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By H. K. Fredrickson

"Midwest Joint Venture (MJV) owns a high-grade uranium deposit in northern Saskatchewan. The deposit is located too deep below surface to be mined economically by open pit methods, and as a consequence, present plans are that it will be mined by underground methods. High-grade uranium ore of the type at MJV, encased in weak, highly altered ground and with radon-rich water inflows, has not before been mined by underground methods.The test mining phase of the project, completed in 1989, had three objectives: 1. to evaluate radiation protection requirements associated with the handling of large quantities of radon-rich water and mining highgrade uranium ore in an underground environment; 2. to investigate the quantity and quality of water inflows into the mine; and 3. to investigate ground conditions in and around the ore zone as an aid in determining the production mining method to be used.With information gained from the test mining project, a mining method for the production mine has been devised. Level plans have been drawn up, ventilation system designed, pumping arrangements made and methods of ore handling considered. All this is to be done in a manner that will be safe for those doing the work underground. Some of the mining methods planned are felt to be unique in that they are designed to cope with mining problems not known to have been encountered before. New problems underground have required new methods to handle them. Remote drilling, blasting, mucking and backfilling form the basis of the planned mining method."

Jan 1, 1992

Jan 1, 1937

By Thomas Schwind, Benedikt Grau, Angelos Gakis

"In Crossrail’s Farringdon station, compensation grouting was successfully utilized to mitigate the surface settlements induced by the tunneling works. The main focus of this study was the platform tunnels, which were approximately 300m long each and were enlarged to platform size using sequential excavation method (SEM) from the existing TBM running tunnels, initially supported using sprayed concrete linings. The rear grouting injections were carried out at a typical distance of minimum 5m above the completed, fully strengthened shotcrete lining hence inducing some additional stresses. In-tunnel monitoring using surveying targets was performed systematically to ensure that no excessive distortion of the initial shotcrete lining occurred. 3D finite element analyses were utilized to back calculate the actual “effective” pressure that was imposed on the completed shotcrete rings against the additional measured in-tunnel displacements due to the compensation grouting episodes. Additionally an assessment of the theoretical, “maximum” allowable grouting pressures that would lead to overstressing of the lining is presented. INTRODUCTION As part of the back analysis performed for the SEM tunnels in Farringdon station, an investigation of the effect of compensation grouting on the two constructed platform tunnels was undertaken. The platform tunnels were 10.6m high by 11.4m wide and approximately 300m long each. They were enlarged to their final shape from the existing TBM tunnels using SCL methods. This study utilized 2D and 3D finite element (FE) analyses in order to back calculate the observed tunnel crown displacement, assessing the equivalent grouting pressures during the rear injection episodes of the concurrent grouting operations. The results provide a useful reference for the order of magnitude of the grouting pressure load cases that can be applied in the numerical simulation of similar cases. Part of the results presented in this paper, were derived by the second author as part of his Master of Science thesis in Technical University of Graz."

Jan 1, 2016

By Roland B. Drugge

The adaptation of pelletizing processes involves the coordination of various factors such as temperatures, gas flows, productions, etc., in order to obtain optimum performance. The consequent problems are considerable due to the large number involved as well as the lack of knowledge of their correlation. It is therefore essential to establish a fundamental theory which takes into account the heat transfer phenomena and the secondary factors which are relevant to the properties of the product. Straight grate and grate kiln pelletizing processes involve heating of a stationary bed on a travelling grate. The heat transfer conditions are therefore of primary importance. This is why this article pays particular attention to conditions which obtain when a stationary bed is heated, in which certain special reactions occur. The conclusions drawn from this penetration of the heat transfer phenomena can be expressed in postulates acting as guide-lines for the establishment of proto-typal processes. Examples of elaborating such processes are also given in the article. Some operational experience at LKAB are also given.

Jan 1, 1975

By Frank Mucciardi, Louca Gourtsoyannis

"Low and inconsistent recoveries of aluminum in the production of aluminum-killed steels triggered an extensive research program on the aluminum dissolution kinetics at McGill University. This research program led to the understanding of the melting kinetics and the hydrodynamics of making additions to liquid steel. It showed the necessity of gradually releasing liquid aluminum at the bottom of the ladle to increase aluminum recoveries.This knowledge was verified in practice at Sidbec-Dosco, where a prototype aluminum wire injector was developed and installed for the first time in Canada. The industrial application showed the validity of the theoretical and experimental results and provided further data for the mathematical modelling of the process.The development of the theoretical understanding, the industrial results, and the interaction between university and industry are presented.IntroductionAluminum is an extremely effective deoxidizer of molten steel; nevertheless, an efficient method of adding it to steel has eluded steelmakers for many years. It is well known that the recovery of aluminum used in the production of aluminum killed steel is both low and eratic, especially when the aluminum is added in the form of bars or lumps from chutes into filling ladles, a practice still common in many of today's steelmaking operations.As the density of aluminum is only about 40010 that of molten steel, it is conceivable that loose bars of aluminum simply ""thrown"" into a filling ladle will resurface before any appreciable dissolution of the aluminum has occurred, particularly if the depth of penetration of the additions is small. Before one can conclude that this is indeed the case, it is necessary to know the melting kinetics of an aluminum addition. Back in 1972, this information was not available; hence an extensive research program which focused on melting kinetics was initiated at McGill University."

Jan 1, 1982

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Jan 1, 1985

By F. Rosenblum, P. Spira

A laboratory test has been developed for assessing the patential hazard from self-heating of sulphides. The technique consists in measuring the rate of self-heating, from near ambient temperature, of a 1-kg sample in an insulated chamber controlled at standard atmospheric conditions. The rate of oxygen consumption is monitored simultaneously. The paper describes the apparatus and test procedure, and reports data on various ores and mill products. Tailings known to have caused problems underground when used as backfill have shown heating rates of over 3°C/h. A preliminary hazard rating guide for backfill based on the known behaviour of several sulphide materials is presented.

Jan 1, 1981

By Marcos Natan da Silva Lima

O presente estudo tem como objetivo avaliar a resistência à corrosão de dois aços alto teor de Manganês com relação à liga 9% Níquel em solução eletrólito 0,001M de NaCl. Para esta pesquisa foram utilizados dois aços com teores de Manganês de 22%, 27% e uma liga 9% Níquel que é amplamente utilizada na indústria petroquímica. Os ensaios realizados para comparação da resistência à corrosão foram o monitoramento do potencial de circuito aberto (OCP) e polarização linear anódica. Foi escolhida a solução de NaCl pela presença da maioria dos tanques de armazenamento e transporte de gás natural se localizarem próximos à região marítima portuária. Foram utilizadas as técnicas de Microscopia Eletrônica de Varredura (MEV), Microscopia óptica e Espectroscopia de Energia Dispersiva de Raios-X (EDS) para a caracterização dos produtos de corrosão. Os resultados obtidos mostraram que os aços alto Manganês apresentaram valores similares de resistência à corrosão, mas ainda inferiores em relação ao aço 9% Níquel.

Oct 30, 2017

By D. A. Sloan

THE Young Davidson mine is of particular interest because of the extremely low costs and also because of the novel method devised for pillar recovery. The outstanding feature of the mining system is its simplicity and regularity. The successful operation of this property was de-finitely assured by the rise in price of gold in 1934. The mine is situated in the southeastern part of Powell township, about forty-five miles southeast of Timmins, Ontario. The local town, two miles distant, in Matachewan, which is twenty-five miles northwest of Elk Lake, the terminus of a branch line of the Temiskaming and Northern Ontario railway. Mining is confined to one of a group of five claims. The ore occurs in a body of syen.ite porphyry which is intruded along a contact of Timiskaming conglomerates with Keewatin altered ruffs and greenstones. This intrusion extends in a north-easterly direction from Davidson creek for 3,300 feet, of which length 700 feet lies within the Matachewan Consolidated property. The maximum width of the mass is 700 feet; it tapers off as it extends east-ward. The Orebody.-The syenite porphyry is of three types, probably all differentiates of the same magma, but only one of them is gold bearing. This is a cherry-red, coarsely crystalline type which forms the central part of the mass. It is more brittle and more fractured than the others, and consequently has allowed easier access to the gold-bearing solutions. The gold has been brought in by a branching system of quartz veins and along minute fractures. It is mainly in the form of very fine 'free' gold, but also occurs in pyrite, which is the chief metallic mineral.

Jan 1, 1942

Computerisation of mine planning activities, specifically in the areas of geology, design and surveying, has been in progress for a number of years in underground metalliferous mines. The combined capabilities of hardware and software have only recently reached the levels necessary to fully support integrated mine planning computer systems in a graphical three dimensional environment. Despite the apparent capabilities of systems now on the market, many sites retain significant parts of their older paper based systems, and employ a mix of two or more software packages. This paper examines what underground metalliferous sites are doing in this area in Australia, and discusses the factors which are inhibiting full implementation of fully integrated computerised systems. Future directions for further developments in the field are also briefly discussed.

Jan 1, 1995

By Mario E. Rossi

Las simulaciones condicionales geoestadísticas se utilizaron como herramientas de control de leyes en operaciones diarias que evalúan la incertidumbre y el riesgo de las reservas mineras en estudios de factibilidad, así como el potencial de mineralización en ciertos ambientes geológicos. Otras aplicaciones incluyen evaluación de recursos y reservas recuperables (dilución), clasificación de recursos y reservas, y optimización de espaciamiento de sondajes. En ese sentido, el presente trabajo señala los aspectos de los esquemas de clasificación de recursos y reservas utilizados en la industria, el cual está basado en el modelo de incertidumbre que las simulaciones condicionales proveen. Asimismo, se analiza el riesgo que este modelo de incertidumbre implica para sub-regiones (unidades) de extracción del yacimiento, como las definidas por fases de extracción mineras o períodos de tiempo.

Jul 17, 2000

By Chris Mark

Pillar recovery continues to be a significant ground control hazard. During the past decade, 10 miners were killed during pillar extraction operations in southern West Virginia. Studies conducted during the past decade have identified a number of "risk factors" that can be used to evaluate pillar extraction plans: Cut sequence Final stump Timber or Mobile Roof Supports Roof bolting Intersection span Depth of cover Roof quality Age of workings For each of these factors, rock mechanics science suggests which alternative would be expected to be more risky. For example, numerical models were used to evaluate different cut sequences, and indicated that less roof convergence occurred with the outside lift method than with the Christmas tree in the particular environment simulated. For many of the risk factors, accident statistics confirm the science. One finding was that currently almost 70% of the retreat coal in southern WV is being mined with MRS. In contrast, timber supports were used in 70% of the past decade's pillaring fatalities. This paper discusses each of the risk factors in turn, presents the relevant accident statistics, and shows how the risk factors can be combined to estimate the overall hazard. It also addresses the use of pillar design to minimize the risk of global stability hazards including squeezes. massive collapses, and bumps.

Jan 1, 2002

By Garry Keizer

The Cortez gold mine is located in Lander County, NR, about 130 km (80 miles) southwest of Elko. Cortez Gold Mines (CGM) is a joint venture between Placer Dome US Inc. (60% operator) and Kennecott Corp. (40%). The claims that supply oxide and refractory ore feed to the Cortez milling facility are located in both Lander and Eureka counties. These include the Cortez and Horse Canyon deposits, all operated by the joint venture. Other deposits outside the CGM claim area include Fire Creek, Buckhorn, Gold Bar and Tonkin Springs. These deposits all occur along the Battle Mountain-Eureka Trend. History Mining in the Cortez district began in 1862 with the discovery of high-grade silver ores. Grades in the hundreds of ounces per ton occurred in an area 7 km (4.5 miles) southeast of the current CGM mill site. Underground mining through the Garrison and Arctic tunnels provided the bulk of the ore production between 1864 and 1895. To a lesser extent, the Mill Canyon area also supplied ore until 1930.

Jan 1, 1993

By John M. Tinsley, Calvin D. Saunders, H. K. van Poollen

In the past few years much con-sideration has been given to the evaluation of the effect of hydraulic fracturing on the productivity of wells. Generally, these studies included the evaluation of fracturing materials, fracture extension and formation damage due to the use of various fracturing fluids. Only little consideration has been given to the characteristics, and in particular the flow capacity, of the fracture itself and its effect on well productivity. This paper presents the results of laboratory investigations pointed toward fie evaluation of the efficiency of various fracrures with special emphasis on the flow capacity of these fractures. Data presented in this paper are the results of both an electrical model study and physical testing. Under consideration are (I) effect of overflush, (2) premature production of well after treatment, (3) "tailing-in" with coarse sand near the end of the treatment, (4) effect of propping agent size and concentration, (5) reduction in effective frac-ture permeability caused by formation caused by formation fines, silt and clays, and (6) effect of various fluids on formation strength and competency. The results of this investigation indicate that the flow capacity of a fracture is affected by any or all of the various parameters mentioned above. The authors believe that a better understanding arid utilization of these factors should result in more efficient formation fracturing. INTRODUCTION Hydraulic fracturing has become almost a standard practice of many companies for stimulating production from old and new wells. Although most companies utilize this service, techniques of application vary widely between companies and areas. Probably too often when a well in an area responds favorably to a particular technique all future wells in the same area are treated in a similar manner. Possibly a modification of the technique would result in a further production increase. Variables, of which many are extremely difficult to evaluate from field results, hamper the selection of procedure changes. Attempts are being made by a number of organizations to analyze statistically treating techniques from production data. This is a very worthy and necessary approach but very possibly laboratory investigations may aid in evaluating some of the variables which tend to affect the results of a fracturing treatment. Some of the factors cannot be studied from practical field experience and only laboratory tests can show the possibilities which might exist. One of the factors which appears to be of major concern today is the flow capacity of the created fracture and how it can be changed. Papers on this general subject have been primarily concerned with the size of propping agent and the extent of fracture. Papers have also been written on the possible permeability damage to formations by fracturing fluids. In addition, it might be possible that another type of flow restriction is prevalent. This would be a restriction of flow through a sand-packed frac- ture caused by foreign materials integrating within the propping agent. This paper presents preliminary data obtained in an attempt to evaluate the effect of some factors affecting flow through sand-packed fractures. No attempt is being made to offer a fracturing technique adaptable to all areas and conditions, but to furnish data tending to show the possible effects which might be caused by variations in procedures and materials. PROPPING AGENT PERMEABLLITIES A hydraulically induced fracture containing sand as a propping agent may theoretically be classified as a packed-sand system. The flow of fluids through such packed systems has been the subject of much research. Although there have been numerous methods proposed for the evaluation of such systems, most writers agree on the general properties affecting their flow capacity or permeability. These properties include porosity, particle size, sphericity and the roughness of the particle. In some methods of evaluation . the particle size and sphericity terms are combined to produce an equation which is a function of the surface area of the particles. In this study the permeability of various fracturing sands was both measured and calculated. The apparatus for the permeability measurements consisted of a 52-in. Lucite tube with a 2.5-in. ID. A screen and drain plug were fitted in the bottom of this tube to retain and hold the sand in place while allowing fluid flow. Two pressure taps consisting of thin, highly perforated

The great Bingham Canyon mine and mill complex of Magna, Arthur, and Bonneville had their beginning in a 300 ton per day operation which was commenced by Daniel C. Jackling and his associates in 1904 as the Utah Copper Company. This small plant was quickly expanded to 500 tons per day and eventually Kennecott, the Utah Copper successor, was processing about 45,000 tons per day of ore in each of its mills, Magna and Arthur. Finally, in 1966, the Bonneville crushing and grinding plant was added to produce a slurry of ground ore for flotation concentration at the expanded Magna and Arthur mills. Bonneville is one of the few railroad dump primary gyratory crushing plants constructed in recent years; another which is described in this Monograph is Cuajone in Peru, started in 1976. Two others are the Climax Henderson unit and the Phelps Dodge Tyrone crusher, which has been fed by trucks since it commenced operation. The Bonneville, Henderson, and Cuajone crushers are all located considerable distances from their mines, hence the use of rail haulage. Shovels in the Bingham pit are 8 to 15 yd capacity; and ore is sent to the Bonneville crusher in 100 ton rail cars, where they are dumped, two at a time, by a hydraulically driven rotary dumper. Ahead of the 54 in. by 74 in. Allis-Chalmers gyratory are two grizzlies, each 27 ft 9 in. (8.46 m) long with bar spacings of 4 ½ in. (114 mm) at the feed end and 6 in. (152 mm) at the discharge end. Grizzlies are sloped at 38. Oversize rocks are re- moved from the crusher pocket by an overhead crane, and the crusher produces a nominal 6 in. (152 mm) product. Rail cars are dumped in the direction away from the crusher, and water sprays at the dumper level and in the rock boxes are used for dust suppression. There are two 72 in. by 19 ft (1.83 m by 5. 8 m) apron feeders under the 500 st cr'ushed ore pocket and two more under the grizzly undersize pocket. Feeder speeds are variable up to 72 fpm. Grizzly undersize is dropped to the feeder pocket by three shelves forming a rock ladder attached to the pocket walls. Each pair of apron feeders, one for crushed ore and one for grizzly undersize, discharge to a 72 in. (1. 83 m) picking belt for tramp removal; this belt feeds one of two 54 in. (1.37 m) feeder conveyors which in turn discharge to a 60 in. (1. 52 m) takeaway belt operating at 175-300 fpm (53.3 - 97.4 mpm). Both picking belts are equipped with metal detectors and tramp iron magnets, and a separate 30 in. (760 mm) conveyor is used to remove tramp material.

Jan 1, 1978

By R. E. Head

Under existing economic conditions, the treatment of gold ores occupies an outstanding position in metallurgical activity. The increased price of gold has automatically brought about a reclassification of old dumps and low-grade orebodies, and many that formerly were considered valueless are now being vigorously investigated and worked. This situation has developed problems that have taxed the ingenuity and ability of metallurgists to the utmost. Some gold ores give a high extraction of gold by standardized methods of treatment; other ores of apparently similar character yield but a relatively small percentage of their gold when treated by the same processes. Surface Contamination The experience gained by microscopic study of numerous gold ores and tailings has shown that there are pronounced differences in the physical characteristics of the minerals composing them, more especially of the gold. Repeated studies of tailings from flotation and cyanidation of gold ores has established the fact that surface contaminations on gold particles are often directly responsible for high gold losses; comparison of gold particles isolated from flotation concentrates and the resulting tailings have showed that the clean gold has been recovered and the tarnished or contaminated gold invariably lost in the tailing. Obviously, it is not possible to make such a comparison of gold ores treated by cyanidation, as the clean gold has been taken into solution; but when the gold found in cyanide tailing has a tarnished or coated surface, one may infer, with a reasonable degree of certainty, that the clean gold has been extracted. This premise is supported by experimental evidence obtained by isolating particles of tarnished gold and exposing them to cyanide solution in small parting cups. In one such experiment, tarnished gold particles picked from a cyanide tailing showed but slight evidence 01 dissolving at the end of 27 days. In this test, a cyanide solution of 1.6 11). per ton was used and the leach solution was decanted arid renewed every

Jan 1, 1939

By G. R. Morin, A. R. Leggett

"In 2010, closure of the Kidd metallurgical plants along with structural changes to the historical ratios of payable metals led to a review of the operating targets at the Kidd concentrator. Previous operating data indicated that there was potential for significant revenue improvements by lowering the copper concentrate grade. An extended plant trial was carried out in 2011 to quantify the revenue impact of lowering the copper concentrate grade from 27% to 25% Cu, as well as identifying the lower practical grade limit. Analysis of plant trial results indicated that the process change increased net concentrate revenue by $2 M to $3 M per year at current metal prices. Increased revenue was primarily associated with higher silver recovery to the copper concentrate while the marginally improved copper recovery was essentially offset by increased zinc losses and concentrate treatment and freight charges.Details of the plant test work results are discussed in this paper along with the financial analysis of the revenue improvements. Key areas covered are major revenue components of the concentrate contracts, validation of revenue improvements through plant testing and sensitivity of the improvements to market and plant variables.INTRODUCTIONThe Xstrata Copper Kidd concentrator commenced operations in 1966 treating ore from the nearby Kidd volcanogenic massive sulphide deposit in Timmins, Ontario (Thompson, Amsden, & Chapman, 1978). Over the years, up to four separate grinding-flotation divisions were in operation with throughput rates peaking at 4.5 million tonnes per year of ore in the mid 1980’s (Kelebek & Francis, 1998). As the mine production decreased and grinding throughput increased operations were gradually cut back and since 1998 two grinding-flotation divisions have been in operation treating Kidd mine ore. One of the idled divisions has been intermittently used for custom milling operations, most recently the Ni/Cu Montcalm deposit, which was milled from 2004 to 2009. Current budget throughput rates, limited by mine production, are 2.4 million tonnes per year of ore grading 1.7% Cu, 4.3% Zn and 55 g/mt Ag. Concentrate production rates are 150,000 tonnes per year of copper concentrate and 155,000 tonnes per year of zinc concentrate. The Kidd mine is the deepest base metal mine in the world with mining operations down to 9600 ft.In 2010 the copper and zinc plants at the Kidd metallurgical site, which previously treated the concentrate production from the Kidd concentrator, ceased operation due to a variety of economic and market factors. Following the metallurgical plant closures the Kidd concentrator products were shipped off site for treatment. The majority of the copper concentrate is now processed at the Xstrata Copper Horne smelter in Rouyn-Noranda, Quebec with the majority of the zinc concentrate processed at Canadian Electrolytic Zinc Ltd. in Valleyfield, Quebec."

Jan 1, 2013

By T. D. Mueller, H. Dykstra

This paper describes a method for obtaining compositions of gas and oil phases in equilibrium with each other at a given reservoir temperature and pressure as lean or enriched gas is injected into a reservoir oil not in equilibrium with the gas. The method is based on calculation of equilibrium constants, K, from the equation log Kp = A + B F, where p is pressure and F is component characterization factor. A and B are constants that vary with a composition parameter. The paper explains how A and B could be determined with a minimum amount of data. The method consists of calculating equilibrium constants as a function of liquid composition and using the K's to calculate changes in compositiort of injection gas as it successively contacts original resetvoir oil at the gas-oil front and the change in composition of reservoir oil as it is successively contacted by the injection gas in the vicinity of an injection well. An iterative process is required. Two examples of the method are given which show how oil and gas compositions and properties are calculated with the aid of a computer. INTRODUCTION The calculation of recovery from lean-gas or enriched-gas injection into a saturated or under-saturated reservoir oil requires a knowledge of how phase properties change in a reservoir between injection well and producing well. To determine the phase compositions or properties experimentally would be difficult and time consuming, especially if one wanted to look at the effect of various injection gas compositions. For some fields an engineer would want to make preliminary calculations on the economics of gas injection before he makes the decision to spend a corrsiderable sum of money for laboratory work to obtain the necessary fluid properties. A method for calculatine compositions and properties based on a minimum of laboratory data is described. The minimum amount of data required is injection gas and reservoir oil composition and injection gas solubility as a function of pressure for an undersaturated oil, or the compositions of mixtures of reservoir fluid with two or more enriched gases, which mixtures have a bubble-point pressure that is the same as the expected average displacement pressure. On the other hand, if the economics appear to be attractive, more laboratory data can be obtained in the form of equilibrium constants for three or more mixtures of the reservoir fluid and injection gases. The method is based on the calculation of K's from the equation, log Kp = A + B F, where K is equilibrium constant, p is pressure and F is component characterization factor. This method of plotting K data by means of the above equation is described by Hoffman, Crump and Hocott.1 A and B are constants that vary with a composition parameter. It is shown how experimental data are used to determine how A and B vary with composition. BASIS OF METHOD EQUILIBRIUM CONSTANTS Equilibrium constant (K) data are required for this method. The K of each component is obtained from an equation that gives the product of K and pressure as a function of component characterization factor. The characterization factor F = b(l/Tb — l/T), where b is a constant characteristic of a given component, Tb is the boiling point in OR at atmospheric pressure of the given component, and T is the temperature of the system in OR. This method of plotting K data is described by Hoffman, Crump and Hocott.l The straight-line equation of the log Kp - F plot is log hp = log Kop +BFi, where Kop is at F =0, i designates the component, and B is the slope of the line. The factor F can be

Jan 1, 1966

By Paul L. Russell

Introduction As the US and world population increases, so does the problem of supplying expanding energy needs. Early requirements worldwide were mostly met by abundant forest vegetation. Man's large scale use of minerals-energy resources is a fairly recent development. It occurred only during the past 200 years. Coal was the first mineral fuel to be used extensively. It was not a large factor in US energy production until the beginning of the 19th century. Petroleum and natural gas use is about 125-years-old. Electricity generation by falling water is about a century old. Atomic energy is only about 30-years-old. Extensive use of it is still in the future. Solar energy is in its infancy. Other energy sources such as wind, peat, and geothermal energy are minor and primarily of local importance. There were three main features of the fuel supply picture during the past 125 years. By 1900 coal replaced wood as the principal energy source. Since 1900, oil and gas have captured many coal markets. And after the mid-1970s, there was a resurgence of coal usage. (Figs. 1 and 2). Wood Until the advent of mineral fuels for energy, the major global energy source was wood. But populations increased and became centered in towns and cities. The problem of a sufficient energy supply first developed as nearby, easily obtained wood was exhausted. This problem became more acute as industries such as shipbuilding, smelting, and pottery making, became more widespread. Results of wood use as an energy source and building material are starkly evident in areas of the world where civilization is oldest. Entire forests were consumed and have never been replaced. In the US, wood supplied 90% of all fuel requirements until about 1850; and more than 50% until after 1880. By 1900, wood still supplied 20% of all energy uses. In the mid-1970s, energy costs increased, due largely to the oil embargo. This increase resulted in an increased wood demand for domestic heating. While large quantities of wood were being used, its share of the total energy market was still less than 3%. By 1983, wood was becoming increasingly scarce and more costly near most large US cities. Wind Wind powered ocean vessels for many centuries. Sailing ships were still common in 1900. Windmills provided an early, cheap energy source and continue to be used today on farms. Design and development of windmills for electricity generation is being conducted. Interest in wind power is growing worldwide where availability of wind currents warrants the investment. Also, US regulation requires power companies to "buy back" electricity generated by windpower in excess of that used by the windmill owner. Coal By the 1600s, England was producing, 900 kt/a (1 million stpy) of coal for heating and other uses. The first coal-fired blast furnace for iron smelting was developed in 1621. Coke-making was developed in 1755, and Watt invented the steam engine in 1769. Illuminating gas made from coal was introduced in London in 1800. During this same 200-year period, France and Germany developed coal production, but on a smaller scale than England. US, bituminous coal was mined near Richmond, VA, as early as 1748. Major bituminous coal-producing states in the early period were Pennsylvania, Virginia, Ohio, Kentucky, Illinois, and Maryland. West Virginia replaced Virginia as a major producer, when that state was formed in 1863. As the US expanded west, new coal deposits were found and exploited. Indiana, Alabama, Tennessee, Iowa, Kansas, Missouri, Washington, Colorado, Wyoming, Arkansas, Oklahoma (Indian Territory), Montana, New Mexico, and Texas had significant coal mining operations by 1900. Today the eight major coal producing states are Kentucky, West Virginia, Pennsylvania, Illinois, Ohio, Virginia, Wyoming, and Montana. These states accounted for about 80% of all coal (lignite, bituminous, and anthracite) produced in the US during the past several years.

Jan 8, 1984

By J. W. Spretnak, V. Kandarpa

Stress -induced ordering of carbon atoms is studied in a series of Fe-Mn-C alloys. A prominent peak is found in the vicinity of 280°C at frequencies of the order of 1.0 cps, with an associated activation energy of 37 kcal per mole. The height of the peak is linearly rekzted to the concentration of carbon in solution. The distortion of octahedral holes by manganese atoms appears to be predominant over carbon-carbon pair interactions. RELAXATION by stress-induced ordering of point defects is expected whenever the introduction of these point defects produces distortions which have a lower symmetry than that of the lattice. Under zero stress, the isolated point defects occupy the crystallographic-ally equivalent positions in the lattice, as these represent states of equal energy. However, if the defect sites are asymmetric, application of an uniaxial stress will split the energy states, and a redistribution of the defects among various states will take place. This is the case of the internal friction peak called the Snoek peak,1 resulting from isolated interstitials in bcc metals. The interstitial sites in this case have tetragonal symmetry. In the case of fcc and hcp lattices, such an effect is not expected from isolated point defects because of the symmetrical nature of the interstitial sites. However, internal friction peaks arising from interstitial diffusion have been reported both in hcp2,3 and fcc4-8 lattices. These peaks are often explained on the basis of stress-induced ordering of interstitial solutes, caused by the deviation of interstitial sites from their cubic symmetry through the presence of nearby defects. In the case of fcc lattices, evidence for interactions of both the substitutional-interstitial4,6,13 and interstitial-interstitial types5'798'14 have been presented by various investigators. The purpose of the present investigation was to study the internal friction peak attributed to the diffusion of interstitial carbon atoms in high purity austenitic manganese steels and to account for the peak in the light of the existing models. MATERIALS The Fe-Mn-C alloys used in the present investigation were made in two different ways, designated as Type I and Type 11. Type I alloys were made from high purity Fe-Mn alloys obtained in the form of 0.04- in.-diam wires from Materials Research Corporation, Orangeburg, N.Y. These alloys were carburized to different levels using gas mixtures of H2 and CH4 at 1000°C. Type I1 alloys were made in this laboratory starting with zone refined iron, spectrographically pure manganese, and spectrographically pure carbon. They were melted in an argon arc melting furnace and drawn into 0.04-in.-diam wires. All the wires were annealed at about 900°C for 3 hr prior to the internal friction experiments. After the measurements of internal friction, the phases in the samples were identified by X-ray diffraction and the carbon determined by the combustion method. EXPERIMENTAL PROCEDURE In the present work, a classical Ke-type pendulum was used. The details of the equipment were described previously by D. T. Peters.9 Dry helium at 40 torr was used in all the experiments. The internal friction, measured as the logarithmic decrement of the torsion amplitude of vibration was determined as a function of temperature, from ambient to about 500°C. The background internal friction was assumed to have the form of the exponential of the inverse temperature and was subtracted from the raw data. The height of the peak was measured at the position of the maximum in the plot of the internal friction versus temperature. The activation energies of the peaks were measured by the peak shift method. The internal friction values for an alloy were obtained as a function of temperature at different frequencies of vibration. The position of the peak changes with frequency, the higher the frequency the higher the peak temperature. The activation energy of the process associated with the peak is obtained using the formula

Jan 1, 1970