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By M. B. Standing, D. L. Katz

Equilibrium data are reported on the composition and densities of coexisting vapor and liquid phases as a function of Pressure and temperature for four hydrocarbon systems prepared from crude oil and natural gas. The data were observed over a range of pressures from 1000 to 8220 Ib. per sq. in. and at temperatures ranging from 35' to 250°F. The compositions of the hydrocarbon systems were such that the critical temperatures of the mixtures were lower than the range of the investigation. Under these conditions, it is shown that the composition of the systems has a marked effect upon both the absolute value of the equilibrium constants and the change of the constants with pressure for the several components comprising the system. From weight-balance considerations, it has been possible to construct the partial phase diagram of a mixture of a 49.9" A.P.I. crude oil and natural gas having a gas-oil ratio of 3660 cu. ft. per barrel. Descriptions of the apparatus used, the analytical procedure, the material used, and the equilibrium measurements are included. A knowledge of the physical behavior of naturally occurring hydrocarbon mixtures is fundamental to the treatment of many problems in the production, transportation and rehing of petroleum. The recent discoveries of pools having formation pressures up to 7500 lb. per sq. in. and higher, and the development of full-scale equipment to operate at pressures of 5000 lb. per sq. in. or more have greatly increased the need for knowledge of the composition, temperature, pressure and density relationships of coexisting vapor and liquid hydrocarbon phases. The following paper presents the results of investigations upon 22 mixtures of natural gas and crude oil of known com~o~itio~ that had been brought to equilibrium under controlled conditions of temperature and pressure. Data are presented upon the densities and compositions of the coexisting vapor and liquid phases for four different systems containing hydrocarbon material having boiling points from minus 2 5S°F. to well over I000 F. The data were observed Over a pressure range from Iooo Ib. per Sq. in to 8220 lb. Per sq. in. and a temperature range from 35' to 2 50°F. The accepted method of computing the physical relationship between two coexisting phases of a hydrocarbon mixture is by use of the "equilibrium constant" K as proposed by Souders, Selheimer, and Brown. definition,

Jan 1, 1944

General discussion on the above subject, presented at the Pittsburgh meeting, October, 1914. GEORGE S. Rice, Pittsburgh,. Pa.-Undoubtedly there is a serious problem through the juxtaposition of gas and oil wells and coal mines, not only at the present time, but possibly of far more serious import for the future. If one examines a map of the coal fields on which is superposed the oil and gas fields of the country, one finds that they overlap to a very great extent. I may say at the beginning that we have not, so far as I know, had any very large number, of accidents growing out of this situation causing loss of life. Such accidents, or rather incidents, that have occurred have all been rather potential possibilities for disasters due to the leakage of gas from wells into the adjacent coal mines. The condition has been most acute in the Fairmont district of West Virginia, and in the general coal fields southwest of Pittsburgh in Pennsylvania. The accidents which have actually occurred were summarized in Bulletin No. 65 (Oil and Gas Wells through Workable Coal Beds) issued by the U. S. Bureau of Mines, reporting the transactions of several meetings in Pittsburgh, February, 1913, of oil and gas well representives, State geologists, State mine inspectors, coal-mine operators, and representatives of the U. S. Bureau of Mines. This conference was partly brought about through a case in which the Pennsylvania State inspection department had brought suit to restrain a gas company from completing a well through a certain mine. The office of the Attorney General of Pennsylvania asked the opinion of the Bureau of Mines upon the matter; about the same time the Association of State Geologists, at its New Haven meeting, asked the Bureau to take up the problem. Accordingly, it was thought advisable to bring together the several, interests. For the most part the danger is, I think, greater in the future, than at the present time Many parts of the coal fields of Pennsylvania, in Greene County particularly, have been penetrated by numerous wells very close together, and with the present practice of having a large coal pillar around each well, has made or will make the mining of many areas difficult or pro-

Jan 4, 1915

By S. M. Parmley

Experience has shown that there are some factors connected with the drying of fine washed coal that are not present in drying of slack coal as normally practiced at cement kilns or pulverized coal plants. It is the purpose of this paper to show some of these variations as they have been encountered under actual operation. The Champion No. 1 cleaning plant of the Pittsburgh Coal Co. uses the wet system of cleaning coal. The plant is equipped to produce commercial sizes or mixtures, the primary sizes being plus 6 in. and 4 by 6 in., which are hand-picked. Minus 4-in. sizes are all cleaned by the wet process. To produce a commercial product as free of moisture as possible, the washed coal is first passed over sizing and dewatering screens, then each size except the — ?-in. coal is conveyed to loading booms by slow-moving draining conveyors. As the +?-in. sizes are not heat-dried, no data will be given, as it is the purpose of this paper to describe the heat drying of — ?-in. coal as practiced at the Champion No. 1 plant of the Pittsburgh Coal Company. To fully appreciate the conditions, it must be remembered that the — ?-in. product is now shipped at a lower moisture content than when received from the mine. The incoming coal in the — ?-in. size will average 5 per cent. moisture and the shipped product 3 per cent. moisture. This is a case where a wet process gives a reduction of both ash and moisture in the fine sizes. The shipped product, therefore, is more uniform in all characteristics, and uniformity of product is the most important single factor in coal preparation. The 0 by ?-in. coal is first dewatered on slow-moving continuous bucket dewatering elevators, running at the rate of 40 ft. per minute, with 48-in. perforated buckets with ¼ by 1½-in. slots, spaced ? by 2?-in. centers staggered. The moisture in the coal discharged from the elevators runs 20 to 28 per cent., depending on the fineness of the product. The discharge from the dewatering elevator is further drained on a drag-type conveyor equipped with ½-mm. wedge wire screen. The conveyor discharges coal with an 18 to 25 per cent. moisture to three centrifugal dryers of the Carpenter type. The Carpenter dryers

Jan 1, 1931

By Augustus J. Bowie

THE discovery of the auriferous cement gravel deposits in and near Deadwood Gulch, Lawrence County, Dakota Territory, in 1876, created a mining excitement, and rush to the Black Hills. Numerous ten and twenty-stamp mills were built on Deadwood Gulch ; Central City and the town of Deadwood sprang into existence, and became the centres of the mining industry of the Hills. Gold quartz deposits were subsequently discovered and located on a so-called mineral belt, extending from Whitewood Creek on the south, to Deadwood Gulch on the north. A northerly group of these quartz locations was purchased in December, 1877, by parties in San Francisco,, who formed a company, incorporated under the name of the Father de Smet Consolidated Gold Mining Company. This company immediately commenced the work of exploration. The deposits were proved to be extensive, and easily mined, but it was found that the ore, on an average, would not. yield more than $10.00 per ton* of 2000 pounds. At this time the cost of supplies and materials in the Hills was high. Miners' wages were $3.50 per diem, and custom mills charged $6:00. per ton. The successful working of mines on "the belt" necessitated cheap milling. The ores being of so low grade,† this could be accomplished only with large and economically running mills. To insure the greatest possible economy, the Father,& Smet mill * Up to December 31st, 1879; the de Smet mine had produced bullion of the value of $718,211.68 from 76,155 tons of ore, being an average value of $9.41 per ton. During the year ending December 31st, 1879, the mine produced 46,344 tons, which yielded $425,360.42, or $9.18 per ton. The total gross cost of working for the corresponding period, including deadwork, exchange, legal expenses, fete., wits $219,914.82. The yield of the mine for the year 1880 (Engineering and Mining Journal, February 5th, 1881), was $600,011, with total expenses of mining and milling, of $313,108. † The first 4083 tons extracted from the de Smet mine, were selected, yielding $62,783.97. The next 25,728 tons produced $230,067.24. The total 29,811 tons were weighed, and worked at custom mills. The yield of the first 409 tons from tine Golden Terra Mine was $11.44 per-ton.

Jan 1, 1882

By G. Swiatnicki, w. Chruscielewski, T. Domanski

INTRODUCTION The inherent errors of simple passive track detectors /SPTD/ used for estimation of exposure to the radon daughters /RnDP/ amount to about 30% /Do75/ while for active detectors only reach about 3 - 4% /Ph78, Du79/. The large error inherent in SPTD is the result of strong dependence of its calibration coefficient on the equilibrium Rn RnDP in the vicinity of detector /0r80/. Nevertheless the SPTD shows many advantages and seems to be very useful for to the estimation of individual exposure of miners to RnDP /Do81a/. For this reason the investigations are undertaken to improve SPTD so that it might provide information about the equilibrium Rn : RnDP during exposure. If the equilibrium would be known then one could use a suitable calibration coefficient yielding a better precision of the measurement. So far, some investigations of the improved type of SPTD have been conducted. One was based on the concept of using the so called diffusion barrier permitting to differentiate the tracks originating from Rn and RnDP /Fr77/. Such improved detector has been named a passive differentiating track detector /PDTD/. However, own investigations have shown /Do81b/ that this type of detector cannot provide precise informations about equilibrium Rn : RnDP. For this reason the investigations of the new idea of PDTD have been carried out. The new model presented below is based on the use of the absorbers of alpha particles. THEORETICAL CONSIDERATIONS The first variant of the concept is as follows: The differentiating detector consists of two portions of detective foil /track detector/ but only one is covered with an aluminium absorber. Thickness of the absorber is carefully selected so that only RaC' particles /7.7 MeV/ can pass the barrier. The second portion of the detective foil records all alpha particles originated from Rn and RnDP i.e. works like a SPTD. The scheme of such a detector, the so called two-fields passive differentiating track detector /PDTD-2/ is shown in the Fig. 1. It is possible to introduce a parameter that can characterize the equilibrium Rn RnDP, namely, the so called track density ratio /TDR/. The definition of this parameter for the considered type of detector is TDR2 = D1•D2-1. One can expect the simple proportion to hold: [ ]

Jan 1, 1981

By L. A. Willey

IN 1953, a method for filtering a liquid phase from solid phases while in equilibrium at elevated temperatures was devised at the Alcoa Aluminum Research Laboratories. Since then, it has been applied to the investigation of liquid-solid equilibrium in metallic systems. In application, the method simplifies the centrifuge filt.ration procedure as developed by Scheil and described by Gebhardt Fig. 1 shows a sketch of the filtering apparatus located within a thermal analysis furnace. It consists of a stainless steel filter crucible A, into the bottom of which a porous carbon filter disk B, 1 3/4 in. diam by 1/4 in. thick, is sealed to the shoulder and rim with carbonaceous cement. The base of the filter crucible, machined to an angle, fits into a tapered seat at the mouth of a stainless steel vacuum chamber C. A connecting tube from the vacuum chamber to a vacuum line also serves as a means for supporting the apparatus from the base of the furnace. An alundum crucible D of 25 ml capacity is placed with- in the vacuum chamber below the filter for collecting the filtrate. Two thermocouples E, in stainless steel protection tubes, which extend from 1/16 to in. above the center of the filter, are mounted on the rim of the filter crucible. One thermocouple is used in differential connection with the furnace control thermocouple to regulate rates of heating and cooling. The other serves as a means for recording the temperature of the alloy prior to filtering. A suitable mold coating is used on exposed surfaces of stainless steel parts to guard against attack upon these parts as well as to avoid contamination of the alloy samples. A mold wash prepared from precipitated CaCO, is suitable for aluminum alloys. The porous carbon for the filter disks and the carbonaceous cement, grade 200 (graphite base), were obtained from the National Carbon Co. Porous car-bon, grade Nos. 10 to 60, has been employed. Grade No. 40, however, appeared to give the optimum advantages in operation. In operation, an alloy sample is remelted and heated to about 50°C above its liquidus temperature. At the same time, the filter apparatus is preheated to about the same temperature. The alloy then is cast into the filter crucible.

Jan 1, 1957

By Walter O. Snelling

In the course of some experiments more than five years ago, made for a totally different purpose than the investigation of the oil used, I placed a small quantity of a transparent yellow lubricating oil in a bomb-like vessel and heated it to a relatively high temperature. At the end of the experiment I removed the oil from the vessel and was amazed to find that instead of bearing any resemblance to the oil which I had put in, it now had the appearance of ordinary crude oil. The green color by reflected light and the rich red-brown by transmitted light were so unmistakable as to at once lead to further investigation. I subjected the material to fractional distillation, and the surprise which I experienced at the appearance of the oil, changed to amazement when I found that it yielded, on distillation, 15 per cent. of gasoline and 30 per cent. of burning oil, and that its constitution resembled crude oil quite as much as did its appearance. Furthermore, the gasoline and kerosene distillates which it yielded were of a clear water-white color, entirely without treatment with acid or alkali, and were entirely free from the odor familiar in "cracked" petroleum distillates. The result of this experiment was quite too remarkable to be credited without further confirmation, and I at once filled the vessel with some of the same oil that I had used before, and again heated to about the same temperature that I had previously used, and for the same period of time. Upon opening the vessel and removing the contents I found, not the material resembling crude oil that I had obtained before, but apparently only the same oil that I had put in, somewhat darkened in color, but nevertheless far different in appearance from the material obtained in the previous experiment. Evidently some condition existed in the first experiment that had not existed in the second test, and here began a series of tests in which I sought by the change of one variable after another to arrive at the identical conditions which must have existed in the first experiment. Only the fact that the bottle of heavy oil used in the first test was still in its place,

Jan 1, 1915

By H. R. Collins

The purpose of pulverizing coal before burning it is to make available every heat unit it contains. Machinery has been developed which will pulverize coal in one operation, delivering it to bins in front of the furnaces at an expenditure of about 17 hp.-hr. per ton, in a medium-sized plant. The cost of the operation depends upon the amount of moisture that must be expelled before pulverizing, the cost of labor, and the cost of coal delivered at the plant. At a small plant, requiring a pulverizer with a capacity of only 1/2 ton per hour, the cost per ton pulverized will naturally be greater than at a plant requiring the largest pulverizer, with a capacity of 7 tons per hour. :The first step is to reduce large lumps to a size suitable for drying uniformly, before passing to the pulverizing mills; this is done in rolls, at a single pass. The second step is the elimination of moisture, in order to facilitate pulverizing to great fineness, while also increasing the heating effect and the temperature attainable when the coal is burned. There are other mechanical advantages in the handling of dried coal. Driers are now manufactured which are able to eliminate moisture without distilling any of the volatile combustible matter in the coal; they are fired by hand or with pulverized fuel. The heat first surrounds the shell of the drier, being confined within a chamber where complete combustion takes place; the heated gases are then led through a duct to the discharge end of the drier and enter the inside of the shell at a temperature not exceeding 300" F. This temperature is maintained by the operator and is indicated by a pyrometer. Volatile combustible matter is not likely to be distilled until the temperature rises above 400" F. On discharging from the drier, the coal is usually passed over a magnetic separator in order to prevent pieces of iron from going to the pulverizer. Two types are used: a magnetic pulley which automatically discharges its collection of iron, and a lifting type, from which the iron is removed by hand when convenient. In the operation of pulverizing, the coal should preferably be reduced until 95 per cent. will pass through a 100-mesh and 70 per cent. through a 300-mesh sieve. Such a product is obviously an almost impalpable powder.

Jan 1, 1920

By C. O. Brawner

Introduction The procedures used to design slopes in rock have only been developed over the past 10 years. Therefore, continued review and assessment of field experience by the publication of case studies is necessary to determine the accuracy of the theory and the application of the design techniques being used. Where failures have occurred, mine management should insist that the slides be 'back analyzed' to determine the shear strength parameters at failure. These values should be used for any pit design as they reflect site conditions more accurately than any other procedure. The previous papers have dealt with the basic concepts and factors that influence stability. To illustrate the application of these concepts in the study of practical stability problems, six ease studies have been selected for review. The examples include failure of tailings dams due to an earthquake, failure and redesign of a slope in overburden, review of failures of rock slopes, calculation of the amount of seepage into a pit and influence of the groundwater on stability, and the design of slopes for a proposed 2,500 ft. deep open pit mine. All but the first example are drawn from the authors personal experience. EXAMPLE 1 Failure of Tailings Dams Due to an Earthquake On March 28. 1965 a serious earthquake occurred in Chile with the epicenter located near the town of La Liqua, about 80 miles north of Santiago. The earthquake had a rating of 7 to 7.25 on the Richter scale. A record of the acceleration was taken at Santiago where a maximum horizontal acceleration of 0.18 g. was recorded. Many small mines are located in this region. Each of the mines provided impoundment storage for tailings with one to three dykes. As a result of the earthquake, most of the dams failed. The most serious failure involved two dams at El Cobre where more than 2,700,000 cu. yd. of tailings flowed into the valley, travelled over 7 miles in a few minutes, destroyed part of the town of El Cobre and killed more than 200 people. The Chilean government commissioned an investigation into the failures. A summary of the findings was published by Dobry and Alvare7 in 1967 (1). The location of the earthquake and the tailings dams which were investigated is shown

Jan 1, 1971

By D. G. Westlake, D. T. Peterson

The saturation solubility of thorium dihydride in thorium was studied by saturation of samples and subsequent analysis. The solubility increased from about 1 at. pct at 300°C to above 20 at. pct at 800 °C. Over this temperature range the log of the solubility was a linear function of the reciprocal absolute temperature for thorium of good purity. The heat of solution of thorium dihydride was found to be 7.9 kcal per atom of hydrogen in crystal bar thorium and 6.6 kcal per atom of hydrogen in Ames thorium, The important effect of hydrogen on the physical and mechanical properties of many metals is receiving increasing recognition. The basic physical and chemical data necessary to help explain the observed effects of hydrogen on the mechanical proper ties of metals are all too often not available. The solubility of hydrogen in thorium metal was determined to provide a portion of this basic data for the thorium-hydrogen system. The absorption of hydrogen by thorium metal and the formation of thorium hydrides was studied by several investigators, beginning in 1891, and their work is summarized in Gmelin's Handbuch.1 No effort was made to determine the extent of solid solubility in the metal in these early investigations and the impure metal available made interpretation of the results uncertain. Nottorf2 reported the results of a careful study of pressure-composition isotherms for the thorium-hydrogen system. The principal interest in this work was in the composition and dissociation pressures of the thorium hydrides. Mallett and campbel13 reported the pressure-composition isotherms for the thorium-hydrogen system up to the composition of thorium dihydride. The saturation solubility of thorium dihydride in thorium was reported to be 13 at. pct at 650°C and 23 at. pct at about 900° C. Mallett and Campbell used thorium metal which had been prepared by the calcium reduction of thorium tetra-fluoride or thorium oxide. The results obtained with thorium of these two types agreed within the experimental error. PROCEDURE AND MATERIALS Procedure—The method used to determine the solubility was to saturate a sample of thorium with hydrogen and analyze a portion of the saturated sample. This method was chosen rather than a determination of pressure-composition isotherms since it was much less time consuming and could be applied at temperatures at which the equilibrium hydrogen pressure was too low to measure. The

Jan 1, 1960

'T. C. MERRIMAN, New Haven, Conn. (written discussion *).-This most interesting paper gives much carefully obtained and valuable data. However, there are two points in connection with the commercial application of such data that might possibly be a source of trouble and mis-understanding. In the first place, the examination of thin sections of annealed brass, subjected to standard Brinell test.(500 kg. on a 10-mm. diameter ball) at thicknesses from 0.075 to 0.150 in. (1.9 to 3.8 mm.) indicates that cold work has been performed on the specimen during the application of the load sufficient to extend way through the section, and have the hardness of the metal affected by the backing. If the usual steel support is used, this effect is very evident. If a piece of soft brass is used as a support, there is an indentation in the surface of the support accompanied by a bulge on the under side of the tested specimen opposite the impression made by the 10-mm. ball. The result is that Brinell specifications (governmental or otherwise) on stock for a certain purpose may not be fair in all cases, since manufacturers, owing to differences in equipment, etc., may not all use precisely the thickness and conditions of stock on which the Brinell specifications were based. This would mean that the government Brinell specification might not give them the temper of stock best suited for their manufacturing methods. The concluding paragraph of this paper, stating that "the hardness of cartridge brass may be determined with greater accuracy by the Brinell hardness measurement than by attempting to judge it from the grain size" appears to be in line with the movement of the last year or so among brass manufacturers toward substitution of the Brinell test for microscopic examination as an acceptance test on cartridge brass. The statement may he strictly true that the Brinell test is the best test for hardness, but it is not, as the statement might readily be construed, a sufficient test of suitable condition. For instance, some cartridge brass might accidentally have been overheated to a point where it would be distinctly unsafe to use for the manufacture of small arms cases and then be so rolled (reduction 3 to 5 per cent.) as to pass perfectly proper Brinell specifications. Under such conditions the microscope would reveal its unsuitability where the Brinell test had failed so to do.

Jan 4, 1919

What is a nugget effect? In the early development of geostatistics, the term ?nugget effect? was coined for the apparent discontinuity at the beginning of many semi-variogram graphs. This name was chosen to reflect the large differences found between neighbouring samples in ?nuggety? mineralizations such as Wits gold reefs. Geostatistical theory assumes that the difference between a sampled value and a potential repeat sample at the same location is actually zero. Included in this ?nugget effect? would be true variation between contiguous samples due to the nature of the mineralization, micro-fracturing, nugget or crystal size and so on. Also included in the nugget effect would be any ?random? sampling variation which might occur due to the method in which the sample was taken, the adequacy of the sample size, the assaying process, etc. Arguments were put forward that ?sampling errors? actually exist at zero distance. Some geostatistical schools actually maintain that ?nugget effect? is all sampling error. This would imply that ?perfect? sampling would eliminate the nugget effect entirely. There is now a dichotomy both in the geostatistical world and in the software packages provided for geostatistical analyses. It may seem academic to argue over whether the semi-variogram model should take a value of zero, a value equal to the nugget effect or a partial value at distance zero. However, the decision can have a profound effect on both the estimated resource and in our confidence on that resource. Whereas most geostatistical texts define the semi-variogram model as taking the value of zero at zero distance, others imply that the full nugget effect should be used at zero distance. For example: ?The nugget effect refers to the nonzero intercept of the variogram and is an overall estimate of error caused by measurement inaccuracy and environmental variability occurring at fine enough scales to be unresolved by the sampling interval3. ?Christensen4 has shown that the ?nugget effect?, or non-zero variance at the origin of the sernivariogram, can be reproduced by a measurement error model. ?The nugget effect is considered random noise and may represent short-scale variability, measurement error, sample rate, etc.5. In many training texts and web courses, the definition of the semi-variogram is ambiguous as the formulae for semi-variogram models is not actually specified at zero distance6,7,8.

Jan 1, 2009

By Jan Visman, James W. Parkinson

GENERAL INTRODUCTION Sample collection, preparation and analysis are all steps in a single process designed to provide information about a much larger amount of material than the sample itself. For example, information about the ash content of a unit train load of coal may be derived from a single, relatively small sample of perhaps a few hundred pounds (liters) extracted from the entire shipment. These three steps can be thought of as links in a chain, allowing the simple analogy that the strength of the chain, or, in other words, the value of the information is only as strong as the weakest link. Sample collection, discussed in Part 1 of this chapter, is critical because analytical procedures can do no more than reveal the characteristics of the sample presented to the laboratory. If that sample does not properly represent the population to be characterized, neither will the analysis of that sample. Sample preparation is also a critical part of the chain. Any change in the nature of a sample between the time it is collected and the time it is analyzed will appear to reflect the nature of the population. For example, sample preparation procedures that allow air to partially dry the coal sample before reaching the laboratory will cause the sample analysis to indicate a lower moisture content than the sample originally had. Other changes can be more subtle. Second only to moisture loss, oxidation reactions generally cause the greatest problems. Oxidation of a coal sample can alter surface properties, changing flotation response. In extreme cases, combustion parameters, such as caking or agglomerating properties or mineral composition, such as pyrite content, can change. Water samples can also be altered by oxidation. For example, ferrous iron can oxidize to ferric iron, possibly changing water quality parameters such as pH. Sample analysis, the subject of Part 2 of this chapter, is the final link in the chain. Obviously, no laboratory procedure can correct for problems created by improper sample collection and preparation. At the same time, poor analytical techniques will create incorrect information about the total amount of coal from which the sample was extracted (the population) in the same manner as improper sample collection and preparation. The key to gaining correct information about the population is to view these three steps as parts of a single process. Following proper techniques through only one or two steps of the process will not produce accurate information.

Jan 1, 1991

By Robert Livermore

It has been a noteworthy feature of the Cobalt camp, that many of the valuable ore deposits have been covered, wholly or in part, by small but usually deep lakes, such as Cobalt, Cart, and Peterson lakes, and with the subject of this article, Kerr lake. In the early days of the camp many water-covered areas, which have since been found to be valuable, were ignored or neglected by owners, or simply perfunctorily staked for water rights. Kerr lake is probably the evidence of a fault or sharp fold, having a general east and west strike, and containing a series of ore-bearing fissures, parallel generally to the strike, although having some notable exceptions to the general rule. On the north side of Kerr lake the formation is of diabase, while on the south side it is of Huronian slates and conglomerates. These Huronian measures extend out under the lake, dipping gently to the north to a lateral distance not yet definitely determined, owing to the lack of development under the deepest parts of the lake. Probably the contact between the diabase and Huronian or else between the former and the underlying Keewatin formation occurs somewhere near the center of the lake. Kerr lake originally covered 45 acres. Of this the Kerr Lake Mining CO. owned 12, the Drummond Mine, 7, and the Crown Reserve Mining CO. the remainder of 26 acres. Since the latter's property was originally entirely under water, it was necessary for the owners to make land for buildings and shaft room. Accordingly, in 1908, a trench was blasted out which deepened the outlet, and lowered the lake 8 ft. As many of the rich veins of the Crown Reserve and Kerr Lake companies were under water, mining was pursued under some disadvantages. Fortunately the rock is tight and solid, and mining has been done sufficiently fay from the surface to avoid unnecessary risk. Nevertheless, there was always danger of encountering open seams, through which too large a flow of water for comfortable working might come. Careful soundings were made over the veins, by means of steel-shod pipe, through mud and water to bedrock, but there was sometimes an element of uncertainty as to whether actual bedrock had been reached. Furthermore, the necessity of leaving safe backs between workings and lake tied up large quantities of ore, and made development in certain

Jan 1, 1915

By Robert Livermore

IT has been a noteworthy feature of the Cobalt camp, that many of the valuable ore deposits have been covered, wholly or in part, by small but usually deep lakes, such as Cobalt, Cart, and Peterson lakes, and with the subject of this article, Kerr lake. In the early days of the camp many water-covered areas, which have since been found to be valuable, were ignored, or neglected by owners, or simply perfunctorily staked for water rights. Kerr lake is probably the evidence of a fault or sharp fold, having a general east and west strike, and containing a series of ore-bearing fissures, parallel generally to the strike, although having some notable exceptions to the general rule. On the north side of Kerr lake the formation is of diabase, while on the south side it is of Huronian slates and conglomerates. These Huronian measures extend out under the lake, dipping gently to the north to a lateral distance not yet definitely determined, owing to the lack of development under the deepest parts of the lake. Probably the contact between the diabase and Huronian or else between the former and the .underlying Keewatin formation occurs somewhere near the center of the lake. Kerr lake originally covered 45 acres. Of this the Kerr Lake Mining Co. owned 12, the Drummond Mine, 7, and the Crown Reserve Mining Co. the remainder of 26 acres. Since the latter's property was originally entirely under water, it was necessary for the owners to make land for buildings and shaft room. Accordingly, in 1908, a trench was blasted out which deepened the outlet, and lowered the lake 8 ft. As many of the rich veins of the Crown Reserve and Kerr Lake companies were under water, mining was pursued under some disadvantages. Fortunately the rock is tight and solid, and mining has been done sufficiently far from the surface to avoid unnecessary risk. Nevertheless, there was always danger of encountering open seams, through which too large a flow of water for comfortable working might come. Careful soundings were made over the veins, y means of steel-shod pipe, through mud and water to bedrock, but there was sometimes an element of uncertainty as to whether actual bedrock had been reached. Furthermore, the necessity of leaving safe backs between workings and lake tied up large quantities -of-ore, -and- made development- in certain

Jan 7, 1914

By Éder Domingos Oliveira, Carlos Antonio Morais, Ruberlan Gomes Silva

The objective of this work was to assess the influence of parameters such as residence time, solid content, and temperature in a process route towards purified rare earth liquor which could be adequate to produce marketable rare earth products. The steps of the process route proposed in this study comprised water leaching of a sulfated rare earth ore and removal of impurities through the purification of the rare earth sulfuric liquor, which is the product of the water leaching step. The sample used in the water leaching experiments was a sulfated rare earth ore produced by the reaction of a beneficiated rare earth ore with sulfuric acid followed by roasting at 700 °C for 2 h in a muffle furnace and cooling at 20 °C. The water leaching experiments showed that the best condition comprised a residence time of 30 min, temperature of 20 °C, and solid content of 10% w/w, resulting in a recovery of 71.3% of the rare earth elements (REE) and a lower impurity extraction, e.g., only 15.0% of Fe, representing the main impurity in the rare earth ore. Even when presenting low amounts of impurities, the rare earth sulfuric liquor still needed to be purified before it can be used in the production of marketable rare earth products. In the study, the rare earth sulfuric liquors were purified in two consecutive neutralization steps, first by dosing the limestone pulp to pH up to 3.5 and secondly by adding the lime pulp to raise the pH from 3.5 to 5.0. The application of two consecutive purification steps at 20 °C resulted in the highest overall REE recoveries (65.4%) throughout all the steps of the process, being 71.3% in the water leaching, 99.6% in the partial purification with limestone pulp, and 92.1% in the purification with lime pulp at 20 °C. Under this condition, there was a total elimination of Fe, P, and Th and a significant reduction in the Al, U, and sulfate concentrations in the purified rare earth liquor. The purified rare earth liquor attained through the experiment is expected to be amenable to be used in the production of marketable products due to the lower impurities/REO mass ratios.

By Brook W. H

The history of the reduction of iron from its ores is lost in antiquity. Archreologists state that man has practised it for over four thousand years. Until comparatively recent times charcoal was the principal reducing agent for making metallic iron from its various oxidised ores. Coke was first used by iron founders in England in 1619, and its use spread gradually throughout the civilised world from then onwards. Indeed, the destruction of forests to obtain the requisite charcoal made the iron founders, both in England and America, extremely unpopular, and this hampered the iron industry to a considerable extent, until coal provided the metallurgist with a new and better fuel--coke which is used by most pyro-metallurgists today. The transition from wood to coal products for iron smelting had a profound effect on the location of iron works. Particularly was this the case in America where, in the infancy of the industry, practically each of the early States had its own iron producers, but as coal and coke replaced charcoal, there came about a concentration of the iron and steel industry around the coking coal deposits, and so today there are huge steel plants clustered around the soft coal basins in Pennsylvania and Illinois.The development of the iron and steel industry in Australia, which is comparatively recent by world standards, has been cradled in the coal basins of New South Wales. The first steel works of any consequence in this country were built near Lithgow, close by the mines which produced the coal for coking.Incidentally, these works were rather unique in that the three essential materials for iron and steel manufacture-ore, coal, and limestone-were found virtually at the works' back door.When, in 1912, the location of the proposed steel works for The Broken Hill Pty. Co. Ltd. for smelting ore from Iron Knob in South Australia was considered, the site ultimately chosen was at Newcastle, almost on top of seams of satisfactory coking coal. One of the major factors which had determined the site of the steel works was the fact that, about 1912, three tons of coal were required to produce one ton of finished steel product, and the relative economics of ore and coal haulage were much in favour of bringing ore to the coal.

Jan 1, 1947

By Kramadibrata S

Over the years many attempts have been made to develop a means of assessing the excavatability of rock. Most of them are based on an empirical rating system while some authorities still propose the use of seismic velocity as a direct predictor of the rippability of a rock mass. On the other hand there are a number of classical models which have been developed to define the cutting force required at the pick or tooth of continuous miners.While these methods have been applied with various degrees of success to the design of excavation systems there is no generally acceptable method of defining the excavatability or cuttability of a rock mass in terms of the machine power required to generate a particular rate of production.The authors have attempted to overcome this deficiency by recording the intact and rock mass properties at a number of surface mines where continuous miners were in use and to use this data to examine the relationships between the relevant dimensionless groups developed from a dimensional analysis of the problem.The results obtained are encouraging and it appears that this new approach to the problem could lead to civil and mining engineers having a means of predicting the cutting power required to excavate rock at a given rate at any site.While this paper does not include a review of past work it refers to a number of key publications which will give the reader an excellent introduction to the subject.It is important to point out that the research work reported in this paper is part of a doctoral research project at Curtin University of Technology on the Influence of Intact Rock and Rock Mass Properties on the Design of Surface Mines. One of the major objectives of this project is to explore the possibility of developing a technique for predicting the excavatability of rock in surface mines at the design stage.This paper therefore commences with a brief reference to the continuous surface miners in use at the test sites followed by some original work by the authors in applying Dimensional Analysis to analyse the macro-performance of mechanical excavators. Following an analysis of the experimental data the authors validate the wider use of the model by referring to its successful application in predicting the power required to excavate rock by continuous surface miners at other sites.

Jan 1, 1995

By Bruce Hebblewhite

"Historically there have been a number of different hypotheses and empirical models developed in an attempt to describe the nature of fracturing above longwall panels in underground coal mining. The motivation for such models varies, ranging from understanding the impact of mining on surface subsidence to back-analysis of caving behaviour in the immediate roof behind the longwall face. One of the most critical motivating factors that is taking on increased importance in many coalfields is the need for better understanding, and hence prediction of the impact of mining on overlying strata, particularly strata units acting as aquifers for different groundwater horizons.This paper reviews some of the major prediction models in the context of observed behaviour of strata displacement and fracturing above longwall panels in the Southern Coalfields of New South Wales, south of Sydney. The paper discusses the parameter often referred to as “height of fracturing” in terms of the critical parameters that influence it, and the relevance and appropriateness of this terminology in the context of overlying sub-surface subsidence and groundwater impact. The paper proposes an alternative terminology for this parameter that better reflects what it is and how it is used. The paper also addresses the potential role of major bedding shear planes mobilised by mining and their potential influence on overlying subsidence and groundwater interference.INTRODUCTIONKnowledge of the detailed nature of rock deformation and failure above any form of large-scale underground mining is always going to be limited to interpretation from a very incomplete set of data. It is extremely difficult, if not impossible, to directly measure the detailed nature of the rock failure, fracture networks, and deformational behaviour above an extracted mining area. Limited techniques such as borehole extensometry can provide some evidence of relative or incremental deformation in the direction of the borehole (usually vertical). However, such data cannot assist below the horizon where full caving has caused major rotation and dislocation of rock blocks and effectively destroyed the instrumentation borehole. Above such a horizon, the data is only valid along the axis and in the direction of the borehole, and to the level of detail defined by the extensometer anchor spacing intervals. It is often also restricted by major shear planes intersecting the extensometer borehole and cutting off the instrumentation below the level of the shear horizons."

Jan 1, 2019

By Craig Collins

The objective was to develop a Real Time Drilling Display System for Rotary Roof Bolting using Analyzing Software to display mapped holes as they are being drilled in real time. The Analyzing Software was developed to analyze and interpret drill sensor data produced during a rotary drilled roof bolt hole by using algorithms and drill sensor data trends to find and locate roof separations or voids. After this software was validated through lab and field testing, its application needed to be tested in a production environment. However, the current method used to record the drill sensor data proved to be cumbersome for the mine to use on a continuous production basis. With different more practical mediums being explored and considered, a Drilling Display System was developed that will both map holes in real time and record the data for later review. With the Analyzing Software's current level of accuracy and with the convenience offered by the real time Drilling Display, the Drill Display System was implemented along with a test mine's current roof control plan and tested as a roof separation "Early Warning System". After installing the Drilling Display System onto a machine, input and suggestions from the mine environment were considered and design changes were made. The real world application provided a design direction that made the information produced by the Analyzing Software both convenient and significant for the test mine. Also, using the Drilling Display System increased the number of drilled holes that were both mapped and recorded. The results from studying these recorded data files lead to refinements in the current algorithms used by the Analyzing Software which increased the overall accuracy and capability of the system A fairly accurate representation of void or separation locations in the mine roof can be determined fromsensor data recorded during production bolting cycle. The immediate feedback on the local roof condition that the Drilling Display System provides, amounts to a quicker and safer alternative to the scratch hole and the video bore-scope examinations that arc performed after the bolting is completed. This project has shown that making the alternative method of examining the roof condition practical for the mine to use is just as important as the accuracy of its results. Based on the results of this field test, this or other methods of reviewing the analyzed drill hole data will be explored.

Jan 1, 2004