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By J. R. Kadambi, C. Shingote, J. Furlan, R. Ke

"Hydrocyclone separators are widely used in various industrial applications in the oil and mining industries to sort, classify and separate solid particles or liquid droplets within liquid suspensions, which are considered to be multi-phase systems. Numerous valuable studies have been conducted in recent years to investigate the flow fields inside hydrocyclones. However, much of the information regarding the performance of cyclones in the literature has limitations, based in some part on the respective current-available theoretical models, and much of it cannot be considered as completely applicable to most real-world applications; many of the studies investigated the flow fields within extremely simplified hydraulic designs that are not representative of the complex geometries or large sizes which are typical in industry. Therefore, in this study, the two phase flow system inside the actual hydraulic geometry of a milling circuit hydrocyclone was explored with the aid of both computational and experimental techniques (Particle Image Velocimetry). In this study, the flow field with an air core has been investigated; in essence, the research was a two-phase problem, which caused some challenges on both the computational and experimental sides. Two turbulence models were used in the numerical calculations: the Reynolds Stress Model and Large Eddy Simulation. The computational studies are mainly focused on the prediction of the dimensions and shape of the air core. Different section planes of the hydrocyclone were selected as planes of interest and then were divided into several fields of view (FOV) for PIV measurements. Two dimensional experimental velocity vector maps were obtained in each of the fields of view. The computational results were validated globally using pressure and flow rate readings at the boundaries and locally by comparison to the PIV velocity vector maps and profiles. Keywords: Computational fluid dynamics (CFD); Hydrocyclone; Particle image velocimetry (PIV); Reynolds Stress Model (RSM); Large Eddy Simulation (LES). INTRODUCTION A hydrocyclone is a device used to separate a solid-liquid, liquid-liquid, solid-gas, or liquid-gas suspensions into at least two separate density or particle size components. Hence, the flow field of the hydrocyclone is generally considered a multi-phases flow system. Application of hydrocyclones have been used in numerous industries including oil, drilling, metal working, and mineral processing since 1891, when E. Bretney obtained the first patent for the predecessor design of all contemporary hydrocyclones. A modern hydrocyclone typically consists a cylindrical section at the top where the liquid is being fed tangentially, and a conical chamber attached to the bottom of the cylindrical section, as shown in Figure 1. The characteristics of the feed stream that is injected into the cylindrical chamber play a significant role in hydrocyclone operation. The inlet is tangential relative to the cylindrical chamber and hence causes a centripetal force in the flow field which is considered as the primary separation mechanism."

Jan 1, 2017

By Alvin W. Schillinger

Drilling was not entirely satisfactory in the search for native copper in the Osceola amygdaloid footwall as mining showed oreshoots to be more continuous than drilling indicated. The Geophysics Department of Michigan Technological University developed a portable induced polarization drill hole probe with a 3 electrode configuration and a sampling radius of 5'± A 20 ma current of alternately reversed polarity is pulsed at 3.5 sec. intervals and readings are taken every two feet. Potential is read during the "on" cycle for resistivity determinations. During the "off" cycle, after 10 ms delay, the I.P. potential is sampled for 10 ms and recorded. Resistivity (ohm-feet) and "S" values (mv/v) are plotted on semi-log paper. A lithologic log is plotted based on resistivity and anomalous "S" values noted. The low resistivity amygdaloid contains sills of high resistivity basaltic trap. Verified by core logs, amygdaloid-trap contacts could be picked to ±.5' from resistivity alone. Lithologic logs of long steel holes, based' on resistivity, now replaced sludge logging. An "S" value of 30 mv/v was established as the lower limit of diagnostic I.P. effects in mineralized amygdaloid and 50 mv/v as a threshold value for ore. Of 914 footwall zones drilled, 315 were probed yielding 203 anomalies. Success increased from 34% for drilling to 72% for drilling and probing. As checked by mining, 46 probe results were 78% correct. Failures were due to the instrument not detecting copper or to the misinterpretation of an anomaly. No false anomalies were found.

Jan 1, 1964

By W. R. Reed, S. Klima, J. Driscoll, T. W. Beck

"Tests were conducted to determine properties of 4 foam agents for their potential use in longwall mining dust control. Foam has been tried in underground mining in the past for dust control and is currently being reconsidered for use in underground coal longwall operations in order to help those operations comply with MSHA’s lower coal mine respirable dust standard (1.5 mg/m3). Foams were generated using two different methods. One method used compressed air and water pressure to generate foam, while the other method used a low pressure air generated by a blower and water pressure using a NIOSH developed foam generator. Foam property tests comprising of a bottle shake test, foam expansion ratio, and water drainage were used to classify foams. The bottle shake tests determined that the lowest concentration of foam agent to be used was approximately 1.5%. Compressed air generated foams tended to have low expansion ratios (10-19) with high water drainage. Blower air generated foams had higher foam expansion ratios (30-60) with lower water drainage. Foams produced within these ranges of expansion ratios are stable and potentially suitable for dust control. The different foam agents produce foams with different material properties. Results of testing eliminated two foam agents for future testing because they had poor expansion ratios. The remaining two foam agents, while having different foam properties, seem to have properties adequate for dust control. These material property tests can be used to classify foams for their potential use in longwall mining dust control. INTRODUCTION About one-half of U.S. underground coal is produced by longwall mining. This technique allows for mining high volumes of coal, producing significant amounts of coal mine dust. This can lead to the overexposures to longwall miners and possibly occupational respiratory diseases black lung and silicosis, which have no cure and can be disabling or fatal. The only method to avoid these occupational illnesses is through elimination of exposure to respirable coal mine dust and crystalline silica (quartz). The current occupational exposure limit for respirable coal mine dust is 1.5 mg/m3 during each shift that a miner is exposed in the active workings of the mine or in mine facilities [1]. When respirable quartz is present, the mine must maintain an average concentration at or below 0.1 mg/m3. If the mine exceeds the 0.1 mg/m3 respirable quartz dust concentration, then the applicable respirable dust standard is reduced, calculated as 10 divided by the percent quartz present [2]. Respirable dust samples collected by U.S. Mine Safety and Health Administration inspectors for the five-year period from 2010 to 2014 found that 5.6% (53 of 943 respirable coal mine dust samples) of the longwall operators on the tailgate side exceeded the coal mine dust standard, and 4.3% (76 of 1768 samples) of the jacksetters exceeded the standard. If silica is present, the reduced standard is implemented, meaning that 11.4% (36 of 316 respirable coal mine samples analyzed for dust) of the jacksetters exceeded the reduced standard. Only 3.8% (18 of 478 samples) of the longwall operators on the tailgate side exceeded the reduced standard [3]. Had the 1.5 mg/m3 standard been in effect during this period, non-compliant samples would have been much higher. For example, over 15% of the samples collected at the tailgate shearer operator occupation exceeded the lower standard."

Jan 1, 2017

By H. W. Hitzrot

Introduction Mechanical classifiers consist of a settling tank with parallel sides and a sloping bottom equipped with a mechanism which continuously agitates the pulp and removes the settled solids. The functions served by mechanical classifiers include the following: (1) allow the particles larger than the desired size to settle in the tank and provide an overflow product with a minimum of oversize particles, (2) produce an overflow product of sufficiently high solids content to meet the requirements of the subsequent processing steps, (3) agitate the pulp and permit separation of the entrapped undersize particles so they can report in the overflow product, and (4) drain and remove the underflow solids from the settling pool. Various types of mechanisms have been used to agitate the pulp and remove the settled solids, which include: (I) endless belt and chain drags, (2) reciprocating rakes, (3) helical flights mounted on a rotating shaft which extends through the settling pool, (4) helical flights inside of a rotating drum which contains the classifier pool, (5) troughed conveyor belts forming the settling pool in the concave curve of the belt closely following the tail pulley, and (6) low-speed impellers agitating the underflow settled in the apex of cone-type mechanical classifiers.23 The principal use of mechanical classifiers has been in closed¬-circuit wet grinding; however, this application has been widely displaced over the past few decades by the hydrocyclone. The hydrocyclone installation, including its feed pump, offers lower capital costs and requires less floor space. Nominally it produces an overflow prod¬uct of higher solids content, compared with a mechanical classifier, for comparable classification size up to 400µm (35 mesh). However, these advantages of the hydrocyclone are being gained at the sacrifice of some of the advantages of the mechanical classifier. The mechanical classifier requires less power and has lower maintenance costs than the hydrocyclone and its feed pump. The underflow product from the mechanical classifier can have a higher solids content and contain less entrapped undersize particles than the comparable product from the hydrocyclone. Consequently, it is capable of operating at a higher classification efficiency. A reduced circulating load24, 25 results under these conditions in a closed-circuit grinding application. Other applications for mechanical classifiers include desliming, dewatering, and washing operations. Sand-slime separation may be achieved using one or more stages of mechanical classifiers. Typical applications would include washing and deslimining of concrete aggre¬gate or sand; glass sand; abrasives; oyster shells; phosphate rock; iron, nickel, and chromium ores; alumina; zeolites; solar salt; and precipitates from chemical processing. As an example for the latter, Bitzer26 recommended the use of mechanical classifiers for countercur¬rent cementation-in-pulp of copper from the leach liquor. Similarly, mechanical classifiers are used in uranium processing27 for separation of the barren sands from the pregnant leach solution. Depending on the particle size and the effectiveness of the sand¬slime separation, the sand product from mechanical classifiers may be drained to solids content approaching 84%. Thus, mechanical classifiers are frequently used as dewatering equipment for various applications, such as: 1) Make a product having suitable solids content for subsequent transport by belt conveyor. 2) Provide a drained product to reduce cost of drying. 3) Provide a drained, washed, heavy sand product for return to the circuit. 4) Recover fine values from dilute plant effluents from coal, sand, phosphate-rock, and iron-ore washing plants and from steel-mill flue¬dust scrubbers. 5) Drain residual reagents of a preceding process stage prior to conditioning with reagents of different characteristics, e.g., cationic¬anionic flotation of phosphate rock, iron ore, etc. Design and Operating Features The proper selection of mechanical classification equipment re¬quires that the properties of the solids and liquid are adequately de¬fined. The following descriptive information regarding the solids is desired: feed rate, chemical and physical composition, density, tem¬perature, size analysis, and desired separation size. For the liquid used, the following data would be needed: feed rate, density, viscosity, pH, and temperature. Settling Area. The pool area required to permit a particle larger than the separation size to settle depends on the density and shape of the particle and on the density and viscosity of the pulp within the classifier pool. These criteria of the solids and pulp determine the settling rate which, in turn, determines the settling area required. Determination of the settling rate by long-tube batch tests can be used to calculate the settling area. A theoretical approach for estimat¬ing the settling area requirements is discussed in Chapter 1, Classifica¬tion Theory. This procedure can be used when neither test data nor empirical data from similar operations are available. Data on settling-area requirements and settling rates have been compiled by the manufacturers of mechanical classifiers. The approxi¬mate settling area required for a particular separation size can be estimated by using the curve shown in Fig. 65. This curve applies for closed-circuit grinding applications. For dilute velocity classification in open circuit, desliming, the length of the overflow weir, W, multiplied by the weir's mean distance from the feed opening, E, is used as the design criterion, instead of the settling area, in sizing the classifier. The area requirements, W X E, for dilute velocity classification can be expressed in terms of volumetric flow, Q, and settling rate, Rs, as follows: W x E = Q/(18.06 Rs) where W and E are feet, Q is gallons per minute, and Rs is inches per second. Overflow Weir. The hydraulic head, and thus the velocity at which the overflow crests the overflow weir, is one of the factors controlling the particle size of classification. Consequently, the length of the overflow weir must be designed to provide the overflow velocity which allows the particles larger than the desired size of classification to settle in the working area of the pool. Obtaining the necessary

Jan 1, 1985

By Victor F. Hollister

This first section of case histories of mineral discoveries, Vol. 3, covers some of the milestones in porphyry copper and skam exploration successes. It was during the era of these discoveries that large, complex exploration departments emerged as permanent adjuncts of large mining companies. The discoveries helped shape how and where exploration activity was conducted. Probably the most important factor in exploration success is management. Few management people are mentioned in this section or in any exploration history, yet the distinction between success and failure ultimately rested with the decisions they made. One of the most successful groups was initiated by Roger Straus and Herbert Brownell when they sought to make Asarco into an important metal producer. Relying on Clem Pollock and L.H. (Bill) Hart to identify and push worthwhile projects, the Asarco management remade the company. Asarco will benefit from exploration successes com- pounded by that team for years to come. Duval is an example of a management that decided to add metal production to its industrial minerals output. When Page Morris was president, that company opened three major porphyry copper mines and became an important factor in metals. Under the Monis management team, Duval became a leader in exploration. At the other end of the scale are managements that snatched defeat from the jaws of victory and by so doing changed the character of their companies in other ways.

Jan 1, 1991

By Lech Bolinski, Joe Shirley

Over two dozen Resin-in-Pulp (RIP) plants have been in operation in the fonner Soviet Union since the beginning of 1970. Currently approximately 50% of the gold produced in Russia, Uzbekistan and Kazakhstan, which equates roughly to 10% of the total world output, is recovered using the RIP technology. An alternative to activated carbon, ion-exchange resin has several advantages for the recovery of gold. Resins are less prone to clay or organic foul­ing and do not require thennal regeneration. They also have poten­tially higher loading capacities, faster adsorption kinetics, and are more resistant to abrasion. The original Russian RIP process was based on engineering expertise developed for uranium ore pro­cessing and has a few major drawbacks. The biggest problem is an 18-30 day resin stripping process, which results in a huge gold lock-up. The problem was recently overcome by the development of a new stripping technology which cuts the time to 12-24 hours. Currently, there is rather limited practical knowledge of the Russian RIP process in the West, but often Western mining com­panies developing gold mines in the fonner Soviet Union are hard pressed by their local joint venture partners to take a serious look into RIP. It is probably only a matter of time before one of these companies will cross Soviet technology with Western engineering and develop a modem RIP process as an attractive option to the current CIP and CIL technologies.

Jan 1, 1996

By Stephen P. Signer

The U.S. Bureau of Mines is developing an intelligent system for roof' control that uses both an expert system and neural networks to improve the capability of mining engineers to evaluate roof support effectiveness for ground control in coal mines. The expert system com- pares roof support capacities with the support requirements estimated to be necessary to maintain entry stability. It does this by evaluating the results of tests on various types of roof support and establishing the maximum allowable load according to anchorage capacity and yield strength of the support. After the user enters geological information (rock properties, geometry of the opening, in situ stresses, bolt pattern parameters, etc. ), the expert system compares the predicted required loading to the roof support capacity and gives the operator advice on the adequacy of the design and how improvements could be made. A good source of the real-time data necessary to allow the expert system to make decisions will come from a roof bolting machine being developed by the Bureau of Mines. Researchers have collected data from a Western coal mine on drill bit position, penetration rate thrust, torque, and rotation rate. Using this information, two neural net- works were developed to identify different types of strata and features in a mine roof such as rock type, rock compressive strength and joint characteristics. The result is a system that can assist a mining engineer with design information that can be constantly updated as mining progresses.

Jan 1, 1992

By Ekkehart Mertins, Hans-Dieter Wasmuth

During the last years many new concepts have been presented for magnetic separation of feebly magnetic minerals by using the new high-duty permanent magnetic material especially in Open Gradient Magnet Systems (OGMS). In this respect KHD Humboldt Wedag AG developed an improved version of a drum-type magnetic separator called PERMOS® with a special magnetic field configuration for the NdFeB permanent magnets, which gives a magnetic field strength of at least 0.7 Tesla and a wide working zone outside the drum. The magnet system within the PERMOS® unit is no longer made of relatively large magnet blocks of alternating radial magnetisation; the system is now made up of a multitude of small NdFeB bars with a specially oriented magnetisation. This PERMOS® Open Gradient Magnetic Separator can be used for dry magnetic separations in "top-feed" and 'free-fall" feed configuration as well as in the conventional wet "drum-type" feed mode. It allows high throughput rates of up to 40 metric tons of solids per hour and meter of drum length without any clogging problems, even when strongly magnetic com-pounds are present in the feed material. The important design features and the actual machine characteristics of the improved PERMOS® units for dry and wet magnetic separation are described and typical examples for their commercial application for industrial minerals and for martitic iron ores are presented.

Jan 1, 1995

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984

By Norbert Patzelt

Today the standard method for grinding ore is the wet process. This is mainly due to the fact that the downstream process is wet and considering the total circuit dry grinding is not the most eco­nomical method for the comminution of the ore. Presently in gold ore processing dry grinding is only applied in cases where whole ore roasting is required downstream from the milling plant or if organic carbonaceous matters and sulfides are the primary reasons for the refractoriness. Since dry ore can be fed to the roaster, a dry grinding mill can advantageously be applied. In cases where d1y grinding could be applied advantageously for instance if the ore contains water soluble material and wet grinding is not possible there are serious concerns prevailing within the industty against this technology due to several reasons. Besides of the aspects as discussed later in this paper, there are several plants in operation with little success and considerable operational prob­lems. The reasons for this often lie in the fact that the several com­ponents of such a plant are bought from independent sources with­out considering the system requirements. It is essential that the dry grinding plants require accurate design between drying, pneumatic transport and grinding of the ore in the circuit. This paper will address some of these concerns raised in the industiy from a mill vendor's view along with discussing the development for the selection and configuration of the various dty grinding circuits.

Jan 1, 1998

By Vernon R. Shorter, Wayne N. Clark

Introduction Coal is the US's largest indigenous fossil fuel resource and is used to power 57% of this country's electrical generation (NCA, 1986). Congress, however, is focusing its attention on the acid rain issue and public sentiment is leaning toward further curtailments of SO2, NO, and particulates from utility and industrial coal-burning facilities. The Cool Water plant has proven that integrated-gasification-combined-cycle (IGCC) technology can reliably produce economical electrical power while limiting SO2, NOx, and particulates to a rate of only one-tenth of the federal New Source Performance Standards for coal-fired plants. Process water requirements at Cool Water are 40% less than those of a conventional coal-fired utility plant equipped with flue-gas desulfurization units for stack gas scrubbing, thus reducing waste water treatment requirements (Clark and Shorter, 1985). The IGCC technology employed at Cool Water has now been demonstrated to efficiently convert coal to electricity in an environmentally superior fashion. The Cool Water Coal Gasification Program is an unincorporated California association of six partners whose commitment is proving that the integration of gasification with combined cycle power production (IGCC) is commercially viable and environmentally safe. The 120 MW Cool Water plant began commercial production on June 24, 1984. The Program is operating it for a five-year commercial demonstration period. The Program is about halfway through this period. Much of that which was intended to be proven is now fact. In this mid-term assessment, the authors report that the com-

Jan 2, 1987

By K. H. McDaniel, L. Griswold

The Waste Isolation Pilot Plant (WIPP) is adding a third fan to supplement its two existing main fans. The third fan will operate in parallel with the two existing fans. An engineering analysis conducted in 1995 identified the need for improved system availability and efficiency. The WIPP, located near Carlsbad, NM, is a US Department of Energy (DOE), Carlsbad Area Office, facility that is designed to permanently dispose of US defense-generated transuranic waste in a deep, bedded salt deposit. Engineers at the WIPP performed a feasibility analysis for modifying, supplementing or replacing its two main mine fans. Since the centrifugal fans were installed in 1988, multiple operational and performance concerns have been identified. A comprehensive engineering study was conducted (McDaniel et al., 1997) to identify and quantify operational concerns, evaluate possible alternatives and recommend an optimum solution. The study indicated that adding a third single-duty centrifugal fan similar to the two existing fans and equipping that fan with Inlet Van Controls (IVCs) was the best choice. This alternative provided the most desirable combination of operational readiness and improving system performance. The fan installation project accommodates several operational and scheduling con¬straints. WIPP was scheduled to receive waste in May 1998. This milestone is proceeded by an operational readiness review. Fan installation could not be allowed to compromise the WIPP 's ability to declare facility readiness. Therefore, operational disruptions during the system tie-in had to be minimized. All engineering for the project has been com¬pleted, and physical modifications were begun in July 1997. The project was completed in 1998.

Jan 1, 2000

By Suresh P. Babu, Joseph W. Leonard, Richard Muter, D. P. Sen

OUTLOOK FOR COAL UTILIZATION In 1975, about 650 million tons of bituminous coal and lignite mined in the United States was used to meet 55% of the primary energy input for electric power generation and 26% of the total industrial consumption.1 Even though this amounts to only 21% of the total energy requirements, it is believed that with the declining natural gas resources and uncertainty of liquid fuel imports, coal will assume a more dominant role in suppplying our future energy needs. A regional breakdown of the coal reserves of the United States along with their characteristics is shown in Table I. At present the United Stares Energy Research and Development Administration (ERDA) and the US Bureau of Mines (USBM) are financing extensive research programs in advanced mining methods and energy conversion processes. The increased demand for wal which may result from these research programs may also create large-scale environmental protection problems. As more and more coal is mined the production of undesirable coal fines and dust also increases proportionally. The USBM and the Environmental Protection Agency (EPA) research in this field includes development in the areas of instrumentation for qualitative and quantitative analysis of wal dust, protective devices, and dust collectors. Dust control by water sprays, steam, and foam are also being investigated. Other aspects of safety in mining include developing methane drainage techniques prior to coal mining, fire and explosion prevention, and neutralization and disposal of acid mine waters. It is the belief of coal technologists that when the emphasis shifts to coal as the primary source of energy. the existing methods of mining and preparation of coal must be improved considerably to supply large quantities of quality coal for energy conversion. It is the objective of the present large-scale research financed by ERDA to produce substitute fuels from coal. Large-scale gasification pilot plants are being constructed by Bituminous Coal Research, Consolidation Coal Co., Institute of Gas Technology, and the Pittsburgh Energy Research Center to produce coal gas of natural gas standards at 1,000 psig. The processes being adopted by these pilot plants are oriented towards recovering most of the sulfur from coal by absorbing H2S and its subsequent conversion to elemental sulfur. The environmental protection standards will be strictly followed in using the results of these pilot plants to construct large-scale coal gasification plants in the latter part of the 1970's. With the increasing awareness of the need for air pollution control, considerable research is being done to minimize sulfur dioxide emission from thermal power plants. Utilization of limestone and dolomite scrubbing systems and caustic treatment of stack gases have proved to reduce sulfur and nitric oxide emissions by about 90 to 96%.3 Currently, about 22 scrubbers are in use in the United States to reduce sulfur in stack gas emission. In the long-term, it appears that optimal combinations will be developed involving both advanced coal preparation technology and wet scrubbing. The sulfur compounds are converted either to elemental sulfur of sulfuric acid. Besides the sulfur compounds, the major byproduct of the coal utilization industry is the large amount of coal ash. The Coal Research Bureau at West Virginia University has developed a practical use for coal ash as a structural material. The coal industry will undergo revolutionary changes when coal becomes the primary source of energy. New coal mining and preparation processes will be introduced to meet the ever-increasing per capita energy requirements. 2. BENEFICATION Introduction Coal preparation, like the beneficiation of ores, is also aimed at the separation of values from associated waste. Even though the unit operations involved when enriching ores and coal are similar, there are certain characteristics unique to coal preparation. Unlike the scattered and concentrated veins of underground minera1 ores, coal is distributed uniformly and very much in abundance in the coal mines. Coal occurs as the metamorphosed sedimentary deposits of vegetable matter, while metal and mineral deposits occur in great varieties of forms, structures, and origin-from igneous to sedimentary and metamorphic. The mined coal, being much more friable than mineral ores, is subjected to less intensive size reduction operations. Ores are usually ground to a very fine size for subsequent concentration, while coal is crushed to a coarser size, since considerably less liberation is needed to achieve the desired degree of separation. Since mineral matter is to a great extent discretely distributed in coal, concentration is mainly achieved by taking advantage of differences in specific gravity between clean coal and associated waste. In general, about 90% of coal is concentrated using specific gravity separation and about 10% by froth flotation. In contrast, these operations are frequently reversed in the beneficiation of ores. In coal preparation almost all fractions separated during various stages are made use of. while large amounts of gangue material are rejected in the concentration of ores. Hence, coal generally has a much lower ratio of concentration than ores. Since ore deposits are frequently more variable and mineralogically complex than coal, beneficiation of ore may require more plant flexibility and point-to-point quality control. This characteristic difference between the mineral ores and coal is attributed to the generally used one-step processing in coal as compared to the multiple-step processing of ores. Dewatering to enhance the energy content of coal is usually considered a major unit operation. Even though thickening and filtration are common for dewatering ore concentrates and cleaned coal, vibrating centrifuges and disk filters find unique application in the coal industry for achieving these types of solid-fluid separation. In coal preparation, thermal drying is carried out in separate units at the mine. While ore concentrates are usually dried at the smelter. Although the drying of ores is seldom considered dangerous, volatile matter and combustible dust are reasons why the thermal drying of coal can be hazardous. Open air storage of ores and mined coal results in partial oxidation. Unlike the oxidation of ores, the oxidation of the pyritic and carbonaceous constituents of coal together with the heat of condensation of moisture contribute to spontaneous heating and the subsequent combustion of coal in storage. Among the other significant difference between concentration of ores and coal are the weight and capacity of processing equipment. Coal processing employs less cumbersome, higher capacity units while the beneficiation of ores usually requires much more massive and frequently lower capacity equipment. Properties of Coal Which Influence Design The widely varying and heterogeneous composition of coal, due to metamorphism, determines the preparation design and uses of the resulting products. Petrography serves as a useful system of classification to identify the coal types present and their distribution. Early

Jan 1, 1985

By E. D. Attanasi, M. W. Bultman, J. H. DeYoung

The 1987 recovery in the US mining industry permitted modest increases in exploration expenditures that year (Fig. 1). Favorable market conditions apparently allowed the gains to be sus¬tained in 1988. There was great uncer¬tainty early in the year as to whether world and US economic growth would be a casualty of the October 1987 worldwide stock market crash. By nearly all measures, the performance of the US minerals industry was strong¬. A weak US dollar and the strong growth of worldwide capital spending resulted in in¬creased US exports of con¬sumer and producer durable items as well as strong domes¬tic demand for metals. In fact, 1988 US steel industry exports reached levels not attained since the mid-1970s. It was re¬ported that steel industry labor productivity is now roughly twice the 1981 level. Among metals with signifi¬cant average nominal price increases in 1988 over 1987 levels were copper (43%), alu¬minum (52%0), and zinc (44%). Even though domestic copper production has grown since 1984 by 30% in mining and 24% in refining, em¬ployment in mining and milling is 4% below the 1984 level. Domestic unit costs are reported in the range of $1.10 to $1.33/kg (50 to 60 cents per lb). The US base metals industry is proba¬bly in its best cash position in many years. However, there has not been a large increase in exploration expendi¬tures. This is due mostly to the percep¬tion that worldwide capacity can easily be increased. Industry reports do show that "low cost" copper producers have been buying reserves from weaker com¬panies. The 1987 Society of Economic Ge¬ologists (SEG) survey of exploration activity indicated that 95 cents out of every exploration dollar spent by US and Canadian firms was for base and precious metals (Economic Geology, January 1989). Press reports, though, make it clear that most of that spending was targeted for precious metals. US claim staking data (Fig. 2) rein¬forces this suspicion. The apparent lag in new claim filings in Alaska and phe¬nomenal growth in filings in the lower 48 states suggests that exploration is directed at relatively low risk targets. In the conterminous 48, more than 40% of the new claims were made in Nevada, indicating an emphasis on gold (although even Nevada had some base metals exploration). Even with a price decline during 1988, the difference between the gold price and average global "cash" production cost of $7.30/g ($227 per oz) still leaves room for an attractive return (Northern Miner, 1988). Exploration plans and results Metals Economics Group (Halifax) indicated in its 1988 exploration survey that exploration expenditures by US and South African firms would be at about 1987 levels. It also reported that Canadian and Australian firms were ex¬pected to reduce their exploration spend¬ing significantly. Firms in both coun¬tries were caught in changes in tax laws. In Canada, the Mining Ex¬ploration Depletion Allow¬ance (MEDA) once permitted generous tax write-offs to in¬dividuals investing in miner¬als exploration. It was to be replaced in January 1989 with the Canadian Exploration In¬centives Program (CEIP). Although structured and ad¬ministered differently than MEDA, CEIP can provide an equivalent or possibly greater financial incentive for min¬eral exploration, for junior mining companies. In 1987, it is estimated that junior com¬panies accounted for two¬-thirds of the off-property (or grass roots) exploration ex¬penditures in Canada. Australia will revoke the tax exemption for gold in 1990. Australian operators, therfore, were expected to concentrate efforts on speeding up pro¬duction during 1988 and 1989. In 1988, Chile and Mali announced special efforts to attract gold explora¬tion by foreign companies. The Cana¬dian government continues to express concern about the write-off (and drawdown) of base metal reserves. Falconbridge did announce a new initiative to identify additional base metal reserves in the Sudbury district. Discoveries Gold deposits dominated discoveries

Jan 1, 1989

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By Robert W. Lawson

INTRODUCTION The two basic types of mine-air compressors are the positive-displacement compressors and the dynamic compressors. Positive-displacement compressors con¬fine successive volumes of air within a closed space and increase the pressure by reducing the volume of that space. Reciprocating and rotary compressors operate on this principle. Dynamic compressors utilize an impeller to acceler¬ate the air, with the increased air velocity converted to pressure through a stationary diffuser. Centrifugal compressors and turbocompressors operate on this principle. Reciprocating Compressors Reciprocating compressors include both single-stage and multistage units. For mine service the most com¬mon is the two-stage intercooled compressor having an atmospheric pressure intake and a discharge pressure of 689 to 861 kPa (100 to 125 psi). The cylinders of a reciprocating compressor may be either single acting or double acting. In the single-acting cylinder, compression occurs on only one side of the piston, while, in double¬acting cylinders, compression occurs on both sides of the piston. Fig. 1 shows a cross-sectional view of a typical two-stage double-acting water-cooled reciprocat¬ing compressor. Fig. 2 illustrates the sequence of opera¬tions occurring in the cylinder of a reciprocating com¬pressor. Most modern mine compressors are double acting and are driven by electric motors. Induction motors are used on low-powered compressors, and synchronous motors are used on larger compressors. Steam-driven compressors once were common and still are applied where a reliable source of low-cost steam is available. For this type of compressor, steam is a desirable power source since a governor that senses the discharge pres¬sure can be used to control the steam inlet. This can be used to regulate the compressor speed to compress only air sufficient to meet the mine's requirements (up to the capacity of the unit). Unlike a constant-speed electric¬driven compressor, no other regulation is required for the cylinders of a steam-driven unit. Rotary Compressors Rotary compressors include sliding-vane, liquid¬piston, two-impeller straight-lobe or cycloidal, and helical-lobe units. Each of these has specific advantages and applications. Of these, the helical-lobe or "screw" compressor has the greatest application for supplying "100-lb air" [689 kPa (100 psi) ] to small- and medium¬size underground mines. Fig. 3 illustrates the operation of a sliding-vane compressor. Fig. 4 illustrates the operation of a helical lobe or screw compressor. Fig. 5 illustrates the operation of a two-impeller straight-lobe compressor. Fig. 6 illustrates the operation of a liquid¬piston compressor. Centrifugal Compressors Centrifugal compressors deliver 0.94 m3/s (2000 cfm) or more of air, and they are best suited to supply¬ing large volumes of compressed air for base loads. The efficiency of a centrifugal compressor declines rapidly when its operation deviates from its design point. If the power cost is not a critical consideration, a centrif¬ugal compressor is attractive as a result of the low initial and installation costs and the high degree of reliability. Fig. 7 illustrates a typical four-stage centrif¬ugal compressor, and Fig. 8 is a cross section of this type of unit. COMPRESSOR SELECTION The process of selecting a compressor requires sev¬eral considerations, including the capacity, the control system, the location, the power supply, and the cooling provisions. Compressor Capacity To determine the required capacity, the total air quantity and pressure requirements must be calculated. The air consumption of various tools, as listed in Table 1, must be included in this calculation. Some of the other considerations include: 1) The correction factor for altitude, as listed in Table 2, must be included in the calculation. While pneumatic tools require additional compressed air as the altitude increases, the compressor produces less. 2) A correction factor must be incorporated for the use of multiple tools. Table 3 lists appropriate correc¬tion factors for rock drills.

Jan 1, 1982

By J. Tuzson, H. McL. Clark

Introduction Wear of pumps, wet cyclones and pipelines by suspended solid particles will impact their operating economies by limiting their useful life, reducing their reliability and increasing the amount of servicing. Erosion is considered a materials problem. The empirical solution consists of testing different materials in field applications. Such tests are time consuming and expensive. The prediction of localized wear rates, life estimates and appropriate material selection demand a better understanding of the slurry wear process in environments corresponding to those found in working equipment. Key words: Wear, Slurry abrasion, Coriolis erosion tester The slurry wear process Solid particles suspended in liquid generally move with the flow, but such particles deviate under the centrifugal and Coriolis forces encountered in pumps. The concentration of particles will increase, and a solid layer or bed may accumulate if the acceleration is directed against the flow passage wall. Large particles in dilute suspensions may impact the wall. dissipating their kinetic energies and removing material. However, particles smaller than about 11)(1 µm encounter great drag when moving through the liquid. and such par¬ticles have little kinetic energy left when they reach the wall. They cannot penetrate the stagnant boundary layers present in pumps. Therefore, small particles tend to form beds that slide along the wall, removing material by a process that could he called sliding erosion or wet abrasion. Wear by impact erosion from large particles in dilute slurries can he predicted by appropriate analytical and empirical methods. However, laboratory tests and calculations of wear rates that correspond to sliding erosion are more likely to correlate with field experience from wear under similar conditions. Specific energy Data from laboratory slurry erosion experiments have shown that the amount of material removed is approximately in proportion to the energy expended in the erosion process. Thus. the specific energy-energy expended in removing by erosion a unit volume of material-offers a simple, practical measure for the erosion resistance of materials. The concept was found to also apply to sliding erosion (Clark et al.. 1997). Generally, the study of the material-removal process in slurry erosion continues to remain a subject of research. Conceptually, if the flow velocities and accelerations could be calculated in a pump or pipeline, the energy dissipated at the passage walls could also be estimated, and the material removal rate would become known from the specific energy determined by laboratory tests. Slurry erosion testing Laboratory devices made specifically for slurry erosion testing may be grouped into the following: jet erosion devices, slurry pots and the Coriolis test machine (Pagalthivarthi and Helmly, 1992). Jet erosion devices use a slurry pump and nozzle to produce a high-velocity slurry jet that is directed at a flat material specimen, either perpendicularly or at some acute angle. Combined impact and sliding erosion are produced. but the processes cannot be separated, and the wear is difficult to assess except by weighing. Slurry pots consist of a cylindrical vessel containing the slurry of interest. In the slurry pot, test specimens are attached to a rotating arm. The samples are rotated in the slurry and experience local flow conditions that result in combined impact and sliding erosion. Wear depth measurements from detailed laboratory experiments have been successfully correlated with results from an analytical model of the process, and specific energy values have been calculated (Wong and Clark, 1995). In industrial use, the samples are weighed to determine the total material loss. Materials are ranked with respect to some reference material. The Coriolis erosion tester (Tuzson, 1984) consists of a rotor with a central cavity, from which two opposing radial passages lead out (Fig. 1). The slurry is fed to the centerof the rotor in one pass from a stirred overhead tank. The slurry particles are subjected to the Coriolis acceleration as they flow out through the radial passages. The material specimens are mounted into the radial passage walls facing the direction of rotation and are eroded by the particles that are pressed against them as they flow out radially. A narrow groove is worn into the sample within a few minutes, using a few gallons of slurry. Material removal is estimated accurately from profilometer traverses across the wear scar at different locations. The energy dissipated by the particles can be estimated, which leads to a well-defined experimental value for the specific energy.

Jan 1, 2000

By Steve Kral

In December 2013, the Colorado Department of Transportation (CDOT) completed the widening of the eastbound tunnel on Interstate 70 (I-70) from two lanes to three. Work on the westbound tunnel began earlier this year and is scheduled for completion in December. The Twin Tunnels project is located in the mountains about 55 km (35 miles) west of Denver. They were bored during the 1960s when I-70 was being built and became the main access to Colorado?s growing number of ski areas and the state?s Western Slope. Throughout the decades following their construction, traffic on I-70 continued to increase and now the tunnels have a major bottleneck. Com-mutes to and from the mountain resorts, particularly on weekends, have become long and slow. Adding a third lane through the tunnels will help alleviate the congestion. (A complete report on the project can be found in T&UC, December 2013, page 61.) Once the eastbound tunnel was completed late last year, CDOT authorized the funding to begin work on the westbound side. That made economic sense, as the equipment used on the eastbound tunnel needed to only be shifted across I-70 to begin work on the westbound side, rather than removing it in December and bringing it back in March when construction began on the westbound tunnel. Work on the westbound tunnel uses essentially the methods that were used on the eastbound tunnel ? removal of the existing concrete lining and using drill-and-blast methods for tunneling widening. The main difference between the two tunnels is that the westbound side requires more blasting on both ends to widen I-70 to three lanes.

By Norbert Patzelt, Johann Knecht

Future ore reserves of major gold producers in North America and elsewhere will consist of refractory gold ores. About 60 % of the current U.S. gold production is mined in Nevada. Several of the large gold operations in this state indicate that 40 to 90 % of their future ore feeds will contain refractory gold. To com­pound problems, average head grades for refractory U.S. gold ores had fallen from 8 - 9 g/mt to about 4 g/mt by 1991. In Latin America, 52 % of the gold operations will require milling for optimal recovery. Deeper reserves are, like in North America, refractory. Therefore a special treatment process is needed in order to allow for the gold ore to be leached downstream in a conven­tional cyanidation process for a high gold recovery. There are several options available for this oxidation process like treat­ment of the gold bearing ore in pressure oxidation, in a biooxi­dation or in a roasting plant depending on the reason for the refractoriness. If organic carbonaceous matters and sulfides are the prima­ry reasons for the refractoriness, whole ore roasting is a pre­ferred method for the treatment of this type of ore. Since dry ore is to feed to the roaster, a dry grinding mill is advantageous. For the dry grinding system, the moisture content and the ore char­acteristic is of importance. Those parameters decide the mode of the grinding system. For autoclave systems common wet grind­ing systems can be applied. This paper deals with technical aspects for dry grinding plants and gives an outlook on and pos­sibilities for new grinding systems.

Jan 1, 1996

By G. W. Luxbacher

Coal mine infrastructure is rarely designed solely on the basis of ventilating the underground workings. Instead the optimization of a mining sequence, the minimization of coal quality variation, reduction of coal losses by adapting to property boundaries and geologic structure (seam inclination, roof conditions, coal thickness, etc.) or other similar considerations have formed the objective for mine plant design and layout. Infrastructure development may also reflect a desire to maximize production early in the life of the mine or to permit flexibility in mining according to market conditions. Requirements for coal mine ventilation, established by the 1969 Coal Mine Health and Safety Act, have mandated changes in the traditional concepts of mine plant design in the United States. Mines, in operation prior to the passage of the Act have experienced severe limitations when attempting to adapt the existing mine infrastructure to meet the requirements of the law. In addition the increased quantities of air required, velocity limitations on trolley haulage and belt lines, the establishment of separate intake escapeways and gob area ventilation have com- pounded the inherent complexity in planning mine ventilation systems. Past practices in mine ventilation planning and design are no longer adequate. Given the complexity of today's ventilation systems, effective planning and design approaches are required. As such, the interrelationship between the overall coal mine plant design and ventilation system design is examined. An approach to ventilation system design constrained by global considerations of mine plant design, is presented. The proposed approach is, however, applicable to the design of new mine ventilation systems as well as to the modification of existing systems.

Jan 1, 1980