tend to contaminate the lead button during crucible fusion and scorification. Nickel, having a slightly greater affinity for oxygen than has lead, would be expected to be oxidized and slagged readily, and would not be expected to appear in the metal phase (lead button) to any appreciable extent. Thus, judged solely on this basis of...

Jan 1, 1960

By J Gutzmer

The Kalahari Manganese Deposit in South Africa accounts for 80 per cent of the worldÆs land-based mineable Mn reserves. Although the ores are primarily mined for export, South Africa is the worldÆs third largest producer of ferromanganese and ferrosilicomanganese alloy, both of which are derived from the smelting of these ores. Mineralogical analysis has traditionally been forsaken for bulk chemical assaying in quality control of smelter feeds, principally because quantitative data could not be easily produced. Yet, mineralogy plays a vital role in determining smelting behaviour, which in turn has economic implications when aspects such as energy and reductant consumption are considered. Smelting is an endothermic process, and the amount of energy input required is dependent on the minerals present in the ore, as well as their relative proportions. Further, knowing the bulk Mn oxidation state in the ores allows for calculation of the amount of oxygen that must be driven off to produce alloy, and hence the amount of reductant required. Additionally, certain minerals can be correlated with potentially deleterious elements, so that ore mineralogy can indicate penalty element distribution. Mineral abundance analysis was conducted on ores from the Kalahari Manganese Deposit using quantitative X-ray powder diffraction. Rietveld refinement using a fundamental parameters approach was applied to micronised ore samples. The mineral percentages obtained were validated using mineral chemical data together with bulk major element data derived from ICP-OES. Bulk oxidation state partitioning was then carried out knowing the Mn valence in the minerals. The results show potential for implementation of the proposed method on a smelting plant, where ore blending could be then implemented to achieve consistent feed compositions to smelters. Selective mining could also be undertaken to obtain desirable ore compositions for domestic or international markets.

Jan 1, 2008

By J. A. Holmes

As the Appalachian mountains reach their maximum development in western North Carolina, we find also in that region indications of extensive dynamic disturbances and alterations undergone by the rocks in' connection with these mountain uplifts. Among the minor results of these changes have been the formation of numerous dikes or a veins " of exceedingly coarse granitic materials, which in some places are mined for the mica which they contain, and in other places are quarried for kaolin. These dikes are filled with quartz, feldspar, and mica, in varying proportions, very coarsely crystallized. Sometimes the mica (generally muscovite), sometimes the feldspar (generally albite or orthoclase) predominates. When the mica is present in considerable quantity, and in large crystals, the dike is usually spoken of as a mica-vein, and is often worked for mica. Sometimes these crystals of mica occur on one side or the other; sometimes on both sides; and sometimes they are largely concentrated toward the middle of the vein, or, it may be, distributed throughout the dike with a considerable degree of uniformity. In many cases the crystals of mica are too small and few to permit the working of the vein as a micamine ; in other cases, very little mica is present, and the feldspar constitutes the larger part of the material. This feldspar of the dikes undergoes, through the weathering action of the atmosphere, certain chemical changes, resulting in its alteration from feldspar into kaolinite—the kaolin of commerce. These dikes vary considerably in size, ranging from a few inches to several hundred feet in thickness, and up to many hundred yards in length. They are generally parallel to the schistosity of the crystalline rocks, which, however, in some cases, they cross at varying angles. The kaolin in those dikes which occur in the Unaka or

Jan 1, 1896

By F. M. Russell

Scheduling the extraction and rates of production for an underground mining operation is a common task carried out during the life of a mine, from the initial feasibility study through to the final production phase. The scheduling process can be complex when several requirements have to be satisfied, for example, match mill capacity and maximize net present value while constrained by mining sequence and availability of resources. Many mines have computer programs relevant to their operation but because they are site-specific they need considerable reprogramming to be adapted to other operations. Other methods currently in use also suffer drawbacks: manual spreadsheets are laborious, computerized spreadsheets cannot handle sufficient constraints, and operations research techniques require a complex mathematical model to be set up. The use of network analysis techniques has been proposed previously but the unavailability, until recently, of easily used computer programs has not allowed this approach to be developed. There are now many programs which can be used; however, the particular program which has been adapted to mine scheduling is Metier's 'Artemis-PC' running on IBM-XT or compatible machine. The use of network analysis offers many advantages, including the use of a logical mining sequence; activities can be as detailed as required; related activities can be readily incorporated, and costs can be easily extracted. A heuristic approach is necessary to optimize the schedule, but the ease and speed of program execution enables this to be readily achieved, while leaving the mining engineer in full control. The paper discusses the problems encountered in mine scheduling and current solutions. It then details the use of network analysis techniques and the application of the Artemis-PC program to solve mine scheduling problems.

Jan 1, 1987

By H. M. Chance

Probably no other prosperous mining district is so little known as the Black Hills. The name leads one to assume that the district is a rolling country, consisting of more or less insignificant hills. Nothing could be more untrue, as the Black Hills is an elevated region rising to heights of 7000 and 8000 feet above sea-level, and broken by deep gorges and ravines. It is, in fact, a mountainous country in all that that term implies ; but the cañons and gorges are of such a character that the country is easily penetrated by railroad-lines, and good wdgon-roads exist to-day, affording access to every portion of it. Aside from the stampede to the Black Hills country in 1876, when gold-placers were discovered, and 15,000 or 20,000 people became infected with the mining fever, the region has never had a mining " boom." Quartz-mining commenced in 1877 or 1878, and has been successfully prosecuted ever since; the output of gold steadily growing, until at present the yield is about $7,000,000 annually. This great production is about equally divided between two entirely different classes of deposits. The first class, comprising the so-called quartz or free-milling ores (of which the Homestake mine is the principal producer), are veins of generally steep dip, in the older metamorphosed rocks, and similar to the veins worked in other mining districts. The second class of mines is located in what is known as the " siliceous gold-belt." The Siliceous Gold-Belt. This belt, more particularly to be described in this paper, lies in the immediate vicinity of Deadwood and Lead City, and covers an area 6 or 7 miles in length by 3 or 4 miles in width. The formation in which the siliceous ores are found is a nearly horizontal series of sandstones and shales, generally

Jan 1, 1901

A PAPER on "Iron and Steel Scrap in Blast-fur-nace Burdens," by W. C. McKee, general super-intendent, Federal Furnace plant, By-Products Coke Corp., Chicago, was published in the October issue of MINING AND METALLURGY, and written discussion by Richard Moldenke was published in the December issue. Following is the stenographic report of the oral discus-sion that took place when this paper was presented at Pittsburgh. J. N. REESE, Monessen, Pa.-I came to the conclusion that Mr. McKee must be using a poor grade of scrap, therefore his loss was high. Mr. Franchot said that there was ex-cess heat in the stack. There is no excess heat in the hearth, therefore, any heat that is used to melt this scrap must come from the stack. By using scrap in a blast furnace you can utilize this excess heat in the stack. The amount of scrap that can be used in any particular furnace involves a direct relation to the efficiency of that furnace. If a furnace is running on a very low-coke rate, you will not be able to use much; if on a high-coke rate, you can use considerable scrap, probably several times as much as that mentioned in these papers-without using any additional coke. This excess heat can be utilized with-out practically any serious mechanical and operating diffi-culties such as might be occasioned by removing gas from a point other than the top of the furnace. I do not think, scrap has any effect on the iron, although I think some foundrymen are trying to prove that it has, and perhaps some open-hearth men may have that idea.

Jan 1, 1927

By R. M. Hudson

Equilibrium constants have been determined for the jeaction C (in steel) + H2O = CO + H2 as a function of carbon content (0.013 to 0.74 ujt pct) and temperature (1200° to 1800°F) by using a flow system in which the water-vapor content of various gas mixtures containing hydrogen, carbon monoxide, and nitrogen was systematically controlled. The equilibrium partial-pressure ratio K1 = PcoPH2/PH2O was obtained for two low-carbon steels and four Fe-C binaries from a series of runs made at each temperature. By comparison of Ki with the analogous ratio for the reaction C (graphite) + H2O = CO + H2, the activity of carbon in steel (ac) was comDuted for each temperature and carbon content studied. From these data values of the equilibrium constants for two other carburization-decarburiza-tion reactions of importance were estimated. The chemical compositions of the two steels and four Fe-C binaries used are given in Table I. Specimens of Steel E were cut from a 0.035-in.-thick commercially box-annealed and temper-rolled lot of a low-carbon low-metalloid steel. Steel F was a commercially hot-rolled rimmed steel that was cold-reduced at the Laboratory to a 0.059-in. thickness. Fe-C binaries A, B, C, and D were vacuum-melted and poured, hot-rolled, pickled, and cold-reduced at the Laboratory to furnish specimens 0.035 in. thick. Sheet specimens were cut from all lots to approximately 3/4 by 3 in. by gage and were vapor-degreased with trichlo-roethylene and weighed before heat treatment. Heat treatments were carried out in a hinged-top Hevi-Duty electric furnace with a Vycor-glass tube. When the furnace was at the desired

Jan 1, 1965

By R. L. Fleischer

IN 1925 a criterion for localized deformation in a single crystal was derived by Taylor and Elam.' They noted that for a single active slip plane the slip plane area is constant, but the direction of the stress with respect to the slip direction changes during deformation so that, when the increase in stress due to this rotation is large enough relative to the work hardening, instability occurs. (By instability is meant continued slip on the active slip plane.) The instability criterion may be written t>dt/d?/m tan2? [1] where t and ? are resolved shear stress and strain, m the Schmid factor for resolving shear stress, and ? is the angle between the slip direction and the axis of tension. The formula points out that regardless of the atomic mechanism involved, for large flow stresses and low work hardening rates, slip on a single plane—once started—should continue. The case of alloying provides a direct possibility of adjusting the parameters in Eq. [I]: As an alloying element is added to a pure metal, the flow stress is raised and the work hardening in easy glide may be either decreased2 or unaltered.3 Since flow stresses in easy glide may rise to the kg/mm2 range for alloy crystals while easy glide slopes may be less than % kg/mm2, 2 it is possible for the inequality to be satisfied. Indeed Garstone and Honeycombe3 report that slip becomes coarser with increasing stress in the easy glide range, where the flow stress would be increasing and the slope would be essentially constant. Another case where high flow stresses and low hardening rates accompany coarse slip is that of quench hardening.4 Maddin and Cottrell report that quenched aluminum crystals display slip markings resembling those of a brass. A similar effect has been observed after irradiation.'

Jan 1, 1960

By P Townsend, F Russell

Palabora Mining Company operates an integrated open pit mine, concentrator, smelter and refinery complex located in the Northern Province of South Africa. The Palabora open pit went into production in 1966, mining 30 000 t/d to produce some 62 000 tpa of copper. With continuous modification and improvement, a mining rate of 82 000 t/d, producing up to 135 000 tpa of refined copper, has been achieved. While these improvements have also enabled the mine life to be nearly doubled, the open pit will reach the end of its economic life by 2003. The Palabora orebody is vertical and the reserve extends to a depth of 1800 m over an area of 700 m by 200 m. Studies to assess the economics of mining the orebody by underground means began in the mid-1980s. These studies indicated that block caving is the lowest cost underground mining method that would generate a suitable return. In March 1996, a decision was made to develop a 30 000 t/d block cave mine that will extend the life of the Palabora Mining Company by a further 20 years. In terms of the relatively high rock strength and the resultant coarse fragmentation that is expected at the drawpoint, the Palabora cave can be considered rather unique when compared to other block cave operations. This paper describes the project in broad terms and details the mining method, particularly with respect to the cavability of the orebody and the design features to deal with the coarse fragmentation. The project commenced in July 1996 and final development of the undercut and drawbells is scheduled for completion in November 2003. This paper describes the status of the project and planning as of April 2000.

Jan 1, 2000

By Katragadda Sarveswara Rao

"In view of the increasing importance of processing lean ores as potential future resources, efforts are being made worldwide to develop suitable technologies for the recovery of both nonferrous, base and precious metals. There are various types of low grade ores present in many countries. Gradually, the mineral reserves for production of metals have undergone complex changes with regards to their grades and particularly the physical and chemical properties of the minerals to be separated from each other. So, there is a need for new processing routes or the upgrading of existing technologies to recover metal values that are otherwise unutilized. Accordingly, an overview of the different types of mineral resources and economical processes, based on ore horizons, consumption of energy and reagents and environmental concerns, is made in this paper. In addition, the paper provides a summary of metallurgical and engineering challenges to be overcome when implementing new technologies and concepts with regard to base metal extraction and sulphur removal efficiency.IntroductionLow grade multimetal complex sulfide ores are likely to become an importance source for nonferrous as well as precious metals in the near future. The principle metallic minerals are chalcopyrite, sphalerite, galena, pyrite, pyrrhotite, arsenopyrite, tetrahedrite, boulangerite and cassiterite. Such ores are located throughout the world as multi-pocketed reserves which have present complex challenges with regard to their grades and particularly the physical and chemical properties of the minerals to be separated from each other. Their compositions can vary widely from one deposit to another and even from one location to another in the same deposit, and this poses a problem for the recovering metals present therein. Successful aqueous processing of complex sulphide ores depends on the composition and mineralogy of the raw materials and on their leaching behaviour."

Jan 1, 2012

By Oliver C. Ralston, James Norman

DIATOMACEOUS earth occurs in deposits widely distributed throughout the nation. The chief producing areas are in the Western States, where many high-grade deposits are known. Eastern deposits of diatomite are unsatisfactory for many purposes either because of impurity or because of the extremely fine particle size. Eastern users of diatomite, on the whole, must pay for transportation of the material from the west coast. The freight rate is necessarily high because of the great bulk of the diatomite. A satisfactory process of purifying the diatomaceous earths of Maryland and Virginia, which are unsuited for wide use only because of impurities, should be of value to eastern users. In the past, benefication of diatomite by wet processes has been limited to classification or sedimentation for the removal of sand from diatoms. Calcination with salt is used to produce a white diatomite when iron is present in the impurities, and small amounts of clay are rendered inactive in the process. No treatment has been developed whereby a large proportion of clay can be removed from a low-grade diatomaceous earth to give a diatomite of very high purity. In August 1937 the Federal Bureau of Mines, in cooperation with Dr. R. C. Williams, of Baltimore, undertook the problem of purifying low-grade Maryland diatomite by Froth flotation. Work was done at the Eastern Experiment Station, College Park. Unless otherwise stated, the crude diatomite used was from the Patuxent River deposits near Dunkirk. This material, consisting of roughly one-third each or clay, quartz sand and diatoms, was representative of the usual low-grade marine diatomite found in the Coastal Plain Miocene beds of Maryland and Virginia and sporadically in other Atlantic coast states. As received, it contained 50 to 60 per cent moisture and probably had been wet since its formation.

Jan 1, 1942

By R S. Suglo, G-K Er

Mineral processing is an essential component of a mine production chain and it accounts for significant proportion of the production cost. Over the past decade, significant research progress has been made in processing plant efficiency using theoretical, experimental and simulation modelling techniques. However, these studies are limited to process plant system as stand-alone units with no interactions with the mining units. In order to maximise the total mining system efficiency and more accurately simulate the system, the processing plant must be modelled as part of an integrated mining system. This becomes increasingly important in continuous surface mining systems, with limited ore processing capacity. In addition, plant efficiency is roughly estimated with known parameters or with simplistic statistical data, without regards to the stochastic processes governing these parameters, such as processing plant feed introduction. In this paper, a comprehensive model of the ore processing plant system is given and the object-oriented design (OOD) of the system is briefly introduced. The joint operation of ore processing plant system and haulage system is analysed. Numerical modelling is used to analyse the ore processing plant system under different conditions. Analysis of the results show that the optimum plant capacity is 0.7 tons/sec (2520 tons/hour). The results also show that the plant system resources cannot be fully utilised because the production rate is always less than the plant capacity. When plant capacity reaches an optimum value 0.7 tons/sec, the production rate is about 0.64 tons/sec. The underlying stochastic processes associated with plant and mine capacities cause this variability. In order to reduce this variability, there is the need to tightly set tolerances for the function variables.

Jan 1, 2003

By H. O. Hofman

The sudden decease, April 1, 1910, in Chicago, of Dr. Franklin R. Carpenter was a shock to his many friends. He died in his sixty-second year, of heart paralysis. To most fellow-members of the Institute who have followed the progress of the last twenty-five years in the mining and treatment of non-ferrous ores, the work of Dr. Carpenter in that line is familiar; but very few, even of those who knew hinl well in the later years, are acquainted with his earlier struggles to satisfy an inborn hunger for knowledge. The story of his life resembles that of the pioneer who opens up new country; overcoming hardships with indomitable courage, meeting doubts and difficulties with an alert and open mind, and exercising in the face of reverses an unconquerable perseverance. It seems appropriate to place on record in the Transactions a brief outline of the life and work of this eminent engineer. During the past 15 years of his life Dr. Carpenter was much interested in genealogy. In the pursuit of this study as an avocation, he traced his own ancestry to the time when ((Hugh," called "le Charpentier" on account of the dexterity with which he had swung his battle-axe as a crusader, settled in Sussex on lands received from William the Conqueror. The Carpenters came to this country with William Penn. A prominent member of the family was Joshua Carpenter, the founder of Christ Church, Philadelphia. Dr. Carpenter's branch of the family moved from Pennsylvania into that part of Virginia which is now West Virginia, where his grandfather settled in Harrison county, while the grand-uncle went further west, and built the first iron-furnace at Hanging Rock, Lawrence county, Ohio. In fact, since early times, the Carpenters appear to have been iron-masters. Franklin R. Carpenter mas born, Nov. 5, 1848, in Parkers-

Jan 1, 1911

The methods employed in dressing 'the tin ore up to the required percentage of purity depend mainly on the type of ore crushed.These may be briefly divided into two distinct classes-viz., q, free: milling ore and a complex milling ore. In the former the ore is practically free from the common metallic sulphides, such as pyrites, mispickel, blende, galena, etc., and 1!' clean concentrate of 70 per cent metallic tin can be directly obtained in one continuous act of concentration; in the latter, however, in order to get a good extraction, it is only advisable to primarily concentrate up to 30 to 40 per cent. tin oxide, the rest being the sulphides of the metals. This concentrate is then roasted, re-ground, and re-concentrated up to 70 per cent. metallic tin on a separate set of tables.In the complex milling ore the tin is intermixed very closely with the sulphides, so that it becomes necessary to collect all the mineral present in order to obtain clean tailings. Stamp mills are the most popular machines for the primary crushing in the district, but whether they are invariably the best for the purpose is rather an open question. No doubt for some of the very tough chloritic ores they are the most efficient, but for a brittle quartzose ore, such as the free milling proposition usually is, the writer thinks Oornish rolls would well hold their own, especially when the tin occurs coarse, as is so often the case with this class of ore. Of course, the great drawback to stamps is the excessive quantity of tin slimed and slime tin is one of the problems of tin-dressing-while rolls give...

Jan 1, 1907

By R Berthelsen

The Peak Gold Mines Pty Ltd (PGM) operations are located approximately 9 km south of Cobar, NSW. PGM currently mines and treats ore from three separate orebodies within 5 km of the original plant. Production from the original Peak orebody ceased in April 2002, with ore currently sourced from underground at New Occidental and Perseverance, while New Cobar supplies ore from an open cut mine. In March 2003, Wheaton Minerals Asia Pacific purchased the operation from Rio Tinto who operated the mine for around ten years. The Perseverance deposit is situated between 600 and 1000 m south of the Peak orebody. Access is via shaft from Peak to 630 m below surface and then by decline to the northern end of the deposit locally known as Zone A. Development of the resource commenced mid 2002 to set up for a long hole open stoping operation at 25 m level spacing. The first production stope was successfully brought on line December 2003. The Perseverance deposit is hosted by sediments and volcanics along the eastern margin of the Devonian Cobar basin and is one of many deposits with the Cobar Gold Field. The Cobar Gold Field has produced approximately 2.5 million ounces of gold since 1880. The Perseverance area had long been considered prospective due to the presence of old workings at the southern extensions of Peak line of lode. Significant mineralisation was first discovered at Perseverance in 1994 when PGM drilled a hole targeted 950 m below the historic workings. The target was based on an apparent plunge to the stratigraphic controls to the Peak Orebody. Subsequent exploration within and around the ægreaterÆ Perseverance area has been significantly aided through the use of downhole electro-magnetic surveys due to the close spatial association between gold and sulfide mineralisation.

Jan 1, 2004

By Robert A. Jankowski, Fred N. Kissell

Recent allegations about tampering of coal mine dust samples have led to suggestions that the coal industry needs a respirable dust compliance sampling method that is less subject to tampering. The sampler designs under consideration include fixed location monitors that could be mounted on a mining machine or located at a stationary location in the mine, and samplers that are worn, like the existing samplers. Fixed-point samplers have a design advantage, since samplers worn on the body will have size and weight limitations. On the other hand, in some cases there are respirable dust concentration gradients, which will dictate careful placement of a fixed-point sampler inlet to determine the exposure of workers. This study examined dust concentration-gradient data obtained by the Bureau of Mines and others over the last decade. As an example of spatial gradients at continuous miners, a fixed-point monitor located in front of the operators cab shows a higher dust concentration than a sampler located on the operator and reduced exposure when the sampling location is behind the operator cab. It may be necessary to use a sampling tube with a machine mounted monitor, to sample in the operators breathing zone in the cab to provide a more representative sample of the miners exposure. For face workers on longwalls or on continuous miner sections using remote control, the study shows significant dust level gradients and gradient variations between typical operator positions and mining machine locations. In addition, if fixed-point monitoring is used to monitor an area, vs the exposure of specific workers, dust control technologies currently employed to ensure reduced dust exposure such as the diversion of dust laden air away from the workers, or through the use of remote control of mining machinery may not indicate the same dust reduction obtained with current sampling techniques.

Jan 1, 1993

By Fred A. Krafft

The time allotted to this paper will permit only the sketching of general principles and practice as generally employed in the industry to maintain and preserve the physical well-being of the miner and his family. Let it be emphasized on behalf of the industry that coal opera- . tors, individually and collectively, have given primary consideration to the appointment of a mine or community physician who is charged with the responsibility of caring for the sick and injured. However, it too frequently happens that with the selection or appointment of the camp doctor many operators wash their hands of further responsibility toward the health and the working efficiency of their employees. The coal industry is so organized in most of the coal-producing areas that not only is the health of the individual miner an important production problem but the collective health of the entire mining community is a fundamental operating problem. This is particularly demonstrated when epidemics of communicable diseases attack entire communities, with resultant loss of life and production and general community distress. The problem and responsibility of maintaining high health standards among miners and their families should be given as much consideration as safety, economical production and efficient distribution are given in the successful operation of coal mines. Company Policy In the inauguration or administration of a medical program for a large mining operation or for an individual mine, the first question to be solved is what the company policy is to be with respect to the following factors pertinent to a medical program: 1. On what basis will the company and the employees finance the program ? 2. Should a contract doctor or a full-time doctor on salary be employed ? 3. What measure of service is to be rendered the employee and his family in return for his monthly contribution to the medical program?

Jan 1, 1936

By A. F. Lucas

Introduction BELLE ISLE, located in the low sea marshes near Atchafalaya Bay, is the southeasternmost of the famous Five Salt Islands of Louisiana. Rising about 80 ft. (24 m.) above the level of the surrounding marsh and being some 800 acres (323 ha.) in extent, it constitutes a striking landmark on the low Gulf coast, and since the eighteenth century, when it played an important part in the operation of the buccaneers, it has been intermittently frequented by man. In 1897, the writer discovered the great mass of salt that underlies the island and shortly afterward pub-lished a brief account of the geology of Belle Isle and the neighboring salt islands.' In that paper he described principally the salt, but pointed out also the occurrence of petroleum and of sulphur in the strata above the salt. Since that time the island has been extensively explored by the drill, and the developments up to 1908 have been described by Veatch2 and Harris," and Knapp.4 The writer's early conclusions as to the probable presence of sulphur in considerable quantity have been confirmed by the later drilling, some of the wells having found quantities of gas containing sulphur dioxide and a number of sulphur-bearing strata as well. Under the stimulus of the present keen demand for sulphur for war purposes, a syndicate of New York capitalists was recently formed to explore the sulphur resources of the island in more detail. Under the writer's direction, six wells were drilled for this syndicate during the winter of 1916-17. The present paper presents briefly the results of this work and sums up the present knowledge of the geology of this highly interesting locality.

Jan 9, 1917