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By Lawrence M. Brennan

Early in 1971 the Peabody Design and Preparation group began engineering on a preparation plant which was to process eight mil- lion tons per year of raw coal. For the first five or six years this product was to be furnished 50% from surface mining operations and roughly 50% from underground operations. By early 1980, 1008 of the production was to come from underground mines. Because of the large annual commitment from Randolph Preparation Plant, it was essential that the plant operate with minimum down-time. As usual, capital costs were to be kept at a minimum. Before we discuss the details of design, it would be best to discuss the overall picture. Underground production to be pro- cessed through Randolph Preparation Plant was to be delivered by means of one mile overland conveyor from Baldwin #1 Underground Mine to the plant. Surface Mine production was to be delivered by means of 100-ton haulage units dumping in a 1,000-ton capacity hopper. Because of the changing production rates in both the underground and surface mines as well as the great difference in size consist, it was decided that it would be necessary to divorce the plant as much as possible from production. For this reason, twin 6,500-ton raw coal silos were constructed to separate the two coals as well as allow blending. The preparation plant itself, was to be designed to Operate two shifts per day, five days per week, at an average rate Of 2,200 tons per hour with a maximum rate of 2600 to 2800 tons per hour. Coal was to be delivered to our customers by means of rail haulage at both unit train rates and plant process rates. The surface mined coal, after discharging into the 1,000-ton R.O.M. hopper, is fed by means of two 72" wide reciprocating feeders onto a 72" wide steel cable belt. The steel cabled conveyor discharges the R.O.M. coal onto a fixed grizzly, scalping out a percent of the 6 x 0 coal. The balance is fed to a 12' x 27' Rotary Breaker. The material passing through 6" round hole screens is added to the 6" x 0" from the scalping screen. The 6" x 0" is conveyed to the top of one of the 6500 ton silos designated for surface coal. Material that is not broken to pass the 6" breaker screen is discharged to a bin and disposed of with plant refuse.

Jan 1, 1976

By James R. Walsh

While flue gas scrubbing with alkaline solutions is one of the oldest methods of removing SO2 from flue gas, a satisfactory scrubbing system for this application, which is both technically and economically viable, has yet to be developed. Interest in alkaline scrubbing has continued because of its basic simplicity and comparatively low capital cost. Drawbacks of alkaline scrubbing are the tendency for accumulation of solids to plug the system, scaling, corrosion and erosion problems and the need to dispose of large quantities of waste. Discussed are instrumentation loop applications to a 90,000 cfm alkali scrubbing test facility, the process problems the instrumentation engineer faces in. selection of equipment, the performance of the instruments selected and the problems encountered during start-up. The Test Facility incorporates centralized electronic process control with a digital data acquisition system provided to log process variables and print off normal (alarm) conditions. Flue gas SO2 content is measured by ultra-violet analyzers and analysis of the total gas stream is by gas chromatographs.

Jan 1, 1973

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

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

By Jan M. Andersson

This paper will discuss the background information about rock cutting tools used for soft ground excavation, their history and evolution. A brief review of the current state of the art manufacturing technology will be presented and some parameters related to the material used in the drag and conical type pick cutters and their impact on their life and performance will be explained In addition, the selection procedures for choosing the right bit for excavation of a particular geological formation will be reviewed A set of guidelines will be provided for initial selection of the tools for various machines and rock formations. The methods for estimation of the tool life in various materials and the influencing parameters along with limitations of the rock cutting tools will be discussed as well.

Jan 1, 1998

By Ed L. Schrader

Kaolin of the Georgia-south Carolina "Clay Belt" are used in a wide variety of filler and extender applications. Sane applications specifically require the performance enhancement delivered by incorporating kaolin into the product formulations. Other applications are less specific in requirements for an inert mineral filler. In order to be a feasible candidate for filling and extending, a mineral or mineral product must be generally inert (except where specific interaction with the filled medium is desired), inexpensive, easily handled and transported in a variety of forms, accessible to a broad segment of the desired market, and must exist in sufficiently large ore reserves as to assure production for many years. Of course, more precise requirements are necessary in specific applications of a prospective filler material.

Jan 1, 1985

By John W. Harshbarger

The Upper Santa Cruz Groundwater Basin lies in the drainage area of the Santa Cruz River in Arizona, and extends upstream from the community of Rillito to the international boundary, see Figure 1. The principal water use area lies between and includes the cities of Tucson and Nogales. A subsurface geologic barrier near Rillito provides a reasonable dividing line between the groundwaters of the Upper and Lower Santa Cruz Basins. The Upper Basin contains the Tucson and Sahuarita-Continental critical ground- water areas. The Tucson Water Use Area encompasses the Tucson metropolitan area and extends from the northern boundary of the Basin south to the boundary between the Tucson and Sahuarita- Continental critical groundwater areas near San Xavier. The San Xavier Use Area is the San Xavier Indian Reservation. The north and south boundaries of the Continental Use Area coincide with the limits of the Sahuarita-Continental critical groundwater area. The first study of groundwater in this valley was initiated by Dr. G. E. P. Smith in 1905 and the results were published in 1910, followed by a series of bulletins of the Arizona Agricultural Experiment Station, University of Arizona, covering the general field of groundwater development and utilization. Beginning in 1939, the Ground Water Branch of the United States Geologic Survey, in cooperation with the Arizona State Land Department and the Arizona Water Commission (AWC), have published a series of bulletins concerning groundwater occurrence and utilization in various areas of the state. In 1975 and 1977, the AWC published the "Arizona State Water Plan" (1), comprising a detailed inventory of water availability and use in the State of Arizona, and a study of alternative futures.

Jan 1, 1979

By L. Barry Phelps

Coal mine drainage, which occurs predominantly as acid mine drainage in the East, is a serious contaminant of surface and groundwater systems (Appalachian Regional Commission, 1969). Treatment of this pollutant can be very costly and is often difficult to achieve (Appalachian Regional Commission, 1969). A major source of coal mine drainage in surface waters is polluted ground- water discharge, therefore treatment of surface water does little to combat groundwater contamination. Prevention of the formation of coal mine drainage and restricting its entrance into the groundwater protects both surface and groundwater systems. The formation of acid mine drainage may be inhibited by preventing contact between toxic spoil, air, and water (Grim and Hill, 1974). This may be accomplished by surrounding toxic spoil with a relatively impermeable barrier, thus restricting groundwater and oxygen flow through the toxic zone. Toxic spoil packages constructed in this manner would minimize the entrance of acid mine drainage into the groundwater system, allowing dilution by uncontaminated water to occur. If successful, concentrations of coal mine drainage would be reduced below toxic levels.

Jan 1, 1981

By David G. Chedgy

It is probably safe to assume that most operating personnel working in coal preparation plants are exposed to noise levels in excess of the MESA Noise Regulations unless they are wearing ear protection. The excessive in-plant noise levels have received little attention until recently. Noise control has not been an important design consideration in existing coal preparation plants. This situation is rapidly changing as plant owners and plant operators are requiring that new coal preparation plants be designed to provide a safe noise environment for operating personnel and also meet MESA Noise Regulations. This paper discusses the noise exposure hazards, noise generating mechanisms, and some possible noise control solutions for typical coal preparation plants.

Jan 1, 1975

By Timothy J. Matthews, Ted Klemetti, Timothy J. Batchler, Daniel J. McElhinney

Unplanned tailgate collapses and difficult mining conditions in U.S. longwalls have necessitated the use of innovative tailgate support systems. Severe ground movement induced by longwall loading has caused operators to utilize secondary support systems that can match yield and strength requirements with expected ground reactions. Standing support performance requirements depend on the nature of the ground deformation behavior, and therefore require a good understanding of the interaction between the support and ground to achieve an optimum roof support system. Mining operations often employ standing support strategies that are unique to their mine. A support system that functions successfully in one situation could fail in another situation. It is important that the design and installation of the standing support systems provide the highest probability of preventing roof falls. Pumpable roof supports are currently being used to provide a safe working environment for longwall mining. A full understanding of the impact of these performance characteristics is necessary to optimize the support application and to provide a foundation for making further improvements in the support performance. Optimization of roof support applications helps to ensure the protection of mine workers by preventing roof falls due to inadequate support design. The load capacity and stiffness characteristics of a 27-in.-diameter pumpable support system was evaluated using the National Institute of Occupational Safety and Health (NIOSH) Mine Roof Simulator (MRS). Previously, tests to evaluate the performance characteristics were performed on a single pumpable support. The goal of this new research is to evaluate the performance characteristic effects of the pumpable support system and compare it to an individual pumpable test examination. This research of a pumpable support system configuration simulates the current installation practices in U.S. coal mines.

Jul 1, 2023

By J. C. Irvine

Meramec Mining Company is a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp. which mines and pelletizes iron ore at the Pea Ridge Mine. The Pea Ridge property is located near Sullivan, Missouri, about 70 miles southwest of St. Louis. The orebody was delineated in the mid-1950's by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drilling in 1956 and 1957 confirmed that the orebody was mineable, Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Production commenced in April 1964. The orebody is overlain by about 1300 feet of flat-bedded sediments. It is tabular, about 2600 feet long and up to 600 feet thick. It dips about 80° and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 150 foot intervals. Crosscuts were driven across the orebody on 190 foot centers. Banks of scopes were mined between the crosscuts by a modified shrinkage stoping method. This was done by undercutting a 40 foot wide by 150 foot long block and blasting horizontal "lifts" drilled on 5 foot intervals. A 65 foot still was left between levels which contained the slushing drift, fingers, crown pillar, and adequate thickness for support. By 1971 a ?lattice? of pillars had been left and mining had progressed to the point that an orderly pillar recovery program was necessary.

Jan 1, 1976

By Charles L. Allie

In June, 1980, Reserve Mining Company completed a 370 million dollar construction project which included process modifications required to reduce the silica content from nine to five percent by weight and increase the iron con- tent of Reserve's pellets from 62.5 to 65.8 percent by weight (dry basis); process modifications and additions required to implement "best available technology" water quality controls; and installation of "best available technology" air emission controls. The new process began operation on June 24, 1980 and is being optimized to meet design objectives. Dry cobbing, fine screening and flotation additions to the concentrating process have reduced the silica content of Reserve's pellet from nine percent to five percent by weight. The process reclaims and reuses all process waters, stores all tailings in a zero discharge tailings basin and meets all air emission standards.

Jan 1, 1980

By L. G. Austin

A program has been written to combine the Klimpel-Austin approximate model for liberation in a binary system with a mill model applicable to ball milling. The model applies for a system where the volume fraction of valuable B in host matrix A is small. The mill model is treated as an equivalent series of fully-mixed reactors. Twenty-three classes of material are considered: pure A; pure B; unchanged AB; 0-10.2 B in AB either as surfac- exposed B or as wholly-contained grains of B; 10-20% B in AB, surface-exposed or wholly- contained;etc.. to 90-100% /b ub /ABm syrface-exposed or wholly-contained. Illustrative cases are given making reasonable assumptions concerning relative rates of breakage and primary breakage distribution functions for each class of material.

Jan 1, 1986

By B. M. Eiter

The impact of fatigue is seen not only in its effect on the job performance of haul truck operators but also on the health of the operator and the productivity at the mine site. Its impact can even extend outside of the mine site to the health and well-being of the surrounding community (Fourie et al., 2010). In this paper, a case study of a small, surface mining organization is presented. The goal is to highlight the fatigue risk management system implemented at the studied mine site. Mine safety personnel who were interviewed discuss the changes made to the infrastructure of the mine, to administrative areas such as the number of shifts and the use of vacation time, as well as the implementation of new technology into haulage vehicles. This paper reviews how these changes are supported in the research literature.

By B. N. Everett

The growing role of Colombia as a major supplier of internationally traded steam coal has been widely discussed in the industry. Although Colombian coal has been produced and exported in small quantities for a number of years, Colombia's position as a major coal exporter began in earnest in 1985 with the first commercial exports from the Cerrejon North Block Project. By way of background: ? The mine is located on the Guajira peninsula in Northeast Colombia, about 90 miles from the coast, as shown in Figure 1. ? In addition to the mine itself, the project involves construction of a dedicated railroad and port, which form a modern, efficient and integrated system. ? The Project is embodied in a 50/50 Association Contract between International Colombia Resources (Intercor), a wholly owned subsidiary of Exxon, and Carbocol, the Colombian state coal company. Intercor and Carbocol share investment and operating expenses equally and Intercor has the additional responsibility of managing operations on behalf of both partners. Each partner is entitled to 50% of the coal produced, and each markets its share separately. ? The project is a multi-seam, truck and shovel operation, based on large reserves (2-3 billion tonnes') of high quality bituminous coal. ? Exports began in the first quarter of 1985 at modest levels through what we call Early Coal Facilities, designed and constructed apart from the permanent coal handling system. ? Production will increase rapidly in each of the next few years, reaching the design capacity of 15 Mt per year in 1939. Based on available resources, however, the resource will support expansion, and continued growth n production into the 1990s appears feasible.

Jan 1, 1986

By Ande&apos, Bengt n

The first Model 480 SALA-HGMS R carousel high gradient magnetic separator was recently manufactured (Figure 1). The carousel's design was based upon the high gradient magnetic separation technology developed and owned by Sala Magnetics Incorporated, Cambridge, Massachusetts. After a short (approximately 300 hours) in-house mechanical and electrical test and break-in period, the carousel was ready for a long term demonstration test. This paper describes the demonstration plant and operational performance of the Model 480 HGMS treating a low grade iron ore tailings. A location convenient to Sala International's main office in Sala, Sweden was desirable because Sala engineers and technicians were to operate the demonstration plant. The Strassa iron ore mine in central Sweden was chosen because it fit most of the criteria that was important for a good demonstration site. The Strassa mine is part of the state owned Svenskt Stal AB. The mine produces about 320,000 tons of iron concentrate per year consisting of 200,000 tons of magnitite concentrate and 120,000 tons of hematite concentrate.

Jan 1, 1980

By Andrew J. Gaber

Beginning in 1961 the Coal Research Bureau of the West Virginia University School of Mines began a program of research to investigate the potential of the magnetic separation process to remove sulfur of the pyritic form from pulverized coal. Using a so-called isodynamic magnetic separator (1) it was demonstrated that significant quantities of pyritic sulfur could be removed from coal. Pulverized coal samples representing several different coal seams were individually screened into sizes in the 48 mesh by 200 mesh (297 to 74 microns) range, and untreated and steam-air treated samples of each were processed by the magnetic separator. The results of the investigations to magnetically remove pyrite from coal were published in two separate reports issued by the Coal Research Bureau (2,3), The isodynamic magnetic separator was designed about 30 years ago for use in geological and mineralogical laboratories (1,4,5) to isolate one or more minerals exhibiting only small differences in their magnetic propertiese Specifically, to accomplish this, the air gap of the unit was designed to provide a nearly constant magnetic gradient both laterally and longitudinally in the separating zone. It is reported that mineral concentrates of 99.5 percent purity can-be produced on closely-sized fractions after several cleanings through the unit. The isodynamic magnetic separator, therefore, was not expected to process coal containing sizes smaller than 200 mesh, or to demonstrate commercial capability to process pulverized coal. The primary purpose of the Coal Research Bureau in using the unit was that of determining if the magnetic process could remove pyrite from coal, and thereby encourage further research directed toward development of acceptable full-scale processing equipment.

Jan 1, 1969

By R. Kuhns

The western coast of South Africa hosts a significant diamond resource within gravels associated with paleo-shorelines, channels and transgressive lag deposits. Diamonds were transported from the Kaapvaal Craton to the coast largely during the late Cretaceous via the paleo-Orange river systems. On shore and near shore environments are underlain by Proterozoic metamorphic and granitic rocks which provide gully and pot hole-type traps associated with paleo-channels, basins, bays and beaches. Inner and mid-shelf environments are covered by diamondiferous, poorly lithified upper Cretaceous and Tertiary sediments. The upper Cretaceous gravels were reworked along the sediment-starved coast during much of the Tertiary. Reworking was mainly achieved during marine flooding events which produced lag gravels along the paleo-shorelines. This process reworked the diamonds from low-grade Cretaceous gravels and concentrated them in higher-grade Tertiary gravels.

Jan 1, 1995

By Peter T. Luckie, Joseph W. III Leonard

INTRODUCTION In a bed of fine coal, water can be found in the interstices created by the particle surfaces and in the pores within the coal particles. The extent of the interstices voidage is obviously a function of the particle sizes and morphologies, as shown by the fact that the bulk density of coarse coal is greater than that of fine coal. After drainage, the voids can be completely or partially filled. By increasing the suction pressure across the bed, as is done in vacuum filtering, the quantity of water in the voids can be reduced. But eventually the increased suction pressure has little influence on the moisture reduction. In addition, the moisture in the pore structure of the coal is unaffected by the mechanical removal of the moisture in the interstices. When this condition is reached, the total moisture content of the coal can only be reduced further by evaporating the water. Evaporative drying is done for one or more of the following reasons: 1. To avoid freezing difficulties and to facilitate handling during shipment, storage, and transfer to the points of utilization. 2. To reduce the heat loss due to evaporation of the moisture from the coal during the combustion process, thus increasing heating efficiency. 3. To decrease transportation costs. The heating value of coal is equally improved for each percentage point reduction in the noncombustibles in coal. It is estimated that a reduction of 1%22 in the total moisture in the coal increases the thermal efficiency of the heat-raising process by about 0. Some coal buyers consider 1 % moisture content to be the equivalent of a 4% ash content.23 It is important to avoid drying the coal to too low a moisture level, since the complete removal of surface moisture to produce a bone dry product can result in a serious dust, wind lost, and safety problem. Surface moisture con- tents of 3% or less for shipped coal are not uncommon. However, the evaporatively dried coal may be lower, Y, to 1%, since the mechanically dewatered coal may be higher. Evaporatively dried coal can be treated (sprayed) with chemicals in order to control the dustiness and any freezing difficulties, thereby providing the producer with more operational flexibility. The evaporative drying of coal is essentially a process of simultaneous transfer of energy, usually in the form of heat, and water, in the form of vapor, from the coal. In order to remove the undesired moisture, it is necessary to raise the temperature of the water from the entering temperature of the feed stream, TF, to the boiling point of the water (usually taken as 212"F), vaporize

Jan 1, 1991

By Thomas E. Shufflebarger

Glass is one of man's most versatile creations. ---The earliest glaze known---is about 12,000 B.C. ---The oldest pure glass is a moulded amulet of about 7,000 B.C.---. Indicative of a knowledge of properly-proportioned, quality raw-material, means were found for producing the first relatively clear, transparent glass at about the beginning of the Christian era. The fundamentals of almost all the modern, technical processes of manufacture were established during the 3rd Century. (1) Glassmaking, practiced in Colonial America, was moved toward the Ohio Valley in response to increasing demands for fuel. This move, beginning in the 1860's resulted in the use of the glass sands contained in the Oriskany (and St. Peter) sandstone. (2) Glass sand,(3) virtually pure, fine-to-medium-grained quartz sand(4) is the principal raw material used in the manufacture of glass; it constitutes proportions which range to a maximum of about 72% and seldom is glass made without it. However, acceptable glass sands are, invariably, those which most nearly fit the glass-producer's economic needs and any method used to evaluate a raw material's potential must finally prove equatable with such needs. As used in the United States, source-rocks of glass sand include sandstone, unconsolidated silica sands, quartzites and vein-quartz. Admixtures of silica sand and other (principally clay and feldspar), readily separable components have been proven viable. However, the silica sand, refined by nature to a high degree of purity, is the desirable product of exploration. If the term, standard, is justified by long-continued use, the Oriskany - and St. Peter sandstones are, together, the nonesuch of the glass sand industry they provide the principal tonnage of melting sand. According to Grimsley,(5) Ottawa, Illinois, glass-sand is considered the standard glass-sand of the West and the analyses of Berkeley Springs sand (Oriskany) compare most favorably with this sand.

Jan 1, 1974

By Melvin A. Jennings

Even though the title of this paper is quite long, the ideas presented herein are very simple. In fact, it occurs to me to be somewhat humble about coming before such a distinguished group with such a simple presentation. However, the application of these ideas has proven useful in at least two copper flotation plants. The purpose in presenting this paper is to inspire some interest in the great stores of data chat the Mining Industry often accumulates without making any attempt at analysis. Unfortunately, many milling engineers rely rather heavily on percent recovery as a strict measure of a concentrator's relative performance. The fallacy of this reasoning is not difficult to see since, by definition, percent recovery is a function of feed assay, tails; and concentrate grade. If copper tailing is constant, and it very nearly is for a strictly sulfide copper ore, percent recovery must fluctuate almost directly to the feed assay. The point that I wish to make here is that it is very seldom justified to compare the percent recovery figure from one period of time to another without considering the difference in feed assay. The problem of comparing results is even further complicated when the ores are of mixed oxide and sulfide mineralogy.

Jan 1, 1978