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By K. J. Shoop

An analysis of an iron ore processing plant showed that its performance was affected by seasonal temperature variations, resulting in reduced grinding efficiency in the winter months. Laboratory and pilot plant studies were conducted to determine the cause of the observed loss of efficiency at lower temperatures. Experiments were divided into two groups: 1) Drop testing, which was performed to determine the effect of brittleness and frozen pore water on the breakage characteristics of the ore. 2) Wet grinding, which was further divided into laboratory tumble tests and pilot plant scale tests. Laboratory tumble tests were performed to determine the effect of temperature on the grinding of iron ore. This was done by varying the percentage of fines and percentage of metamorphic clay material added to the iron ore grinding charge. The pilot plant scale tests were conducted as a scale-up of the laboratory tumble tests and to determine the effect of temperature on slurry rheology, and how this changing rheology affected the grinding of iron ore. The combined results of drop testing and wet grinding were used to make some predictions of the performance of plant scale autogenous grinding. In drop tests the breakage characteristics of the iron ore had no noticeable change in size stabilities for dry and water saturated samples over the temperature range studied (-25°C to 25°C). The effect of frozen pore water was not observed due to the low apparent porosity of the iron ore. The addition of fines to the test charge caused a decrease in the amount of new fines produced for both the laboratory tumble tests and pilot plant scale tests. The production of fines further decreased when tests were performed at the colder temperature. The increased energy consumption at lower temperatures was dominated by rheology phenomena. The addition of metamorphic clay material to the iron ore test charge resulted in more fines being produced in the laboratory tumble tests. This was due to the metamorphic clay material grinding easier than the iron ore. This caused an increase in the amount of fines in the mill, which in turn, would prematurely slow down the grinding of iron ore because of an increase in the viscosity of the slurry.

Jan 1, 1995

By J. J. Harrington

The pH control system at Kennecott Copper Corporation's Nevada Mines Division concentrator consists of two independent systems, lime dilution and lime addition. Lime dilution is accomplished in a mixing tank where slaked lime and water are introduced proportionately. The desired dilution of the milk of lime is determined by the electrical inputs provided by eight pH probes, one per grinding section. The inputs are averaged and applied to a time base generator. The summation output of the time base generator determines the proportion of lime and water that are to be mixed and applies this input to a voltage sensitive switching device. The switching device provides a signal to one of two relays which will add either slaked lime or water, dependent upon the summation output. The entire cycle is based on a two-minute repetition rate to allow for sufficient mix time. Responsivity is controlled by maintaining a limited maxi- mum solution level in the mix tank. Lime addition is controlled independently within each grinding section. The electrical input from the pH probe in a given section controls the solenoid valves which provide lime to each of the ball mills. The valves remain open within a time range capable of maintaining the pH of flotation feed at a given set point.

Jan 1, 1975

By E. O. Strahl

In the search for economic mineral deposits, the need for efficient mineral or chemical assays is of prime importance. Costly analytical methods and long delays in obtaining assay results severely limit the effectiveness of an exploration program. The assay problem becomes acute once a potential ore body is located and large numbers of samples are generated in close interval drilling programs. In order to avoid the delays and high costs of extensive assays, the exploration geologist or mining engineer is likely to be required to curtail his drilling program, composite samples, or restrict the required analyses to a limited number of critical elements. As a result, the definition of the potential ore body may be less than needed for final economic evaluation. The assay requirements for bauxite exploration usually involve a considerable number of total elemental and empirical analysis. Since the Bayer process for extracting alumina from bauxite involves the selective dissolution and reaction of specific minerals from the bauxite, a direct chemical analysis for elemental composition yields only a portion of the data required to effectively describe the potential economic value of ore body or deposit. To fully characterize the economic value of a bauxite requires digestion, mud separation and mud settling tests under conditions approximating a commercial Bayer plant. A close approximation of the process behavior of a bauxite also can be calculated from a combination of the chemical and mineralogical analyses. By determining the elemental chemistry, mineralogy, and Bayer plant processing characteristics of a few samples of a given bauxite, correlations can be derived that permit defining a bauxite in terms of a given set of process conditions.

Jan 1, 1977

By J. A. Harrison

Most coal being used in the United States has been subjected to one or' more processes of coal preparation. The physical and chemical characteristics of a coal have a marked influence on the ultimate uses that coal may have. The selection of the proper preparation processes may similarly be important in producing the best practical coal product for any particular use. The physical composition and distribution of physical constituents in the coal will partially determine the efficiency of the preparation process chosen and conversely distribution and concentration of coal macerals will, to a degree, be controlled by the preparation process. A number of preparation processes have been evaluated to determine this interrelation. If a coal having' certain specifications is desired, these evaluations trill assist in providing information needed for the production of this coal.

Jan 1, 1962

By R. M. Wilson

Concentrating plants, in common with other types of industrial plants, cannot operate without lectric power. The elctric distribution system and utilization equipment; therefore, must be carefully considered in all phases of plant design. Two important aspects of proper distribution and usage of electric power are the selection of voltage levels and the methods of system grounding. These are not necessarily the most important design considerations, but are perhaps more controversial than other major points, Modern plants demand increasing amounts of power which, since electric power is the product of voltage and current means higher voltages or currents. Higher voltages require more insulation, higher currents more conductor. The equipment designer?s problem is to obtain in his transformer, cable, switchgear, controller, motor, etc a balance of insulation and conductor to obtain a satisfactory design at lowest cost. To provide the most economical system, both in initial and operating costs requires the system designer to select the proper voltage level. Adequate voltage also permits expansion to meet future requirements, and results in better operation of utilization equipment, Relatively little advancement has been made in recent years in improving conductors; however, great progress has been made in improving insulation. Thus modern plants, in general, now find it more economical to use higher voltages than they did previously. For example plants which only a few years ago would have used 2400 volts as a system voltage often find nowdays that 4160 volts is their proper choice. If the plant requires 10,000 kva br more, then an even higher voltage, such as 13,800 volts, will prove economical. However, for the average concentration plant, with less than 10,000 kva demand, the choice of distribution voltages is between 2400 and 4160 volts, and 480 volts will be used for the low voltage utilization system.

Jan 1, 1958

By W. J. Martin

At the termination of mining, the quality of ground water in the ore-zone aquifer must be returned to a condition predetermined by the permitting agency. To immobilize the redox-sensitive elements (e.g., uranium, arsenic, selenium, molybdenum), it has been suggested that a reducing agent be circulated through the leached ore zone. We have conducted laboratory experiments to test the ability of sodium sulfide as such a reducing agent. Sodium sulfide lowered the redox potential of the effluent such that insoluble minerals containing the redox-sensitive elements should be stable. The uranium concentration of the solution decreased by more than three orders of magnitude. In column experiments, by the time ten pore volumes of the sulfide solution had flowed through, sulfide minerals were forming. The experimental results indicate that the addition of sodium sulfide should cause the chemical reduction of the leached sediments, thereby producing long-term stability of the ground-water/sediment system.

Jan 1, 1986

By F. Ikumapayi

In order to predict and minimize detrimental production problems due to the recycling of process water in sulfide ore processing, the influence of major species, calcium and sulfate in process water on sphalerite flotation was investigated through Hallimond tube flotation, zeta potential, diffuse reflectance Fourier transform infrared spectroscopy (FTIR) and X-ray photoelectron spectroscopy (XPS) measurements using sphalerite mineral sample. Assessment of process water species in flotation was done using deionized water, process water and simulated water containing calcium and sulfate ions in experiments. Hallimond flotation shows increased sphalerite float-ability in process water compared to deionized water, but no significant effect on the presence of calcium and sulfate ions in deionized water using isobutyl xanthate as a collector. The presence of calcium ions reduced the sphalerite?s negative zeta potential, while at higher concentrations, a charge reversal occurred, at about pH 11. FTIR and XPS studies revealed the presence of surface-oxidized sulfoxy and carbonate species on sphalerite at pH 11.5 in deionized water, process water and water containing calcium and sulfate ions. These surface species do not influence xanthate adsorption.

By Donald E. Ranta

An ore zoning technique developed by Climax geologists during the past forty years has been used to contour drill-hole assay values on cross-sections for fourteen porphyry molybdenum deposits in western North America, including the Climax and Henderson orebodies. Detailed geological studies of these large, complex deposits are essential in determining ore controls, which are the factors most critical to the zoning of MoS2 assays and to conceptualizing geologic models. The technique provides a detailed representation of the size, shape, grade distribution, and geologic continuity of the individual deposits. It demands a disciplined approach and adheres to rigorous rules while maintaining flexibility for adaptation to widely varying geologic conditions. By defining the limits that guide and control manual, computer-assisted conventional, and geostatistical reserve estimates, it exerts a strong geologic influence on the estimation process. A combination of ore zoning and geostatistical analysis of molybdenum deposits produces a significantly more accurate, precise, and defendable geologic reserve estimate than other methods.

Jan 1, 1983

By Raymond A. Bradbury

Martin County Coal Corporation's operations are located on Wolf Creek in Martin County in the eastern part of Kentucky. The mining complex was started in 1971 and has grown to a 2.273 Mt/a steam coal operation. All of the production goes to Duke Power Company's Belews Creek electric generating station in North Carolina. Production comes from four seams that are mined by deep and surface mining methods. Surface mining accounts for 55% of the production and the remainder comes from 2 underground mines. Continuous miners are used exclusively underground. The preparation plant was originally designed to take coal at the rate of 907 Mt/h but that has gradually been increased to 1134 Mt/h. Other cleaning equipment includes Daniels DMS vessels, Model 77 Deister tables, 355.4-mm R. & S. Hydro- cyclones, Krebs 254-m classifying cyclones, Daniels froth cells and in 1980, 3 Krebs 508-mm heavy media cyclones were added to recover some lost and misplaced material. CMI centrifugaldryers and a Hey1 & Patterson size 100 thermal dryer are also part of the coal preparation equipment.

Jan 1, 1981

By R. L. Thoms

Compressed air energy storage (CAES) has been initiated recently in West Germany with a 290 MW peaking-power facility utilizing twin cavities with volumes of 150,000 m3 each in the Huntorf salt dome (1). The implementation of CAES in several geologic media is being studied carefully in the United States (2); however, the use of cavities for CAES in salt domes or salt anticlines and the associated long-term stability of these cavities is the specific topic of this paper (3). The cavities represent an essential element in a CAES plant, and thus should be designed and operated to perform with "stability" over the planned life of the overall facility. As employed herein, cavity stability implies the extent to which an acceptable amount of cavity storage volume can be utilized with economic maintenance over the planned life of the overall facility. In this sense stability is relative to the planned use of a storage cavity. Figure 1. depicts the salt deposits of the United States (U.S.) (4). The Gulf Coast region includes the only confirmed salt dome basin in the U.S., although salt structures also exist as anticlines in the Paradox Basin of Utah and Colorado. Because of the relatively small amount of elevation change within the Gulf Coast, CAES in salt dome cavities appears to hold considerably more potential for energy storage in this region than more traditional "pumped hydro" systems.

Jan 1, 1979

By Roger J. Araujo

A photochromic material is one which undergoes a change of color when exposed to light and reverts to its original color when the light is removed. Review articles carrying extensive bibliographies have been written by Brown and Shaw (1) and by Schwab and Bertelson (2). Organics. Schwab and Bertelson divide the reactions in different organic photochromic materials into some half-dozen categories; Windsor (3) reduces these to the following three main classes of general interest, based on how they work. 1. Stereoisomers. Absorption of light breaks one of the chemical bonds in a. ring molecule, thus allowing the molecule to unwind and form a different geometrical arrangement. The reverse process is a reforming of the bond. Examples of this class are the spiropyrans and the anils. 2. Dyes. A triphenyl methane dye, for example, is oxidized by energetic light; the absorption characteristics of the positive ion so formed are different from those of the original electrically neutral benzene rings. 3. Triplet States. In the class of polynuclear aromatic hydrocarbons, ground-state molecules are excited first to a singlet state by irradiation, and then go, via the lowest triplet state, to an excited triplet state. Visible light is absorbed in the triplet-triplet transition.

Jan 1, 1970

By Dominc C. Torre

INTRODUCTION Rapidly increasing capacities of equipment used to load and handle material at the face require improved haulage equipment and better methods of transporting the mined materials or "muck" to the shaft or slope and out of the mine to the processing plant on the surface. Increasingly, those transportation demands are being satisfied with underground belt conveyors. Such con¬veyors provide the most readily available, most efficient, and most reliable method of carrying materials in a continuous and rapid flow from the face loading equip¬ment to the surface plant. Almost universally, belt con¬veyors are used in coal mines and those nonmetallic and metallic mines operating in relatively flat bedded deposits. Belt conveyors are available in a wide range of capacities, mainly limited by belt width, belt speed, and weight of the material to be transported. They can be used on the level, on grades, on the surface, or under¬ground. They provide a rapid and constant flow of material, transporting the material as quickly as it is mined. Length of belt conveyors is almost unlimited. Where the haulage distance exceeds the practical capacity of a single conveyor unit, two or more conveyors can be used in series. Single conveyors more than 1.6 km (1.0 mile) long are common; multiple-unit conveyors in series operation are transporting materials over distances exceeding 16.1 km (10 miles). Depending upon mate¬rials being transported, job conditions, cost of the terrain being spanned, and alternative haulage methods avail¬able, conveyors can be conceivably economical for even greater distances. Table 1 lists maximum belt speeds recommended for most materials, relating belt speeds to belt widths. Table 2 lists capacities of typical conveyors of various widths and speeds, transporting commonly mined mate¬rials of various weights. ADVANTAGES OF CONVEYOR SYSTEMS For several years after their initial introduction, belt conveyors were available only with rigid side frames or a supporting structure of steel and concrete. That con¬struction limited underground applications, due to the cost and relative inconvenience of extending or retract¬ing the belt conveyor in the confined underground environment. However, during the last 20 years, wire-¬rope side-frame belt conveyors, as shown in Fig. 1, have become widely used in underground applications. Even wider use can be anticipated with their application to the fast movement of materials in surface mining operations. Advantages of the wire-rope side-frame belt con¬veyors were realized first in panel or section belt haulage in coal mines. Soon afterward, use of those conveyors was extended to gathering and main-line haulage. The mobility of the wire-rope side-frame conveyor is the principal reason for its success. The conveyor can be extended, retracted, dismantled, moved, and reassembled easily, even in relatively confined under¬ground spaces. The intermediate structure consists of rope support stands, wire-rope side frames, wire-rope tie-off stands, and the necessary carrying idlers and return idlers. An important and very attractive feature is the ability to suspend the conveyor from the roof in an underground mine, as shown in Fig. 2, leaving more room for improved housekeeping and allowing passage of face-service equipment, haulage vehicles, and other mining equipment. The conveyor's open construction allows quick and easy visual inspection of its com¬ponents for failures or malfunctions; it also facilitates checking and correcting conveyor-belt alignment. Other advantages contributing to the attractiveness and widespread application of wire-rope side-frame conveyors in main-line, slope, and surface installations include: lower capital costs, lower installation costs, lower maintenance costs, reduced impact on idlers and belting, reduced spillage and minimal cleanup, improved conformation to undulating terrain or mine bottoms, and the ability to span most obstacles without heavy structural support. CONVEYOR CLASSIFICATIONS Underground belt conveyors are classified normally according to their function. Generally, conveyors are classified into one of four main categories: 1) Panel or section conveyors normally receive the material to be transported directly from the face-haulage equipment. 2) Gathering conveyors are secondary haulage units that usually receive material from two or more panel conveyors. 3) Main-line conveyors transport all the material mined in the underground operations to the slope or shaft. Normally, they receive the material from two or more gathering conveyors. 4) Slope conveyors generally operate in tandem with the main-line conveyor, providing continuous Table 1. Maximum Belt Speeds Recommended for General Use Run-of-Mine Hard Ores and Belt Width, Coal and Earth, Primary Crushed mm (in.) m/s (fpm) Stone, m/s (fpm) 356(14) 1.52 (300) 1.52 (300) 406(16) 1.52 (300) 1.52 (300) 457(18) 2.03 (400) 1.78 (350) 508(20) 2.03 (400) 1.78 (350) 610(24) 2.54 (500) 2.29 (450) 660(26) 2.54 (500) 2.29 (450) 762(30) 3.05 (600) 2.79 (550) 914(36) 3.30 (650) 3.05 (600) 1067 (42) 3.56 (700) 3.05 (600) 1219 (48) 3.56 (700) 3.30 (650) 1372 (54) 3.56 (700) 3.30 (650) 1524 (60) 3.56 (700) 3.30 (650) 1676 (66) 4.06 (800) 3.81 (750) 1829 (72) 4.06 (800) 3.

Jan 1, 1982

By A. W. Khair

This paper deals with the causes and mechanisms of pillar failure in a southern West Virginia coal mine. The conclusions are based on in-mine observations and analysis of mining methods, eyewitness account of the sequence of events, topographic and geologic conditions, and mechanical properties of mine roof, coal, and floor. The study indicated that massive pillar failure was caused by the collapse of the mine roof. The collapse of the mine roof was due to the combined effects of high lateral compressive stress and high extraction ratio causing excessive bending of the roof beyond its limit. It was assumed that the higher lateral compressive stress under such a shallow cover was probably due to the drastic change in the structural geology of the area and possible residual stresses. Eased on the analytical study it was concluded that the sudden release of elastic energy (shock) due to the rupture of the roof caused the squeezing of approximately eighty mine pillars.

Jan 1, 1983

By John Turner, David Carey

The Robbins Mobile Miner demonstrates the successful combination of hard rock excavation technology and the latest electronic control technology. The mobile miner, a partial face hard rock tunneling machine, features a fully integrated control system relying on a Programmable Logic Controller (PLC). This PLC controls all functions: tramming, steering, boom swing, cutting depth, conveyor operation, air scrubbers, hopper loading and unloading, etc. The machine is fully instrumented allowing the PLC to monitor the results of all commands on a continuous basis, to record data, to evaluate developing problems, or to shut the machine down if necessary. The paper describes the development of the Robbins - Pasminco automated mobile miner and early testing at Pasminco Mining's Broken Hill Southern Operations, New South Wales, Australia.

Jan 1, 1993

By H. Sevim

Placement of coal combustion byproducts into abandoned sections of underground coal mines for the purpose of subsidence mitigation has been investigated in Illinois. Several handling and transportation systems have been examined. Prototype pneumatic and hydraulic placement systems have been developed and demonstrated. The demonstration site, Peabody No. 10 mine in Pawnee Illinois, is selected as a case study. It is found that, for a project of 5 years, transporting 150,000 tons of byproduct annually from a power plant located 32 km (20 miles) to the mine would cost $1.80 per ton and placement into underground would cost $3.07 per ton.

Jan 1, 1998

By A. R. Wilde

The Muruntau gold deposit is situated in the Kyzyl Kum Desert of Uzbekistan. It is currently being mined in the world's largest open pit gold mine with production believed to be of the order of two million ounces per annum. The open pit measures about 3.5 by 2.5 km and extends to a depth of 350m. The gold ore resource in the Muruntau deposit, including production, is about 170 million ounces of gold (Anonymous, 1996). This paper presents a summary of the deposit geology and its regional geological setting. Information used in this paper has been generated from observations made on two visits to the mine and its environs, and a review of the Russian-and English-language literature.

Jan 1, 1999

By M. G. Hester

The ore reserve modeling of stratiform uranium deposits to be mined by selective underground mining methods must be done with consideration of the erratic nature of the ore forming processes that produced the mineralization. Estimation methods that model the uranium grade as a smooth transition between high grade and low grade drillhole intercepts tend to greatly overestimate the ore tonnage. Care must be taken to first establish realistic ore pod geometries and then estimate the grade of the reserve units within the pods.

Jan 1, 1985

By C. D. Litton

This paper describes two new gas diagnostic techniques which have been used successfully to determine: 1) the combustion status of underground mines which have been sealed due to fire (Littol1 Ratio); and 2) the location of fires in abandoned, underground mines or waste banks (Kim Ratio). The Litton Ratio makes use of the rate of change of CO, calculated on both an air-free and methane-free basis, from mine gas samples and compares these values with those expected from a normal, combustion-free environment. The Kim Ratio uses the concentrations of total hydrocarbons, C2 ?C5, less the methane, to arrive at ratio values which increase as combustion intensity increases. The details of these two new techniques, their application to real situations, and their limitations will be discussed.

Jan 1, 1989

By Walter von der Linden

On April 14, 1975, Consolidation Coal Company began to operate a longwall face in their Shoemaker Mine using shield supports. This was the first time that shield supports were used in the United States (1) 1/. Shortly thereafter, Kaiser Steel Corporation started a shield longwall face at their York Canyon Mine (2). From October 1975 through May 1976, three more longwall faces were commissioned in mines located in south-western Pennsylvania and northern West Virginia. In June 1976, four installations had completed their first panel and two installations had completed the subsequent move to the second panel where coal production is now well under way. During this time (April 75 to June 76) the total number of shield support face operating months, including moving time, was 40 and the combined face advance was approx. 17, 000 feet (5200 m) with face lengths varying from 450 - 550 feet (137 - 168 m). While the shield support experience in the U.S. is still limited on the basis of time and mining parameters, the operating results of all faces have been very satisfactory. New record 24 hr production figures were set and - far more important - long term production figures could be sustained on an impressively high level even when taking face moving time into account. The purpose of this paper is to review the performance of shield type supports in the U.S. and to assess their value as longwall supports for U.S. underground mining conditions with regard to roof control, maintenance, operator convenience, safety and productivity.

Jan 1, 1976

By M. Tanaji

Over the last many decades, a significant amount of research has gone into the domain of slurry transport. However, design engineers still face many challenges with respect to prediction of pressure drop, critical velocity and other design parameters as a function of Solids % and Particle Size Distribution (PSD). The industry requirement is to transfer the slurry at the maximum concentration as possible (above 30% (volume %)) to make slurry transport more economically viable and to reduce water consumption. To facilitate the design and scale-up of slurry transport in pipelines and in process plants, there is a need for a correlation that can predict slurry pressure drops over a wide range of operating conditions and physical properties of different slurries. The objective of this study is to overcome the limited range of applicability and validity of existing correlations and to develop a generalized but more rigorous correlation applicable to a wider range of slurry sys-tems. The existing Wasp et al. (1977) method is based on multi-phase flow modeling approach. This study attempts to modify this approach by considering material-specific values of Durand?s equation co-efficient and by defining flow regimes based on particle Reynolds number. When compared with experimental data, the modified Wasp method proposed in this study predicts the pressure drop for slurry flows more accurately than other available correlations. Also, the pro-posed method requires minimal test/experimental data for a particular slurry system and can be extended over different input conditions. An iterative computer algorithm is developed to calculate the criti-cal settling velocity and pressure drop in a pipe as a function of Solids % and PSD. The solution method can easily be implemented in de-signing slurry pipes, design validation, and studying the different slurry transport scenarios. The modified method can also be extended to accurately predict pressure drops in dynamic pressure flow networks used in commercial process simulators.

Feb 23, 2014