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By S. C. Sun, A. M. Gaudin

This paper presents the results of a. study of the relationship of the electrokinetic or zeta potential to flotation phenomena. The relationship is established on the experimental evidence that mineral behavior in flotation and in cataphoresis is generally governed by the chemical properties and amount of the reagents adsorbed at mineral surfaces. Cataphoresisl-5 or electrophoresis is the name given to the movement of a charged particle in a liquid medium under the influence of an applied electric field. Mineral behavior in cataphoresis, as indicated by the magnitude and sign of its ionic charge, can be represented by the term zeta potential. In this paper, the zeta potential represents the difference in potential between the immovable liquid layer attached to the surface of a solid particle and the movable liquid layer in the body of the liquid, and is the result of the tendency of ions on the surface of the particle and within the liquid boundary layer to establish a condition of equilibrium. A high positive zeta potential means that the mineral surface is dominated by cations, a high negative zeta potential indicates domination by anions, and a low zeta potential indicates that the surface is influenced almost equally by cations and anions. In terms of observed quantities the zeta potential is given by the relation _ 4µ(300)² Z= kx in which z is the zeta potential expressed in in volts, µ the viscosity of the solution in poises, u the velocity of particles in centimeters per second, x the impressed potential gradient in volts per centimeter, and k the dielectric constant of the solution in c.g.s. units. A convenient source for the derivation of the equation is the book by Thomas. The term zeta coefficient is an expression coined to denote the difference between the magnitude of the zeta potential of a mineral suspended in solution A, and that of the same mineral suspended in solution B, which consists of solution A and an additional chemical agent, but is at the same pH. For example, the difference between the zeta potential of galena particles in distilled water and that in any chemical solution at the same pH is a zeta coefficient. Or, when the zeta potential of a mineral particle is measured first in distilled water, then in a mixture of water and copper sulphate, and finally in a mixture of water, copper sulphate and xan-thate, we can calculate the zeta coefficient

Jan 1, 1947

By A. M. Gaudin, S. C. Sun

This paper presents the results of a. study of the relationship of the electrokinetic or zeta potential to flotation phenomena. The relationship is established on the experimental evidence that mineral behavior in flotation and in cataphoresis is generally governed by the chemical properties and amount of the reagents adsorbed at mineral surfaces. Cataphoresisl-5 or electrophoresis is the name given to the movement of a charged particle in a liquid medium under the influence of an applied electric field. Mineral behavior in cataphoresis, as indicated by the magnitude and sign of its ionic charge, can be represented by the term zeta potential. In this paper, the zeta potential represents the difference in potential between the immovable liquid layer attached to the surface of a solid particle and the movable liquid layer in the body of the liquid, and is the result of the tendency of ions on the surface of the particle and within the liquid boundary layer to establish a condition of equilibrium. A high positive zeta potential means that the mineral surface is dominated by cations, a high negative zeta potential indicates domination by anions, and a low zeta potential indicates that the surface is influenced almost equally by cations and anions. In terms of observed quantities the zeta potential is given by the relation _ 4µ(300)² Z= kx in which z is the zeta potential expressed in in volts, µ the viscosity of the solution in poises, u the velocity of particles in centimeters per second, x the impressed potential gradient in volts per centimeter, and k the dielectric constant of the solution in c.g.s. units. A convenient source for the derivation of the equation is the book by Thomas. The term zeta coefficient is an expression coined to denote the difference between the magnitude of the zeta potential of a mineral suspended in solution A, and that of the same mineral suspended in solution B, which consists of solution A and an additional chemical agent, but is at the same pH. For example, the difference between the zeta potential of galena particles in distilled water and that in any chemical solution at the same pH is a zeta coefficient. Or, when the zeta potential of a mineral particle is measured first in distilled water, then in a mixture of water and copper sulphate, and finally in a mixture of water, copper sulphate and xan-thate, we can calculate the zeta coefficient

Jan 1, 1947

By Rajesh Minda

160 numbers of double bulb underreamed piles were constructed based on the safe loads specified in IS code at Tripura. The pile load tests yielded a much lower capacity. A detailed investigation work was carried out to find out the reasons for anomaly. The work included soil exploration followed by extraction of two piles bearing numbers 72 and 151. It was seen that the bulbs were not formed as per specification. Some comments on IS Code have been made at the end.

Jan 1, 1996

The general nature of the special constructional features that are required to provide safety from explosion hazards on machines approved as permissible by the Bureau of Mines is fairly well known to mining men who are at all interested in safety. These requirements are outlined in the schedules, under the provisions of which the investigations of such machines are conducted. Also, the method of conducting inspections and. tests of such machines has been described in Bureau of Mines Bulletin 305, Inspection and Testing of Mine Type Electrical Equipment for Permissibility, published in 1929. A more detailed knowledge of the special safety features of a particular machine is often useful and sometimes essential for inspectors, safety engineers, maintenance men, and engineers in selecting machines for a particular application. Three bulletins that give information of this nature have been published by the bureau, viz., Bulletin 313, Permissible Storage-Battery Locomotives and Power Trucks. Bulletin 343, Permissible Coal-Handling Equipment. Bulletin 382, Permissible Coal-Cutting Equipment.

Jan 1, 1937

By P Duplancic and C R Lilley, R Hudson, D Beck

For deep block and panel caving mines, undercut or extraction level instability can have as much impact on overall recovery as underbreak or the premature ingress of waste. If the area of the mine affected by excessive convergence is small, recovery is usually possible. However, there are several examples of entire blocks that were abandoned prematurely due to footprint collapse. Clearly, such global instability has a major impact on recovery and must be avoided – this is one of the essential tasks for the design team.The design problem is complex. The planning engineer must ensure that the footprint has sufficient capacity to withstand the maturing loads in the cave, accounting for the effects of stress path, geological structures and the rock mass conditions. These are often uncertain and only sometimes quantified.The demand on the footprint will also be influenced by decisions and actions taken during production, especially related to draw, which can worsen a developing problem. Perhaps most importantly, the design decisions that favour stability and lead to lower deformation also negatively impact on idealised cave flow. Many engineers aim to minimise the size of extraction level pillars, to improve theoretical future resource recovery but in fact this can reduce recovery if widespread deformation ensues due to the undersized pillars.In 2014, PT Freeport Indonesia undertook detailed numerical analysis to quantify the effects of variability in rock mass conditions and geological structure on overall footprint performance. The analysis accounted for a range of stress paths and maturing cave loading scenarios sourced from a calibrated whole-of-cave deformation-flow model. The intent was to replicate some of the diversity and distribution of pillar deformation outcomes that may be seen in practice at Freeport.The discontinuum finite element analysis was able to replicate damage localisation around persistent structures, a realistic distribution of drive deformation and show how small areas of instability can evolve into major problems if poorly managed. The limits of pillar size and ground support as solutions to a number of different instability problems was also investigated, showing that pillar size alone was not a sufficient remedy for all adverse circumstances.In this paper, we describe the workflow for numerical analysis of footprint deformation. We discuss the compromises and considerations for balancing the sometimes competing needs for efficient extraction levels pillars, ground support and prudent draw strategy.CITATION:Beck, D, Hudson, R, Wilson, A D, Duplancic, P and Lilley, C R, 2016. An investigation of factors affecting the long-term reliability of extraction level pillars in deep block and panel caves, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 295–304 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

A numerical method for discriminating between replacement and sedimentary processes is applied to a folded stratiform sulphide occurrence. It is shown that numerical treatment of chemical analyses from a single diamond drill hole allowed a rapid assignment of the ore to the syngenetic sedimentary class and thereby enabled a revision of the evaluation program concerned with consequent reduction in costs.

Jan 1, 1973

By L. M. Kaas

Production planning techniques, particularly a production scheduling model are proposed for use in taconite mining in Minnesota. Applications of the scheduling model, on a two-year mine plan, which evaluates the equipment involved, are detailed. Data and tables are included. INTRODUCTION Fifty years ago Walter Swart, a mining engineer investigating the Minnesota taconites for the Jackling interests, made this report: "There is no doubt in my mind that this taconite mining is a game worth playing and that it can be played to a successful finish. It seems now only a question of men, money, methods and details."' The prosperous industry envisioned by Swart and other taconite pioneers is today a reality. Its growth has been spurred by technological advances in mining and beneficiation, depletion of natural ore reserves, and improvement of the economic climate. However, the massive problems that were solved to make this development possible are only a foreshadowing of the questions of "men, money, methods and details" that will be encountered in the future. The increasing demands for higher quality taconite pellets have already forced producers to examine better methods of quality control for their operations. THE NEED FOR PRODUCTION SCHEDULING The maintenance of fairly constant crude ore characteristics is of major importance to mill quality control. Through careful mixing of ore from storage bins, the concentrator staff can partially reduce the deviations of the feed analysis from the analysis desired. However, the limited effectiveness of bin mixing systems has caused some mining companies to utilize large capacity blending piles to reduce grade variations. Although the analysis of material being reclaimed from such a pile is relatively uniform, raw materials still must be scheduled from the mine into the pile in such a manner that the average analysis of all material is close to the feed analysis prescribed by the mill. PRODUCTION SCHEDULING MODEL Minnesota taconite mining companies have used three analyses to estimate how a certain ore will react to concentration. These factors are the crude magnetic iron content of the ore, the iron content of the plant concentrate produced from the ore at a standard grind, and the grindability index of the ore. The actual laboratory tests used to determine these factors have been described by Apuli.2 These data are obtained from analysis of diamond drill cores, and are the established criteria used by mine planning engineers to formulate taconite production schedules. Variations in the grade of mine output should be controlled by any production scheduling system. However, to be truly practical, production scheduling methods must also consider many of the less tangible variables that affect mine operation. These include pit layout, available equipment, and ownership of ore leases. The number and scope of the variables inherent in this type of mine planning suggested that a computer simulation could be used to solve the problem. The preparation of information for use with a computer simulation program was the initial phase of this project. Meaningful data was developed from logs of diamond drill cores. Analyses were related to their respective areas of influence by means of mine maps, Fig. 1. The three assay variables and tonnages for each area were punched on cards. In addition to these values, a location code number was punched in the first eight columns of the cards. This code is explained as follows: Column 1 Bench number Column 2 Quadrant location as determined by mine reference lines Columns 3-5 East-west coordinate Columns 6-8 North-south coordinate The heart of any simulation techniques is a well engineered mathematical model. To develop a model for taconite production scheduling, the relationship between the three assay variables had to be established. Because changes in each analysis factor affect the mill circuits in different manners, a weighting system was used to determine the relative im-

Jan 1, 1967

By Jeffrey A. Schaefer

The Olmsted Locks and Dam project is located on the Ohio River approximately 16.5 miles upstream of its confluence with the Mississippi River. It consists of twin 1200 ft. locks with floating and fixed approach walls, five 110 ft. wide tainter gate bays, a 1400 ft. wide navigable pass boat operated wicket gate section, and a 400-ft. section of fixed weir (Fig. 1). The project will have a total cost over 1.0 billion dollars, and as such is one of the largest civil works project underway in the Corps of Engineers.

Jan 1, 2001

These proceedings consist of papers presented at two Bureau of Mines Technology Transfer Seminars for the purpose of disseminating recent advances in mining technology related to coal mine fire and explosion prevention. Primary, although not exclusive, emphasis is placed on underground coal mining. The papers address research accomplishments ranging from methods for detecting mine fires to remote sealing systems for extinguishing coal mine fires. Each paper represents the current state-of-the-art in the respective research categories. Presentations include research occurring from both Bureau in-house and contract efforts.

Jan 1, 1978

By Narasimhan K. S, Sastri S. R

A typical high silica low grade magnesite from the Salem area of Southern India, was used in studies to develop a suitable flotation system involving commercially inexpensive sodium oleate as the collector together with other reagents which can give a product of the required grade with an acceptable recovery. It was found that the presence of metallic ions such as Cu+, Fe+++ or Ca++ introduced by the addition of their salts, or by their formation with the addition of hydrochloric acid to react with magnesium and calcium carbonate in the me, can act as effective depressants for silica. The addition of 'hydrochloric acid' was found to give best results.

Jan 1, 1980

By Kenneth O. Bayard, John D. Sullivan

"This repot continues the series of papers being prepared at the Southwest Experiment Station of the Bureau of Mines in cooperation with the department of mining and metallurgy, University of Arizona, on factors involved in the leaching of ores. A list of previous publications is given in the bibliography at the end of this paper.Four factors are essential to the leaching of any copper-bearing ore: (1) A solution that will attack the copper minerals must get into the body of the ore particles; (2) the copper minerals must be dissolved by the solution; (3) the solution containing the dissolved copper must find its way out of the voids of the rocks into the main solution; and (4) the copper in solution must be recovered by some means of precipitation.This paper considers the third step -- the extraction of the water-soluble copper. From leached ores. The usual commercial practice in leaching copper ores consists of: (1) Crushing the ore to pass through a screen with approximately one-quarter or one-third inch openings; (2) charging the crushed ore into a fat; (3) adding, usually by upward percolation, a solution that will attack the copper minerals; and (4) washing the copper-laden solutions out of the ore with solutions weaker in copper than the pregnant solution, and finally with water. A countercurrent system of leaching and washing is usually employed. In some plants the solutions are circulated during leaching and washing, and in others they are not. In some plants the wash solutions are added by upward percolation, while at Chuquicamata the so-called ""piston-displacement"" method of washing is employed.The following condensed resume of the leaching cycle of the plant of the Inspiration Consolidated Copper Co. is given to present to the reader a picture of the steps involved in a typical leaching cycle. This plant was recently described in detail.4 The ore, crushed to pass openings 11/2 inches in diameter, passes over a screen with ¼-inch openings. The undersize from the screen goes to a Dorr howl classifier that removes the primary slime from the circuit. The overflow from the classifier is treated in a slime-leaching plant. The oversize of the ¼-inch screen passes through rolls, where it is crushed to pass a 3/8-inchscreen. The undersize from the 3/8-inch screen join with the coarse material from the classifier, and the combined products become the feed to the leaching vats. There are 13 concrete lead-lined leaching vats. The complete leaching cycle is 13 days. Eight tanks are under acid-iron content at all times; three are being washed, one tank is being filled, and the one is being excavated. The electrolyzed solution from the tank house, to which amount of new acid has been added, is put on the oldest tank of ore under need This solution travels from tank to tank and finally emerges from the newest ore"

Apr 1, 1931

By P Staples, G Lane, S Morrell

The design of semi-autogenous grinding (SAG) mill based comminution circuits for the treatment of competent ores, similar to those at Geita and Boddington, involves the same process of sample selection, test work, data analysis and data modelling/interpretation as that used for 'average' competency ores. However, over the past couple of years a number of issues have arisen that are common to the design of circuits for the comminution of 'highly' competent ores. The impact of these issues on project viability is generally more pronounced when treating hard ores than average to soft ores due to the greater impact on capital and operating costs. The issues encountered have included: standard test procedures have been modified; test equipment has worn or been damaged; different procedures yield different data and varying interpretation; and modelling and empirical calculations have been based on poor benchmarks, or used incorrectly in the case of JKSimMet, yielding misleading outcomes. The purpose of this paper is to present, discuss and clarify some of the issues associated with conducting test work and designing comminution circuits for the treatment of 'highly' competent ores in order to reduce thelevel of conflict arising from interpretation and application of test work data. Specifically, the issues associated with the bond crushing (impact) and rod work indices measurement, the various SAG mill specific energy tests, and the interpretation of the resulting data will be discussed in the context of case studies.

Jan 1, 2009

By S. L. Windes

This report presents physical and petrographic properties of approximately 60 types of rock from operating mines or mineral-investigation projects and supplements Repot of Investigations 4459, Physical Properties of Mine Rock (hereinafter referred to as Part I). The physical properties evaluated include the following: Apparent specific gravity Apparent porosity Compressive strength Flexural strength (modulus of rupture) Impact touchiness Abrasion hardness Scleroscope hardness Modulus of elasticity (Yourg?s modulus) Modulus of rigidity Specific damping capacity Longitudinal bar velocity

Jan 1, 1950

Provide long-term measurement of mine sidewall closure in backfilled stapes subject to flooding. Approach A water resistant extensometer is made by enclosing a rod-actuated transducer in a telescoping tube which is enclosed within an oil-filled outer housing. How It Works To assemble a water resistant extensometer, a cylinder -shaped transducer and two support rings are pushed into the small-diameter section of the telescoping tube where the transducer is held in place by friction. The outside of the small-diameter section of the telescoping tube is then greased and slid into the slightly larger-diameter section. See drawing. The cylinder-shaped transducer contains a potentiometer whose actuating arm is attached to a stainless steel extension rod that passes through one end of the telescoping tube. A similar rod is attached to the opposite end of the transducer body, and extends through the other end of the telescoping tube. A coil spring inside the large diameter end of the telescoping tube extends the tube

Jan 1, 1978

By A Tawadrous, P D. Katsabanis, S Kim

A series of small scale tests, have been performed to establish the effect of powder factor and initiation timing on the grindability of rocks. The blasts were performed in high quality granite blocks, which were cut from stone prepared by dimensional stone quarry operations. Three different types of granite, Barre, Stanstead and Laurentian, were tested. The charge consisted of detonating cord placed in five holes with water as a coupling medium, or concentrated charges of equal mass of high exolosive, fully stemmed or air decked. Powder factor was varied from 0.39 to 1.17 kg/m3. Charges were detonated simultaneously or with a small delay and in some experiments, detonation waves collided in the middle of each borehole. All fragments were collected and screened, and drop weight impact tests were performed on selected size fractions, using different levels of impact energy, to calculate the resistance of the rock to impact breakage. The work index of some of the blasted samples was measured using standard Bond mill grindability tests. The experiments generally showed increased softening of the rock with powder factor, manifested by the increasing initial slope of the relationship between t10 and impact energy. The rate of change of the initial slope decreased with increasing powder factor. Changes in the work index were small and consistent in trend with the results of the single fragment impact characterization tests. Charge configuration, timing as well as in-hole collisions of detonation waves affected the initial fragmentation but had little effect on the impact breakage of the samples.

Jan 1, 2009

By M. Arlandi, J. C. Monge

In this paper, the finite element method is used to obtain deformations of a rock tunnel. More than 3,000 numerical simulations and calculations were carried out. The analysis of this results has allowed to achieve mathematical formulations and graphs, to obtain in advance support monitored displacements in tunnelling construction, so they can be predicted before excavation and supporting woks. The application of the presented formulations and charts will aid engineers to control tunnel monitoring, reducing decision time when problems occur. The analysis is limited to cases involving elastic behaviour of the terrain for tunnels of average to good quality in rock, and to excavation sections of average or smaller dimensions for road tunnels. The results show that, in all cases, mathematical expressions to obtain induced displacements around tunnel excavations, have similar pattern than Kirsch (1898) expressions for circular excavations. As corollary of the achieved results, a method to estimate in situ natural geostress ratio K0=sh/sv , by means of convergence measurements, is presented. In diesem Beitrag wurden die Verschiebungen eines Tunnels durch die Methode de Finiten Elemente berechnet. Mehr als 3000 numerische Simulationen und Berechnungen wurden ausgeführt. Die Analyse dieser Resultate hat es ermöglicht, mathematische Formeln und Grafiken zu entwickeln, die zur Voraussage der Verschiebungen in Tunnel dienen, auch unter Betrachtung der verschiedenartigen Abstützungen. So können die Verschiebungen vor den Erdarbeiten und dem Bau der Abstützung abgeschätzt werden. Die abgeleiteten Formeln und Grafiken werden vor allem einen grossen Dienst für Ingenieure bei der Mithilfe zu zeitsparenden Entscheidungen im Tunnelbau und bei der Tunnelkontrolle leisten. Die analysierten Fälle sind auf eine elastische Modellierung des Baugrunds beschränkt, sowie auf eine mittlere bis gute Qualität des Felsens und auf die Aushebung von Abtragungsabschnitten mittlerer oder geringer Grösse für den Strassenverkehr. Die Resultate zeigen, dass in allen Fällen die hergeleiteten Formeln zur Berechnung der durch den Tunnelbau verursachten Verschiebungen um den Tunnel die gleiche mathematische Struktur haben wie diejenigen von Kirsch (1898) für kreisförmigen Bodenaushub. Als Korollarium der ausgeführten Berechnungen wird eine Methode zur Abschätzung der vorherrschenden in situ Geospannungs-Verhälniss K0=sh/sv angegeben, die von den Messungen der Verschiebungen im Tunnel ausgeht. Dans cet article, la méthode des éléments finis est utilisée pour obtenir les déformations d’un tunnel creusé dans la roche. Plus de 3000 simulations numériques et calculs ont été effectuées. L’analyse de ces résultats a permis de définir des modèles mathématiques et des abaques afin de prédire, avant l’excavation et les travaux de soutènement, les déplacements auscultés pendant la construction du tunnel. L’application de ces modèles et abaques permettra donc aux ingénieurs de contrôler les déplacements durant l’auscultation du tunnel et de réduire le temps de décision lorsque des problèmes surviennent. L’analyse est limitée au cas du comportement élastique du terrain pour des tunnels creusés dans de la roche de moyenne a bonne qualité, et à de sections d’excavation de petites à moyennes dimensions pour des tunnels routiers. Les résultats montrent que, dans tous les cas, les expressions mathématiques permettant d’obtenir les déplacements induits autour des excavations de tunnel, présentent des similitudes avec celles de Kirsch (1898) pour des excavations circulaires. Comme corollaire aux résultats obtenus, une méthode pour estimer le rapport des

Jan 1, 2003

Up to 19 chemical parameters were studied in flotation pulps in six Australian lead-zinc concentrators with the aim of obtaining a better understanding of the nature of, and changes in, the chemical environment. This paperdiscusses pH, redox potential, and oxygen concentrations, which were all measured continuously.Two plants operated at natural pH in slightly acid circuit and one at natural pH in alkaline circuit, and pH was stable at all these plants. In two plants varying amounts of cement fill were present in mill feed and pHwas variable. At the other plant pH was controlled with soda ash and lime.Redox potentials became more positive as the pulp passed through the plant. The minimum change was 20 mV and the maximum about 250 mV. Potentials tended to be variable at any point early in the circuit, but were more stable later on. In some plants reagent additions had no noticeable effect on redox potentials. However in the two plants operating in acid circuit the redox potential of .zinc rougher feed was markedly affected by changes in copper sulphate addition rate. Sodium sulphite additions had an effect at other plants.Oxygen concentrations in grinding circuits tended to be low (30 per cent of saturation) and very variable. Flotation was usually conducted in pulps at least 90 per cent saturated with oxygen. Addition of sodium sulphite had a marked effect on oxygen content in one plant, but less effect in two others.Statistically significant relationships were found in some plants between pH and redox potential, redox potential and oxygen saturation, and oxygen saturation and pH.Improved control of copper sulphate and sodium sulphite additions could result from applications of this work.

Jan 1, 1970

By J. Betzhold

Knowing more about an orebody makes it easier to exploit it profitably. Many orebodies are made up of different rock types, alteration zones or mineralogical ensembles: oxides versus sulphides or facies with different metallurgical properties for example. Being able to model the spatial layout of these different facies is vital to good mine planning and management. Kriging cannot be used to generate numerical models of orebodies with complex geometric patterns because it gives too smooth an image. Over-smoothing has the meaning of changing the natural selectivity. Conditional simulations are required to do this kind of model respecting the natural variability: honouring the natural selectivity, dilution and ore loss. Standard geostatistical simulations cannot duplicate many of the complex spatial relationships seen in the earth sciences, such as differently orientated facies, sets of conjugate veins or ore types where certain facies cannot touch. After a presentation of the case study, it will be shown why traditional methods for simulating indicators like the sequential indicator simulations, or the truncated gaussian method, cannot be used to simulate it. Plurigaussian simulations are then introduced and shown to be more powerful than the previous methods in their range of application. This is confirmed by the case study treated; a Chilean copper deposit, for which the plurigaussian simulations are seen to respect the different orientations of the individual facies as well as honouring the facies seen in drill holes, and respecting the proportion of each facies even if this varies in space (e.g. increases or decreases with depth). Key words: geostatistics, indicators, conditional simulation, plurigaussian method, copper variability.

Jan 1, 2000

By H. J. Sutherland

This paper reviews the results of the subsidence research prop ram at Sandia National Laboratories. The manuscript highlights the following: the application of empirical methods (profile functions) to the subsidence above longwall panels in the U.S.; the use of the "rubble model" to describe the behavior of broken strata as it distends when it falls to the Mine floor (or top of the rubble pile) and then is subsequently compacted as it is loaded by overlying elements of strata; and, the application of physical modeling techniques (centrifuge simulations) and numerical techniques to study the failure mechanisms in highly structured stratigraphy. The capabilities of the latter two are illustrated ')y comparing their predictions to the results of a field case that has complicated stratigraphy.

Jan 1, 1982

By Gary Brendel

India produces about 70 Mt/a (77million stpy) of coal ash (fly ash and bottom ash) from burning about 200 Mt/a (220 million stpy) of coal for electric power generation. The average ash content is 35%. In contrast, the United States produces about 70 Mt/a (77 million stpy) of coal ash from burning 800 Mt/a (880million stpy) of coal for power generation. Its average ash content is less than 10%. In 1998, GAI Consultants Inc.(GAI) embarked on a federally funded, multiyear project to assist the government of India in solving its growing coal-ash management problem.

Jan 1, 1999