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By David N. Hedges

This paper discusses the actual recovery and installation of a longwall in a bituminous coal mine in Western Pennsylvania. The coal bed mined is in the Lower Freeport seam which averages 42 inches (1.07m) in height locally. Drainage from the mine is approximately 3.5 million gallons of water per day (13.25 million liters per day). This water deteriorates the mines predominately fire clay bottom making it extremely difficult to move equipment. Support removal is accomplished by pulling the 6 ton (5443 kg) shields out under their own hydraulic power and transporting them by track over four miles (6437m) co the next longwall panel. This paper also discusses the changes that will be made to aid in the succeeding longwall recoveries and installations.

Jan 1, 1984

By R. D. Smith

The Bureau of Mines (BOM) evaluated the technology, capital investment, and annual operating costs of emission control for further reductions of lead in ambient air at the East Helena, Montana, smelter, and the Glover, Herculaneum, and Buick smelter-refineries in Missouri. This study seeks to analyze the technology available to the domestic primary lead industry for attempting to further reduce lead emissions below 1985 levels. Through enclosure of concentrate unloading and handling areas and all lead-processing buildings, followed by venting these enclosures to the atmosphere through baghouses and the application of appropriate local controls, emissions can theoretically be reduced by a total of 2,424 pounds of lead per day from 1985 levels at the three Missouri operations. Using conventional technology, aggregated additional capital costs of $250,014,000 and annual operating costs of $18,507,800 are estimated for the four operations, resulting in a total cost of $0.063 per pound of refined lead. Continuous drossing techniques would assist in meeting OSHA standards in process areas. Continuous drossing would change capital and operating costs to $258,344,800 and $16,104,700, respectively, for a total cost of $0.062 per pound of refined lead. First quarter, 1986 open market price for refined lead was $0.184 per pound.

Jan 1, 1986

By Richard E. Sabaski

The demand placed on the mining industry in the past 15 years for increased production, coupled with lower grade ore and higher cost capital, has resulted in a noticeable increase in grinding mill sizes. The pressure to increase mill size continues. During the mid sixty's and early seventy's, ?larger? conventional ball mills were in the range of 13 - 16-1/2 f t. diameter (4 - 5 m) increasing by the mid 70's to 18 ft. (5.5 m) diameter. At the same time, autogenous mills expanded from 24 - 32 ft. (7.3 - 9.1 m) to the present 34 and 36 ft. (10.b and 11.0 m) diameters. These size pioneers of 16-1/2, 18, and 36 ft. diameter mills from Allis-Chalmers have been free from structural problems, because they were analyzed, modeled, designed, and field tested to meet the challenge.

Jan 1, 1981

By J. Philip Bromberg

Any effectuation of a viable program of air pollution control will' require a basic understanding of the chemistry of atmospheric processes, the fate and migration of SO2, and the ultimate health effects of the sulfur oxides themselves. The migration and ultimate fate of Sop discharged into the atmosphere will be discussed with particular emphasis on the various conflicting views and how these relate to our present emission control regulations. Our present lack of knowledge is a severe detriment, and consideration will be given to those particular areas where future research may best be directed to provide the answers to a number of perplexing questions.

Jan 1, 1976

By R. B. Elver

The St Lawrence Seaway, a deep-water canal system between Montreal and Lake Ontario, was opened for navigation on April 25, 1959. Prior to the opening of the new Seaway, small boats known as canalers and a few deep-sea carriers, primarily from Norway since World War II, were the only ships able to journey past Montreal into Lake Ontario through the old Canadian St. Lawrence canal system. Thus it was necessary for many tons of bulk and general cargo to be shipped all-rail to and from Atlantic ports or to be transhipped to and from canalers. With the opening of the new Seaway, large Great Lakes bulk carriers and a much larger percentage of the world's shipping fleet are now able to sail directly from the sea to the heart of industrial North America. In this paper, the historical events leading up to the construction of the Seaway and the physical nature of the Great Lakes transportation system are briefly reviewed. Following this, an attempt is made to evaluate the effects of the new Seaway on Canada's mineral industry. Because the Seaway has been in operation for only one season, this paper must be considered only an interim report. Because of the importance of the Seaway to iron ore, this commodity will be dealt with in detail, although the effect on other mineral commodities such as coal, petroleum, nonferrous ores, iron and steel will be mentioned. At this time the author wishes to thank the various company and government officials who provided both factual information and critical opinions.

Jan 1, 1960

By Paul Soros

There has been a steady increase in shiploading capacities in the last 20 years. One of the reasons has been the general increase in ship sizes, with the largest vessels currently around 280,000 DWT in the bulk trades. Obviously, the large vessels can accept loads faster and their economic utilization justifies higher loading rates, up to a point. The high capacity facilities have loading rates far beyond this point. The economic logic of these facilities is based on the fact that for a given annual throughput the capital and operating cost of a single high capacity system is less than that of two smaller systems. Thus, where the projected throughputs are large, there is a great incentive to reduce capital costs by maximizing capacity. Tubarao, in Brazil, designed for a 100 million ton per year throughput with one finger pier and one high capacity berth is a successful demonstration of this approach. In the current state of the art, the high capacity systems have shiploading capacities in the range of 10,000 - 16,000 TPH, for annual throughputs of 20 million tons and larger per loading berth.

Jan 1, 1982

By Wen Lan Chen

In the plant where the 800m diameter cyclone was tested, 500 STPH of 50m x 0 raw coal was pulped in a heavy medium tank and pumped originally to four 700mm heavy medium cyclones. For the test one of these units was replaced by an 800mm diameter cyclone with the subsequent feed pressure for both units sets at approximately 15.0 psi. Since this was a feed to zero plant, an appropriate bleed of heavy medium to the dilute medium circuit was established to en- sure that the cleaning of the 500m x 0.5mm raw coal would not be adversely affected by high levels of 0.5mm x 0 fine coal present in the heavy medium. The medium gravity measured varied between 1.41 and 1.43. During a three-hour test on June 10, 1983, the raw coal sample was collected from the feed belt while the clean coal and refuse samples were collected separately at proper locations along the product streams of the two different diameter cyclones. All of these samples were shipped to the laboratory for screen and sink- float analyses. The raw coal screen and ash analyses are shown in Table 1.

Jan 1, 1984

By Willard C. Lacy

Survival of the mining industry as a viable economic entity in the United States is being seriously threatened by declining grades of ore reserves, rising operational and capital costs, and increased governmental taxes, restrictions, and subsidizing of third world competition. Discovery of major high-grade deposits is becoming increasingly unlikely. International competitiveness of the U.S. deposits will depend upon 'Yankee Ingenuity', innovations in mining methods, metallurgical treatment, and financial arrangements, and a viable national strategy. This will require not only effective management for control of operations, but management that encourages breakthroughs in technology, financing, labor relations, and in public and governmental understanding. The mine geologist and the exploration geologist both have important roles to play in all aspects of this fight for survival of the industry.

Jan 1, 1984

By Shi Wang, Hongbin Liu, Meiliang Wei, Wu Liu, Xiaojun Wang, Yuxian Ke, Xuepeng Song, Tiejun Tao

The study investigates the influence of water migration on the fluidity of fresh cemented paste backfill (CPB) with and without flocculant under variable shear rates. The research utilizes nuclear magnetic resonance (NMR), rheological tests, and scanning electron microscopy (SEM) to analyze water migration, rheological properties, and microstructure evolution. The findings reveal that water migrates to more stable states as shear rate increases, and the addition of flocculant enhances water migration and increases shear stress and apparent viscosity. The study establishes relationships between water content and rheological parameters, providing insights into the flow characteristics of fresh CPB for pipeline transportation.

May 4, 2022

By J. L. Coburn

The Huari-Huari ore body is located in Southern Bolivia, at an elevation of approximately 3,800 meters above sea level. The ore is a complex massive sulfide, slightly marmatitic sphalerite, in a fault fissure filled hypo- thermal vein in Paleozoic shales and quartzite-altered sandstone. The sphalerite has common associations with pyrite and quartz and is frequently intricately intergrown with stannite, chalcopyrite, and pyrrhotite. The sphalerite also contains numerous inclusions of pyrite, marcasite, galena, boulangerite, jamesonite, quartz, arsenopyrite, and cassiterite. A research program was undertaken to develop a flowsheet for the recovery of separate silver, zinc, and tin concentrates by selective flotation. Based on this experimental work and subsequent flowsheet development, engineering and construction of a 500 metric ton per day concentrator was undertaken. This paper describes the process that was developed for the Huari-Huari ore and the unique aspects of the Huari-Huari concentrator.

Jan 1, 1977

By Robert P. King

In April, 1975 the Colorado School of Mines began monitoring rock mass response associated with a long- wall shield demonstration project near Raton, New Mexico. This project was sponsored by the U. S. Bureau of Mines (DOE) and performed at Kaiser Steel Corporation's York Canyon Mine. Two of the results of the intensive instrumentation effort are the empirical subsidence and horizontal strain models presented herein. These models are based on data gathered during mining three adjacent longwall panels (4N, 5N and 6N). A total of 251 permanent surveying monuments were monitored over the three-panel area. Lines of monuments were oriented along panel centerlines, perpendicular to centerlines and diagonal (45O) to centerlines. Active monuments were monitored twice a week, with less frequent monitoring of stable or slow moving monuments. Primary subsidence monitoring equipment consisted of an AGA Geometer Model 75 and a Wild T-2 Theodolite with -a Wild-Sercel DI-3S distomat.

Jan 1, 1979

By A. Vilinska

The adhesion of Leptospirillum ferrooxidans bacterial cells onto the sulfide minerals pyrite and chalcopyrite was evaluated using two different physical-chemical approaches; thermodynamic and extended DLVO theory. For the parameters incorporated into calculations, the zeta potentials and contact angles of powdered solids and bacterial cells were acquired experimentally. The Hamaker constants essential for Lifshitz-van der Waals inter-action calculations were calculated following two different methods: macroscopic and microscopic. Adsorption tests were carried out at physiologic conditions to estimate the amount of cells adsorbed onto a mineral surface and the extent of alteration of that mineral surface in biobeneficiation. The free energy of adhesion was found to be negative for both minerals, indicating that the adhesion is energetically favored and preferred. The interaction energy diagrams of the total interacting force was also negative for the cases where the particles were charged oppositely; in the remaining cases, the total force was attractive after overcoming an energetic barrier caused by the repulsive electrostatic forces. Under the conditions of the adsorption test, the experimental results are in agreement with the theoretical; this suggests that the physical-chemical forces are crucial for bacterial adhesion.

Jan 1, 2011

By Elham Najafi, Sina Ghassa, Ata Akcil, Hadi Abdollahi, Sandeep Panda, Zohreh Boroumand

The present study investigates the influence of ferrous iron (as FeSO4) and ferric iron (as Fe2 (SO4)3), and pyrite (FeS2) on the ability of bacterial leaching of a high-grade sulfide Zn–Pb ore. In this regard, shake flask experiments were carried out at 5% (w/v) pulp density of the ore sample (having 40.7% Zn and 12.4% Pb initial metal content) using a consortium of mesophilic iron and sulfur-oxidizing acidophiles. A concentration of 0.04 mol/L of ferric iron in the leaching media was found to be optimum for zinc extraction without affecting growth of the microorganisms. Under this concentration, the dissolution of Zn, Pb, Cd, and As was found to be 57%, 0.2%, 0.03%, and 9.9% in 25 days. Using ferrous iron in the media, 0.16 mol/L of Fe2+ was found to be the optimum concentration for efficient bacterial growth and metal dissolution (54.6% Zn, 0.08% Pb, 0.03% Cd, and 10.2% As) from the sample in 25 days. On the other hand, using pyrite as the source of energy for bacterial growth, an initial 12-day lag period was observed when compared to the effect of ferrous iron in the media. Under the optimum concentration (test with 0.24 mol/L iron in the form of pyrite), the dissolution of Zn, Pb, Cd, and As was found to be 39.8%, 0.1%, 0.03%, and 10% in 25 days. The surface chemistry analysis indicated formation of a sulfur layer over the particle surface that hindered reagent diffusion and affected metal recovery through bioleaching.

By William G. Freeman

In comparison to many mineral commodities, barite has a low unit value. It is mined and sold in low volumes in high-grade (90-98% BaSO4) forms to basically one industry that uses approximately 75% of the free world production. Based upon the above criteria, all factors must be considered in determining if a barite deposit is economical; such as, the type of deposit, mining and milling methods required, transportation to the markets, tariffs, local political conditions, etc. In other words, barite mines are made--not found. The length of this paper precludes any indepth discussion of geology, mining and metallurgy of barite. Most of this information can be found in data already published, such as D. A. Brobst's bulletin1 on the Barite Resources of the United States.

Jan 1, 1971

By Paul R. Kruesi

A patent-applied-for process has been developed in which zinc chloride is extracted by tributyl phosphate in kerosine, and stripped with ethylene glycol or ammoniated ethylene glycol. The anhydrous zinc chloride is precipitated from the glycol as the ammine and zinc chloride is prepared by thermal decomposition with recycle of ammonia. The process permits the production of high grade zinc from scraps and drosses. Its applicability to the recycling of zinc will be illustrated.

Jan 1, 1984

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Bruce R. Lipin

Chromite, the only ore of chromium, occurs almost exclusively in mafic-ultramafic rocks and their weathering products. However, not all kinds of mafic-ultramafic rocks are favorable hosts for possible low-grade chromite deposits. For instance, chromium-bearing basalts are not favorable because they commonly contain < 0.3% Cr2O3, and chromite is almost always less than 50 um in diameter, if it is present at all. Chromite-bearing ultramafic rocks in stratifonn and alpine complexes tend to be either ore grade (> 25% Cr203) or very low grade (< 2% Cr203); intermediate rocks are scarce. In most complexes large tonnages exist only in the < 2% Cr2O3 group and the composition of the chromite is commonly unsuitable for metallurgical use. Some nickel laterite districts contain more than 108 tons of material with 2.0% to 5.5% Cr2O3, and hence, chromium could be a byproduct of nickel in some laterite districts.

Jan 1, 1982

By D. J. MacKinnon

Smooth, compact, dendrite-free, 24-h zinc deposits have been electrowon from zinc chloride electrolyte using a diaphragm cell. The cell consisted of sealed anode compartment con- taining DSA anodes which were separated from an open cathode compartment by Dynel cloth membranes. Fresh electrolyte was fed to the cathode compartment where zinc deposition onto an aluminum cathode was aided by air-sparging. The catholyte passed through the diaphragm into the anode compartment from which chlorine gas and spent anolyte were withdram.

Jan 1, 1980

By John E. Caffrey

United States Steel Corporation began mining coal in southern West Virginia at the turn of the century, and until the late 40&apos;s production was confined to underground mines. Surface mining was initiated in our southern West Virginia -holdings as a means of recovering coal that could not be obtained by underground mining methods. After observing and evaluating surface mining methods employed by other companies, some of which operated on lands of U. S. Steel on a lease basis, we undertook the establishment of a surface mining operation. Our method, as, initiated in the early 50&apos;s, remains basically the same today. Variations have come about primarily by equipment modifications. The method consists&apos; of a pioneer cut to establish the coal outcrop, a drill bench for drilling, blasting, and removal of the overburden from the top of the coal. The land is rugged except for the flood plains of some of the larger streams. The mountains have steep slopes, averaging 30 degrees, with relatively sharp ridges and very narrow valleys,. Geologically, the coals_ lie in the Pocahontas formation of the Pottsville group in the Pennsylvanian system. This means that the coals in our area of operation lie at the base of. The coal measures in the Appalachian range. The coal that we have mined is primarily the Pocahontas nos. 3 and 4 seams. We have mined negligible amounts of other seams in the Pocahontas groups above the no. 4 seam. The Pocahontas no. 3 seam, which is the lowest mineable seam geologically in this area, lies anywhere from 700 feet below: the valley floor up to 400 feet above the floor. This is occasioned by the Anticlinal formation which gives us grades Of about 9% at the anticline, decreasing to about 1 to 1-1/2% on the extreme flanks. We have the opportunity to engage in surface mining where the coal seams intercept the surface above the valley floors.

Jan 1, 1972

By D. L. McCain

In considering slurry transportation for Consolidation Coal Company&apos;s Loveridge Mine in Northern West Virginia, many things were taken into account. The advantages of slurry transportation in comparison to shuttle car, belt and rail transport have been given in previous publications." These include increased safety and the potential for increased productivity. In addition to these considerations, the slurry transport system installed in an existing mine provides an alternate or additional transportation mode to the one already existing. In the Loveridge Mine, this is rail haulage. Since the rail system is a single track system and the distances involved are several kilometers, it is not always possible to provide for haulage of loaded cars, return of empties, supplies and maintenance in an efficient manner. Slurry transport of the coal from the production faces to the preparation plant allows the rail system to fulfill the remaining transportation needs in an efficient manner. Consol initiated development work in slurry transportation in the early fifties leading to the construction and operation of the 174-km (108-mile) Cadiz-to-Cleveland pipeline which carried 1,200,000 metric t/y (1,300,000 t/y) from 1957 to 1963. Development work on coarse coal slurry transportation for underground mines started in 1970 with feasibility tests on coarse coal at Conoco&apos;s Ponca City, Oklahoma, slurry test facility. A series of tests were conducted in several locations as development work proceeded on the various subsystems, mechanical equipment and controls required to produce a working system to follow a continuous miner in a room-and-pillar type mine, then to combine the coal production from several mine faces for mainline transportation to the preparation plant. The face system, including the crusher-pump vehicle, hose hauler vehicle, slurry booster pump units, return water unit and 900 to 1,200 m (3,000 to 4,000 feet) of pipeline was brought together for the first time and tested as a system at Consol&apos;s Robinson Run Mine in northern West Virginia in 1975. A pipeline wear test at the Loveridge Mine where 300,000 tonnes (330,000 tons) of minus 127-mm (5-inch) run-of-mine coal were pumped from the raw coal bin into the preparation plant and a large scale test in Ponca City on a multiple feed sump where several slurry flows could be combined to provide surge capacity and a constant reclaim rate for mainline transportation completed the preliminary testing for a complete system. A decision was made to commit the system to a mine, and the south half of the Loveridge Mine was chosen. This site required the transporting of coal from one longwall and two continuous miner development faces. To prevent startup and unexpected problems in the slurry system from shutting down production of coal from the longwall face, the rail car loading point from the main section belt was chosen as the feed point for the slurry system and the design would allow for rapidly switching from one transport mode to the other. Scheduling of required development entry work and the developmental nature of the system led to a decision to install only one continuous miner face slurry transport system rather than two for this first system. The type of continuous miner in use at the mine for the foreseeable future was the Joy 2BT6 twin boring machine. This machine cuts a stable, arch-shaped entry and was therefore allowed to mine about 26 m (85 feet) of entry before roof bolting was re¬quired. Since peak production from this type of machine was not high, a smaller transport system could be used with it.

Jan 1, 1981