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By R. K. Mehta, S. K. Dhar, R. R. Mallepali

Introduction The incentives for developing a distributed form of the comminution model include: • for describing the heterogeneous comminution behavior of multiphase mineral systems (i.e. realistic ore bodies, such as complex sulfide ores consisting of chalcopyrite, galena, sphalerite, pyrite and siliceous minerals) and •for generating, for various locked and free particles, in situ, microscopic, breakage behavior such as the mode of fracture (intergranular vs. transgranular), the relative rates of breakage and the breakage distribution. However, the computational difficulty and the effort made in estimating a large number of parameters (Andrews and Mika, 1975; Wiegel, 1976; Herbst et al., 1985, 1988) do not justify the utility of this approach. Also, obtaining the results experimentally is very tedious and time consuming if using this approach. Although recent work in this area (Choi, Adel and Yoon, 1987; Mehta, Adel and Yoon, 1989, 1990) showed good results for liberated particles, the quality of the estimation and prediction was poor for locked particles. Attributing this to the complex breakage behavior of locked particles, it was concluded that additional constraints, and possibly modification in the breakage distribution functional form for locked particles, were necessary. However, further mathematical analysis revealed that the method previously used was in error. The purpose of this technical note is to present the correct solution methodology used in solving the general model equations of a multicomponent mineral system containing multiple classes of locked particles. Previous work and connection The general form of the comprehensive, distributed-component comminution model has been described as (Herbst et al., 1985) [d (m (0) -[ I - B ] S_ m(t) = A m(t) (1)] where [A] is the fundamental matrix/state companion matrix, commonly referred to, in the comminution literature, as the "mill matrix." Note that for a distributed-component comminution model, this is a full, general, real matrix, unlike a lower triangular matrix for a simple grinding model or a lumped-binary grinding model with one locked-particle class (Herbst et al., 1985, 1988). Also note that all of the underlined letters refer to matrices. Herbst et al. addressed the difficulty in arriving at a closed form of the analytical solution of Eq. (1) via similarity transformation [IT J inv (T)]]. This difficulty occurs for several reasons: •[A], as a general real matrix, may have complex eigenvalues and complex eigenvectors. •More importantly, there is no guarantee that a set of eigenvectors of [A] will turn out to be independent. An attempt was made to solve this problem by implementing a decomposition method. With this technique, the mill matrix was decomposed into a sum of lower and upper triangular matrices [(i.e., A = L + U)]. Then the solution was obtained using similarity transformation (Mehta, Adel and Yoon, 1989, 1990). [e(At) = e(L+U)t = e(Lt).e(Ut) _ [ TL Jt- (t) inv(T_L)] [-Tu Ju (t) inv(Tu)]] [But, unfortunately, e 1+U = e t . eu is true only for scalars and not for matrices (i.e., eL+u# eL- eU) unless they commute. In other words. eAt = e(L+u)t= ekt - eU' if LU = UL. Hence, this decomposition scheme must be replaced by a mathematically acceptable method known as the "splitting method"]

Jan 1, 1994

By T. V. Shockley

The Gulf Coast, from New Orleans, Louisiana to Brownsville, Texas, is one of the largest high energy consuming areas in the United States. This region has areas of very high industrial density. The lower Mississippi River from Baton Rouge past New Orleans has a high industrial density. Lake Charles, Louisiana, the Orange-Beaumont-Port Arthur area, Houston (4 separate areas) and Corpus Christi all have the same high industrial density. This situation developed due to good transportation, good labor conditions, good weather and the abundance of cheap energy-gas and electricity. Obviously the latter had the greatest impact. The availability of cheap natural gas made Gulf Coast sites very attractive to energy intensive industry such as metals and petro-chemicals. The area was also very attractive to refinery development due to the large supply (endless?) of crude oil available. As we all know, the energy situation has changed dramatically in the last five years. The cost of a MMBTU has risen from the 17¢ to 19¢ level to a price of $1.95 to $2.10. The price of interstate gas, as defined by the National Energy Plan recently passed by Congress, was approximately $1.98/MMBTU escalated to October, 1978. It goes without saying, that this has had a major impact on doing business in the Gulf Coast. As fuel prices soared, different companies responded in different ways; but overall, industrial energy consumption along the coast was reduced 20% to 25% from 1974 to 1975. Since 1975 production and energy use has been increasing, however, production at some plants is still significantly below plant capacity.

Jan 1, 1979

By Ann M. Hagni

One of the most costly factors in the production of aluminum is the consumption of energy in the form of electricity. Energy consumption is increased as refractories in the Hall-Heroult reduction cells are corroded by penetration of the fluxes cryolite (Na3AlF6) and villiaumite (NaF). To decrease thermal conductivity and refractory waste, protective barrier linings are installed. Dry powder barrier linings form stable reaction products at cell operating temperatures. Four dry powders composed of vermiculite, corundum, wollastonite, and anorthite are characterized mineralogically by X-ray diffraction, reflected light microscopy, scanning electron microscopy, and cathodoluminescence microscopy in this research project.

Jan 1, 1998

By J. Guzman

Introduction The Carlin Trend is North America's premier gold producing district. It is located in northeastern Nevada's Elko and Eureka Counties along a northwest trending belt about 65 km (40 miles) long and 8 km (5 miles) wide (Thorstad, 1989; Jones, 1989). This trend is the worldwide reference site for epithermal, sedimentary rock-hosted microscopic gold deposits. At least 19 deposits have been discovered to date, varying in size from 933 kg to 1.08 kt (30,000 to 35 million contained oz) of gold (Fig. 1). Newmont Gold Co. and its parent, Newmont Mining Corp., jointly constitute the largest mineral right holders in the district. They own or control more than 1000 km' (386 sq miles) in and around the Carlin Trend and own all or part of I6out of the 19 mines and prospects identified to date. Since the initiation of Newmont's exploration activities in the Carlin Trend in 1961, 2.24 kt (72 million oz) of cumulative gold resources have been identified. Cumulative production from all mines since the start-up of Newmont's Carlin mine in 1965 to the end of 1989 was about 202 t (6.5 million oz) (Jones, 1989). The incentive of sustained high gold prices and innovation in processing technology resulted in a significant acceleration of gold output over the last few years. Newmont Gold alone produced more than 43.5 t (1.4 million oz) in 1989. That is equal to 22% of the cumulative 1965 to 1988 output, and an almost 200% increase over its 1986 output. The same incentives produced even more spectacular exploration results. In each of the last five years, net additions to reserves and resources outpaced current production by substantial margins. These facts demonstrate the spectacular past prospectiveness of the Carlin Trend and the success of focused, multi-disciplinary exploration methods that made it possible to more than offset the recent accelerated depletion of gold resources. However, is this situation sustainable? How long can the mining companies along the Carlin Trend keep on finding resources faster than they deplete them? These are some of the questions that motivated this study. The authors have not quantified the future potential for gold exploration in the Carlin Trend nor established a deposit discovery path. But strong indications were discovered that the [ ] Carlin Trend remains a relatively immature exploration district and that the potential for significant new discoveries is high. Methodology and data The approach chosen to address the above questions was simple. The authors compiled deposit size data, measured in contained ounces of gold resources, for all known deposits along the Carlin Trend (Table 1). The resource information was obtained by adding cumulative historical production (adjusted for mining losses and metallurgical recovery) to 1989 year-end published resource inventories. In a mature exploration area, where most deposits have been discovered, this distribution would be expected to approximate lognormality and would plot along a straight line on a lognormal probability scale. This result was found in previous work by Allais (1957) and recently confirmed by Cox and Singer (USGS, 1986) in regard to various types of mineral deposits in several regions of the world. It was also found to hold true for oil and gas pool size distributions (Arps and Roberts, 1958; Kaufman, 1962; McCrossan, 1969). [ ] The data used were compiled by Newmont Exploration geologists. The purpose of the study is to make inferences about the underlying geologic processes in the district and the maturity of the exploration effort. Therefore, deposits were not classified according to ownership but according to geologic occurrence as known from current information. Newmont's Post and Barrick's Goldstrike and Betze deposits, for example, are shown as a single occurrence to reflect the actual geologic setting. The cumulative frequency distribution of deposit sizes was plotted on lognormal probability paper (Fig. 2). The abscissa shows the cumulative fraction of deposits at or below a certain deposit size and the ordinate shows the deposit size in thousands of ounces of contained gold resources.

Jan 1, 1992

"Foreword by Ren BryceSenior engineer, Freeport-McMoRan, Oro Valley, AZ, USAThe eighth annual Mineral & Metallurgical Processing Division (MPD) student poster contest was held in February at the national SME Annual Conference & Expo in Denver, CO. A total of 33 students presented their work at the poster competition. The competition started in January when the students submitted a one-page abstract for judging. In February, they prepared and presented a poster that explained their research. A panel of judges went to each student and listened as the students explained their research in five minutes.The students worked hard at their research and then set aside time to complete all the necessary steps to enter the competition. The MPD recognizes the effort required to enter and as a result, provides monetary rewards to all participants. The first-place winner receives $1,000; the second, $500; the third, $250; and the fourth through seventh, $200. All remaining students receive $100. All of the students also receive a complimentary ticket to the MPD Luncheon. As in past years, there are separate levels of competition for undergraduates and graduates.Graduate winnersKaustubh Shrimali won first place for researching “The influence of starch in amine adsorption by quartz in the reverse flotation of iron ore.” Venkata Atluri received second for studying “The effect of polysaccharides on bubble attachment at talc surfaces.” Jennifer Galvin placed third for investigating “A novel approach to alkali pug bake for the extraction of rare earth elements from monazite concentrate.” In the fourth through seventh places were Amy Richins, Sumedh Gostu, Brett Carlson and Chris Vass.Undergraduate winnersNestor Santa earned first place for exploring “Copper recovery from waste printed circuit boards.” Cassidy O’Connell received second place for conducting “Performance analysis of CMC in high talc copper sulfide flotation systems.” Catalina Vanegas won third place for examining “Cyanidation tests and pretreatments to recover gold from a refractory ore in Colombia.” Kylie Boyce, Michael Archambo and Pritam Sinha were in the fourth through sixth places.Everyone benefits from attending the student poster competition. The students are able to display their work to future employers, and industry professionals can talk to up-and-coming engineers. Next year, there will be another competition for more posters and networking. Come and enjoy the event and the posters."

Jan 1, 2017

By W. F. Hahnert, T. L. Ward, K. E. Stoll, P. A. S. Ralston

Fuzzy logic is well suited to the control of non-linear, time-varying, and/or any system for which it is difficult to obtain an exact mathematical model. To design a good fuzzy controller, a thorough understanding of the desired process behavior is required. Many papers have suggested that fuzzy controllers are robust. If they are, it is reasonable to explore whether or not they respond better than conventional controllers to errors in the feedback signal due to imperfect measurement or sensor degradation. It may be the case that simpler or less expensive sensors could be used in some processes. The research presented here relies on an integrated array of computer programs that model and simulate (1) the process to be controlled, (2) common feedback instrumentation, and (3) both fuzzy logic controllers (FLC's) and conventional controllers. Various types of feedback sensor imperfections are imposed upon the system, and the performance of the two controllers is compared. Each controller is tuned to its optimal performance in response to a setpoint change using perfect feedback measurement.

Jan 1, 1993

By N. Parra, M. N. Herrera, T. Vargas, C. González, B. Escobar

The bioleaching of mineral sulfide pulps in a bench-scale rotating-drum reactor was studied. The reactor contained horizontal paddles on the inner wall for lifting the mineral pulp. The performance of the reactor was tested by bioleaching a refractory pyritic gold concentrate in the presence of pure Thiobacillus ferrooxidans (ATCC 19859). The concentrate was efficiently bioleached at a 50% pulp density with the drum operating at from 0.5 to 1.0 rpm. Attempts at bioleaching the concentrate at the same pulp density used in a conventional stirred reactor showed severe inhibition of bacterial activity. The results demonstrated that the drum reactor allows the process to operate at much higher pulp densities than conventional stirred-tank reactors. This results in improved bacterial activity because of better attrition control of the mineral particles.

Jan 1, 1999

By S. Kelebek, G. W. Smith

Four samples of galena, each prepared differently, dry ground chalcopvrite and real gar have shown various quantities of liberated sulfur corresponding up to nine equivalent monolavers on their surfaces as determined by the solvent extraction/UV analysis of sulfur using ethanol. The pretreatment of a dry ground galena sample with acidified ammonium chloride or neutral, saturated sodium chloride solutions rendered the mineral readily floatable. The collectorless floatabilities of galena pretreated with the former solution and of drv ground chalcopvrite have been evaluated in reference to the natural floatability of sulfur using the critical surface tension of floatability approach with aqueous methanol solutions. It is shown that the amount of sulfur- on these samples can be correlated with their flotation rate, which was found to he first order within the critical surface tension range. The larger the concentration of surface sulfur, the faster the sulfide mineral floats, thus, the greater its flotation rate constant.

Jan 1, 1990

By R. Mira, S. Rosa, P. Souza

This paper shows the historical data, local geology and mineralization, the mining methods applied, current production,, the planning staff, the resources of hardware and software available and the computer methods used. The activities are described starting with the driving of ramps/research levels, their dimensional parameters and positioning, driving of production sublevels with the same parameters plus those concerned with the stability of the rock mass during the development and the mining itself. Then, the detailing of the ore body is described together with the stopes planning, the openings stability analysis, and the drilling patterns and explosive types are discussed.

Jan 1, 1998

By P. J. D. Lloyd

INTRODUCTION The mining of gold commenced on the Witwatersrand in 1884, and reached a mature state by about 1910, when nearly 20 million metric tons were mined. Mining has continued on the central Witwatersrand, the area of the first mining, at a similar rate until recent years. As a result, well over 1 billion tons of gold mining wastes have been dumped along a line over 80 km long, centred on Johannesburg and stretching from Randfontein in the west to Nigel in the east. (see Figure 1). The population of this conurbation, which straddles the line of mining, has grown to about 5 million. In 1922 uranium was first discovered as a constituent of the ore (and thus of the wastes). Investigations in the 1940's revealed that values near Johannesburg were low, but that, at the western and eastern ends of the central Witwatersrand line, and in the new areas centred on the towns of Carletonville, Klerksdorp and Welkom up to 200 km from Johannesburg, values were adequate to permit uranium recovery. Commercial production commenced in 1952, and by 1979 over 100 000 tons of uranium oxide had been produced. There thus arises on the central Witwatersrand nearly 1,3 billion tons of slimes containing an average of about 0,005% uranium and about 1,8% FeS2. The material grading is about 70%-75µm, and the primary mineral is silica (over 85%). The material is contained in dams constructed hydraulically. A typical dam has an area of about 100 hectares and there are over 6000 ha under slimes dams in the area of concern. The treatment of these dams for uranium recovery commenced in 1977. The situation has thus arisen whereby potential uranium ores have unwittingly been dumped in significant quantities within a major metropolis over a period of nearly a century. The purpose of this note is to review the experience arising from the potential hazard over a considerable time.

Jan 1, 1980

By B. V. Knelson

Although the concept of centrifugal recovery has been around in one form or another for many years, it has not been until recent times that it has became clearly understood am properly put into perspective. Simplistic in theory, it is when one begins to examine the variables involved that the true complexities of this matter begin to emerge. In creating an enhanced g force concentrator, there are many aspects both positive and negative that are encountered.

Jan 1, 1989

By Ammar Henchiri

INTRODUCTION A current need of the phosphate industry is the development of process technology for treatment of high carbonate content ore. About 75 to 80 percent of the world phosphate reserves (1) are associated with carbonate gangue, which is not amenable to treatment by the existing techniques. Numerous laboratory and pilot plant studies have been devoted to carbonate-phosphate separation. Different techniques were involved (washing, sieving, attrition, flotation, calcination, and leaching). For technical and economical considerations, flotation appears to be an efficient process for this problem. Separation of carbonates (calcite and dolomite) from igneous phosphate is applied in industrial plants in Brazil, Finland, and South Africa. But, for sedimentary phosphate, no industrial realization exists at this time, with the possible exception of some plants in China. The main problems with sedimentary phosphates are: • high substitution by CO3-2 and F- for P04 -3 in the apatite lattice. • larger surface area of the phosphate particles because their surface is porous and irregular. • tendency of sedimentary minerals to produce more slimes in conditioning than the crystalline minerals, hence, larger dosages of collectors are required for sedimentary minerals. Investigations reported in the literature indicate that most processes developed for sedimentary carbonate-phosphate separation involved inverse flotation of carbonates using fatty acids as collectors (2), (3) associated with phosphoric acid (4), monosodium phosphate (5), F- (6), or SiF6 -2 (7) as phosphate depressants. Later investigations, conducted by A. Henchiri and G. Baudet (8), (9), (10), (11), (12), (13) in the BRGM technology department, of phosphoric esters revealed that these reagents are selective collectors for carbonate-phosphate separation. The above investigators developed a process which used phosphoric esters to collect carbonate minerals and sulfuric acid or sodium fluorosilicate as depressants for phosphate. Extensive comparative tests for different processes on the same phosphate sample show that the process developed at BRGM is one of the most promising methods proposed for sedimentary phosphate-carbonate separation. It should be pointed out, however, that phosphoric esters make abundant froths. The handling of these excessive froths will be a problem in an industrial-scale plant. This paper outlines test work performed in order to optimize and to improve the efficiency of the flotation process on a sedimentary phosphate ore selected from the Sra Ouertane orebody. The Sra Ouertane flotation process, illustrated on Figure 1, first removes carbonate minerals by inverse flotation and then recovers phosphate as a froth product, leaving quartz as a cell product.

Jan 1, 1993

By Z. D. Yu, J. E. Gebhardt, D. I. Phillips, S. X. Jiang, R. Z. Zhang

The potential application of flotation for the treatment and processing of a fine-particle coal stream was investigated for the Chaili Coal Washery of the Zaozhuang Coal Mine Administration, Shandong Province, China. A goal of the test work was to evaluate the application of the Microcel™ flotation column and compare its performance to conventional flotation. Small-scale flotation tests were performed in the laboratory and on-site with fine coal slurry samples from the plant thickener underflow, i.e., the washing plant's undersized reject material. Flotation tests were conducted on-site with a 7.6-cm diameter Microcel™ column and with small-scale conventional cells. Column flotation tests were performed to determine a recovery-grade relationship for the fine coal feed, which contained about 24% ash, and to obtain sufficient data to enable scale-up to a large diameter unit. A primary objective as to establish the maximum throughput capacity of a Microcel™ flotation column while operating to produce a clean coal product with ash content of 7.5-7.8%. Results of the test work indicate that this product quality could be achieved at feed rates of 15-19 tph to a 3-m diameter Microcel™ flotation column. Lower ash products, i.e., 5-7% ash, were obtained but at lower column throughput capacities. Comparative flotation tests, conducted with conventional flotation cells, indicated that a product ash of only 8% or greater was achieved in a single-stage test.

Jan 1, 1995

By L. B. Phelps

The nation is confronted with thousands of acres of abandoned coal lands. In Pennsylvania alone, these lands have been estimated to cost $15 billion to reclaim (Pennsylvania Bulletin, 1985). One way to deal with this problem is to provide a mechanism by which the active mining industry can undertake to remine certain of these abandoned mine areas. Remining is classified as reaffecting previously mined areas in order td extract the remaining coal reserves while improving the environmental quality of these lands through land restogation and abatement of pollutional discharges. However, there currently is a disincentive for operators to plan to extract reserves on these abandoned mine lands which have preexisting pollutional discharges. Therefore, they presently avoid them. This is because of the "potentially unlimited liability of post- mining treatment for any discharges that would remain after mining? (Pennsylvania Bulletin, 1985).

Jan 1, 1987

By E. Suoninen, K. Laajalehto, L. Kartio

Efficient separation of pyrite from other sulfides is of great importance for the metallurgical performance of flotation processes. In this work Fourier Transform Infrared Spectroscopy (FTIR) and X-ray Photoelectron Spectroscopy (XPS) were used to study the surface chemistry of pyrite in depression, activation and xanthate adsorption under conditions of controlled potential. Modifications of pyrite surfaces after treatment with depressants (lime, sulfur dioxide, sodium cyanide) and after activation with metal ions (Cu, Pb) were studied. The principal adsorption product identified on pyrite was dixanthogen whose formation started at about +0.15 V and +0.25 V (vs, SHE) in ethyl and amyl xanthate solutions, respectively. Copper xanthate was formed on copper(II) activated pyrite. Activation mechanism of pyrite by copper(II) salts is likely to be electrochemical where copper occurs as copper(I) on the surface of pyrite. Effective depression is achieved by sulfur dioxide and sodium cyanide. Depression at high pH is due to formation of iron(III) hydroxides. Calcium ions do not affect the electrochemistry but adsorb on pyrite and reduce the surface sites for dixanthogen adsorption.

Jan 1, 1995

By Dennis Ofiara, Paul Zick

Two 8.2 m diameter flood control tunnels are being constructed in the Taylor Shale Formation beneath San Antonio, Texas. These tunnels use precast concrete segments as final lining. This paper describes the features of the segments chosen for this project. The tunnels are being excavated with a hybrid Tunnel Boring Machine. This paper also describes the features of the TBM which make the machine compatible to the type of segments chosen.

Jan 1, 1989

By Takashi Sakaguchi, Takehiko Tsuruta, Akira Nakajima

Some attempts were made to recover gold from aqueous systems using immobilized persimmon tannin. This adsorbent adsorbed gold from solutions containing hydrogen tetrachloroaurate (III) with high efficiency, whereas there was minimal adsorption from solutions containing gold (I) sodium thiomalate and sodium dicyanoaurate (I). The adsorption of gold was rapid, and was affected by the pH of the solution, temperature, external gold concentration and amount of adsorbent. Adsorbed gold was easily desorbed with 1 M thiourea solution, indicating that immobilized persimmon tannin can be repeatedly re-used for the recovery of gold using adsorption-desorption cycles.

Jan 1, 1995

By Eung Ha Cho, Olufemi Olajide

A laboratory study of dense-medium cycloning with Redstone seam coal has been conducted. The investigation was performed using a 203 mm (8 inch) Heyl-Patterson dense-medium cyclone. Two underflow orifices of 31.8 mm (1.25 inch) and 44.5 mm (1.75 inch) in diam were used. The effects of feed pressure and particle size were determined on the cycloning performance. In addition, the cyclone performances with coal-rich and refuse-rich samples were also investigated. The results of the experiments were analyzed using basic hydrodynamic principles. The increase in the feed pressure increases the specific gravity of separation (defined as the specific gravity of a particle which has an equal chance of reporting to either the overflow or underfow) for both underfow orifices. The specific gravity of separation increases as the particle size decreases with the smaller underflow orifice. However, an opposite relationship has been observed with the larger under flow orifice. The separation becomes sharper with the smaller under flow orifice than with the larger underflow orifice especially for fine material. Also, the separation becomes sharper with an increased feed pressure for a coal-rich sample. However, this effect is less significant for a refuse-rich sample. This means that a high feed pressure is required for a coal-rich sample to maintain a good separation. It also means that applying a high feed pressure provides a good separation regardless of the type of feed coal in terms of ash content in the solid.

Jan 1, 1990

By Alexandr A. Abramov, Svetlana M. Angelova, Oleg P. Maljuk

In this paper the effect of the ratio of grinding and grindable size classes (ore lumps, particles and grains) and the guantity of intermediate size classes in the feed of the primary wet autogenous grinding "Cascade" mill is studied and the features of the influence of these factors on the mill capacity are shown. On the basis of the experimental data from Lebedinsky mineral processing plant, where magnetite guartzites are processed, a coefficient of the "total ratio of size classes" in the feed of 9x3 m "Cascade" mill was estimated. It helps to get an evaluation of the effect of ore size classes ratio in mill media on the mill capacity and to work out some ways for its optimization in closed or semiclosed technological circuites of primary autogenous grinding. The operation of the specific energy effects for some intermediate size classes, removed from the mill as circulating load, is proposed to improve its efficiency before regrinding in "Cascade" mill and decrease negative influence of such classes accumulation in the mill media.

Jan 1, 1995

By John P. Weir

The coal industry is ever changing and complex. This article discusses some problems that must be faced and their projected economic effects. Coal's dominant market is for electric power generation. While growth rates in the 1980s will be nominal compared to historic growth rates, such growth will be substantial in absolute terms because it comes from a very large base. Higher growth is expected in the 1990s. History When looking toward the future, an examination of the past is helpful. Figures 1 and 2 present production and consumption statistics (expressed in short tons) on bituminous coal and lignite in the US from 1945 onward. After an all-time peak of 572 Mt (631 million st) in 1947, total production dropped to 356 Mt (392 million st) in 1954. Sustained growth in total production did not begin until the 1960s. Coal production from under-ground mines declined from more than 400 Mt (440 million st) in 1947 to about 295 Mt (325 million st) during 1965 to 1970. A further decline ensued until a strike in 1978 reduced production to only 218 Mt (240 million st). Underground mines produced 283 Mt (312 million st) in 1983. Coal production from surface mines has increased steadily since the early 1960s. While some of this increase has come from mines east of the Mississippi River, most has resulted from surface mines developed in the West. Production from surface mines west of the Mississippi River was 224 Mt (247 million st) in 1982. Total coal production from under-ground and surface mines, other than Western surface mines, continues to approximate levels of 40 years ago. In spite of several wage increases, increasing productivity in surface and underground mines enabled the coal industry to hold the average value f.o.b. mine to about $5.50/t ($5 per st) from the end of World War II to 1969. Under-ground mine productivity peaked at more than 13.5 t (15 st) per man-day in 1969. The Coal Mine Health and Safety Act of 1969 was a major factor in underground productivity declining to 7.6 t (8.4 st) per man-day in 1978. However, pro-

Jan 7, 1984