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By James A. Decker

Hanna has developed three production planning systems to help the mine engineers determine the sequence of mining from available ore supples. The Production Scheduling System was developed to give the staff a tool for long range planning by sequencing the removal of blocks from an open pit property. It was needed to determine the year of removal for each block so that the optimum present-valued pit limits could be established by our pit design program. Since in this type of planning short term grade control is less of a factor than it is in daily or monthly planning, the system gives the program greater control and more latitude in block selection. It permits only full blocks to be taken.

Jan 1, 1984

By Gordon T. Clark, Stephen H. Gleaton

This paper discusses the conceptual design of cut-and-cover tunnels currently being considered as one of the replacement options for the earthquake damaged, elevated double-deck highway and seawall built in downtown Seattle 50 years ago. The 4.8 km (3 mile) long, 6-lane improvement consists of 1.6 km (1 mile) of at-grade roadway with aerial interchanges, 1.6 km (1 mile) of large two-level cut-and-cover tunnels along the Elliott Bay waterfront, and retrofit of the 50-year old, 1.6 km (1 mile)long Battery Street tunnel and north transition to grade. Construction of deep cut-and-cover tunnels, in complex glacial geology overlain by hydraulically placed fill immediately adjacent the existing sea wall, will require extensive use of concrete diaphragm walls, ground improvement and dewatering. The potential liquefaction and lateral spreading of the fill materials in this active seismic region pose significant design challenges and risk.

Jan 1, 2003

By Z. Rogovin

A Knowledge of the amount of time the material spends in a continuous mill in a closed-grinding-circuit is essential for circuit performance analysis and scale-up design. A convenient method for acquiring this information involves the use of tracers in residence-time-distribution (RTD) experiments. In a typical wet process the RTD measurements are usually done by injecting a pulse tracer into the mill inlet and detecting it in the circuit exit. In this article a simple to use mathematical method for the calculation of the closed circuit material residence time in the mill is developed. Using this method there are no restrictions on the location of the tracer injection or detection points. Any convenient location in the closed circuit can be chosen without affecting the simplicity of the method. This method also enables the calculation of the degree of mixing in the mill. The use of this method for the analysis of impulse response data obtained from several commercial grinding circuits is presented.

Jan 1, 1986

By Benno Ring

"When designing segmental lining at a cross passage, modern solutions try to bear the loads transferred around the opening within the rings by using shear elements in the ring joints. Commonly used shear systems providing coupling are cam & pocket, shear keys, bicones and steel dowels. These solutions have advantages and disadvantages, but most importantly they have limits in their capacity to stabilize the lining at cross passages. With the help of 3D-FE-analyses for an opening of one segmental ring these different solutions are compared taking various ground conditions into account. As a result the capacity of the individual system is determined. Notice is made regarding the applicability of the system as well as the consequences for the design and the verification. INTRODUCTION The design of the lining at a cross passage is among the most difficult tasks when designing TBM-tunnels. At one time the bedding of the lining disappears due to the excavation behind the lining, at another the ring-structure is interrupted when parts of the lining are removed for the opening. The trick is to transfer the loads around the opening into other segments and to areas of the lining which are still supported by the ground. In the rings immediately at the cross passage this creates a distribution of forces within the segments that strongly deviates from that of the regular lining. The design needs to consider the boundary conditions of the respectively chosen solution for stiffening of the opening and its peculiarities in the modeling as well as in the verifications. In the past segments of steel or an internal strutting of steel (DAUB 2013) placed around the later openings were commonly used, see Figure 1. The application of a strutting from the outside to the rings is restricted to such circumstances when for example a connection from a shaft to the tunnel can be made in suitable ground conditions (Ring and Glomp 2012). Though steel segments provide a robust solution, they are expensive due to their laborious manufacture, the high proportion of construction steel required and the time consuming installation in the tunnel (Häfliger 2013). A strutting of steel requires extensive design and verification. The resulting structure usually weighs several tons and must be erected as well as demounted within the confined space of the tunnel, requiring suitable transport and assembly equipment. The cross section remains restricted, thus causing a bottleneck for TBM operations and hindering the works and installations for the cross passage works."

Jan 1, 2016

By John S. Thayer, Frederick E. Brinckman, Kenneth L. Jewett, Gregory J. Olson

Interest and research in biorecovery of metals has focused mainly on cell-based metal leaching and precipitation. However, production of non-enzymatic metabolic products by microorganisms related to metals bioprocessing also deserves attention. For example, it is well known that biogenic inorganic (e. g. H2S04 ) and organic acids play a role in metal dissolution. There are other metabolic products of microorganisms which can catalyze useful metals transformations. These include methylated metabolites such as alkyl halides which can dissolve metals from ores and also reduce certain metal ions in solution to pure metallic particles. This paper describes some of our research on metal dissolution and precipitation by such metabolites. Examples include the production of high purity elemental metal particles of gold, platinum and palladium by volatile metabolic products.

Jan 1, 1989

By Paul S. Murphy, Eric W. Oehrke

INTRODUCTION The global market for sulphur has experienced dramatic structural shifts during the past 10-15 years. After a number of years of dominance as sulphur suppliers, the world's Frasch producers were confronted with large increases in nondiscretionary sulphur production from sour crude oil and gas, while escalating energy and exploration costs reduced the economic incentives to replace dwindling sulphur reserves. Sulphur demand also experienced substantial growth, but the majority of this increase was filled by nondiscretionary production. The result was essentially flat demand for Frasch sulphur during this period. This paper briefly summarizes Freeport Sulphur Company's outlook for global supply and demand for sulphur through the year 2000, and examines in more detail the history and anticipated shifts in trade patterns. World trade in elemental sulphur (brimstone), despite year-to-year fluctuations, has been on an upward trend in recent years, and should continue to grow through the end of this century. The type (discretionary vs. nondiscretionary) and location of forecast sulphur supply sources has potentially broad implications for the world's Frasch producers, as they compete in a world market which is increasingly supplied by non-discretionary production. SULPHUR DEMAND Sulphur is a versatile and essential element used in a wide variety of manufacturing processes. Almost 90% of world sulphur consumption is in the form of sulphuric acid. The relatively low cost, ready availability, and physical and chemical properties of sulphuric acid make it the most economical acid for a number of chemical processes. Although sulphur is used in the manufacture of many products, the largest consumer by far is the phosphate fertilizer industry. The majority of sulphur demand growth over the past 30 years has occurred in this area, and today it accounts for nearly 65% of global sulphur consumption in all forms. Due to the relative importance of phosphate fertilizers, sulphur demand is typically discussed in terms of fertilizer and non-fertilizer (or industrial) uses, and this approach will be followed here. We at Freeport believe that global consumption of phosphate fertilizers should grow at an average annual rate of about 3.5% through the year 2000, which is roughly equal to the growth rate experienced during the past ten years. The basic assumption underlying this forecast is that world population will continue to grow significantly, while the total acres of cropland will grow only slightly. Therefore, in order to feed a growing population, higher crop yields will be necessary, which in turn will require the increasing use of fertilizers. This expected growth in phosphate consumption will, without doubt, fluctuate from year to year, and we Frasch producers have painful memories of the large dip in demand during 1982-83. However, we believe that the fundamentals for growth are very good, and that future demand will move back onto the historical trend line. In the recent recession the use of sulphur in the non-fertilizer or industrial sector dropped sharply, and it now appears that future growth will not return to the old trend line. Most of the non-fertilizer uses of sulphur and sulphuric acid involve older industries which are very mature, such as oil refining, steel, pulp and paper, pigments, etc. The use of sulphur in some of these industries is expected to exhibit little or possibly even negative

Jan 1, 1986

By J. H. Bassarear, P. F. Thomas

Large-scale equipment can reduce both capital and operating costs for concentrators. Large semiautogemus (SAG) mills can be effectively utilized in big mills but they must be kept fully loaded to tab advantage of their great potential. With varying hardness ores, controlling throughput is difficult and careful design must take this into consideration. A variable speed drive on large primary semiautogemus mills can overcome this problem to a considerable degree. This paper provides a base case using standard accepted technology of two wound rotor single speed 4476 kW (6000 hp) motors to drive the primary SAG mills of a hypothetical 36,364 t/d (40,000 stpd) typical porphyry copper mill. One other fixed speed drive will be described and jive variable speed drive configurations will be reviewed. Capital costs for the entire grinding circuit, from crusher product to flotation feed, were developed for the base case and each of the alternatives. A technical evaluation comparing the advantages and disadvantages of each drive is presented.

Jan 1, 1987

By B. Yarar

Samples of coal from an adit in the Fording River District of British Columbia showed hydrophilic properties between depths of 0-24.5 meters and was naturally hydrphobic at higher depths. Laboratory experiments indicated that weathered samples (i.e., those mined at 0-24.5m) could be upgraded by froth flotation at pH<2 following slimes removal. The i.e.p. of samples varied regularly between pH=1.6-4.1, the weathered fractions being at the lower end of the pH scale. These values followed a pattern comparable with the flotation behavior of samples. A patchwise structure for coal surfaces determining the resultant disjoining pressure, which in turn determines flotation-non flotation relationships is postulated.

Jan 1, 1982

By Carmel Bofinger, Maureen E. Hassall, David Cliff, Eric Stemn

e Springer Nature remains neutral with regard to jurisdictional claims in published maps and institutional affiliations.

By Leonard W. III Joseph, William H. Buttermore

Noise control is an important consideration in the design and operation of coal preparation plants. It is essential to understand and control noise sources to provide a safe environment for operating personnel. The average noise exposure levels in a typical coal preparation plant generally range between 95 and 100 db. An operator should not be exposed to noise levels of this magnitude for more than 4 and 2 hr, respectively, according to MSHA (Mine Safety and Health Administration) regulations. The major noise sources in a coal preparation plant include noise from impacting the inside surfaces of steel chutes, vibrating screens, coal crushers, pumps, motors, fans, and blowers. The magnitude of noise varies considerably from one situation to another. NOISE EXPOSURE IN COAL PREPARATION PLANTS Noise regulations established by the Occupational Safety and Health and- ministration (OSHA) are accepted by MSHA relative to coal preparation plants. Fig. 14-43 shows a graph of the time of noise exposure in hours per day to which a worker can be exposed for each A-weighted sound level. According to this graph, a worker is permitted to spend 8 hr in a 90-db noise environment, but that time must be halved when the A-weighted sound level increases 5 db. The noise dosage a worker receives is determined by the ratio of the length of time the worker spends in a particular noise environment divided by the length of time he is permitted to spend in that environment. If the worker is exposed to several different sound levels, his total dosage would be the sum of each of the individual dosages. The equation for determining the dosage is: where C is the actual duration of exposure at a given steady state noise level and T is the noise exposure limit for the level present during the time C According to MSHA regulations, the dosage, D, should not exceed one (1) for any worker for a full day of work. As an example of typical plant noise levels, vibratory screens and crushers emanate noise levels in excess of 90 db at a distance of 3 ft (0.9 m) with no material flow. With material flow, noise level increases 1 to 10 db depending

Jan 1, 1991

By M. G. Ayat

Coal degradation is the breakage of coal particles to finer sizes due to the continuous forces of attrition and gravity to which coal is subjected during mining, handling and preparation. The objective of this paper is to determine the extent of degradation that a coal slurry experiences when transported through pumps and pipelines. The examination of degradation results for a number of different coal slurries have indicated that the extent of breakage is dependent upon the physical properties of coal and the time span of circulation. These results have also indicated that certain size fractions of coal are more susceptible to breakage than others.

Jan 1, 1986

By Brain W. Lawrence

Sublevel stoping generally is a large-scale open stoping method. It sometimes is referred to as long-hole or blasthole stoping. This method usually is applied to strong ore bodies that require minimal support and are surrounded by strong country rock. The ore body should be fairly regular in shape and well-defined. The dip of the footwall normally should be sufficient to allow broken ore to gravitate freely, although the method has been adapted successfully to mine some of the flatter ore bodies. Sublevel stoping is not dependent on the width of the ore body. Widths of less than 6 m (20 ft), however, make the utilization of long-hole drilling techniques more difficult. The main criteria for sublevel open stoping are competent ore and stable host rock, regular ore boundaries, and a footwall dip that exceeds the angle of repose of broken ore. Basically, the method entails providing access to the ore body at various subintervals between the main haul- age levels in order to drill and blast the intervening ore. Stope drilling is carried out from drilling drifts on the sublevels, and the ore is blasted in slices towards an open face, which generally is vertical on the downholes and may be inclined towards the open face for the up holes. The blasted ore gravitates to the bottom of the stope and is collqcted through drawpoints. Fig. 1 illustrates a typical sublevel stoping mine. The method requires extensive ore body development with relatively high capital expenditures. How- ever, much of the development is in ore, and production costs are comparatively low. Productivity rates are in the 13.6 to 27.2 t (15 to 30 st)/manshift range. The drilling, blasting, and loading operations are performed independently, and equipment utilization is high. Large outputs can be obtained with few units and limited personnel. Dilution with waste rock may occur if ore boundaries are irregular or if caving occurs, but 100% of the ore within the stope usually is recovered. Pillar recovery sometimes is a problem. It is at this point that most of the waste dilution occurs. At the time data was gathered for this report (1973), the sublevel open stoping method was used by more than 21% of the metallic ore mines in the US (9 out of 42 mines producing 1088 + t/day (1200 + stpd) and provided more than 12% of the total metallic ore production from these mines. DEVELOPMENT Access to a sublevel open stope mine is gained by a slope or shaft that is normally sunk in the footwall of the ore zone, away from any possible effects from blasting or other production operations in the stoping process. Main level interval is selected and usually varies from 45.7 to 121.9 m (150 to 400 ft), depending on the vertical extent of the ore body, the number of drawpoints required to maintain output, and the eventual stope height. The main haulage drifts are located at the bottom of the stope horizon, either directly in ore or in the footwall, with access crosscuts at intervals. Raises are driven up through the ore to connect to the level above and provide access and ventilation to the sublevel drilling drifts, which are driven horizontally in the ore for the length of the stope. Depending on the drawpoint system to be used, the stope is either undercut by a raise-and-cone system, from a scram or slusher drift, or simply provided with a drilling drift from which access is made at intervals to the main haulage level. A raise driven either at the end or in the middle of the stope (if stope access is from both ends) provides a start for opening up a vertical slot across the stope to establish a free or open face for long-hole drilling and blasting. Sublevel stoping has proved to be a relatively safe and economical system.

Jan 1, 1982

By C. J. B. Dreyer, H. A. Robinson

Only South Africa can be considered in terms of resources of amphibole asbestos. There are two asbestos producing fields, one in the Northern Cape Province (producing only crocidolite), and one in the Eastern Transvaal (producing only amosite). This chapter discusses the banded ironstone sequence in which amphibole asbestos occurs, and the genesis of the fiber is briefly reviewed. Also, the properties and characteristics of amphibole asbestos are discussed and compared to those of chrysotile. Approximately 300 000 t of arnosite and crocidolite are produced annually. Crocidolite has a potential for a fivefold increase of production chat could be sustained beyond the year 2000. An estimate of the resources of the various producing fields in South Africa is made. The mining methods are also reviewed and evaluation techniques are discussed. The only two types of amphibole asbestos exploited at present in South Africa are amosite [the white colored amphibole asbestos) and crocidolite (the blue colored asbestos). As is well known, crocidolite has been mined in South Africa from the early part of this century, and amosite since the early 1930&apos;s. Since these dates, asbestos mining has developed into a significant mining industry. The annual production of crocidolite is approximately 170 000 to 200 000 t. and amosite production is in the region of 80 000 to 100 000 t. CLASSIFICATION OF AMPHIBOLE ASBESTOS Four types of amphibole asbestos are produced economically in the world today: (1) crocidolite, the fibrous equivalent of riebeckite, (2) arnosite, the fibrous equivalent of grunerite, (3) trernolite, and (4) anthophylite. The names trernolite and grunerite apply equally to the crystal and fibrous forms of these minerals. For comparison and clarification purposes, the six known varieties of asbestiform minerals known in the world are shown in Table 1. All the asbestos minerals were originally developed as cross fiber seams in veins or fractures, or were controlled by the lithology. Post- fiber effects caused slip, sheared, and folded fiber seams. Crocidolite and amosite occur in banded ironstone of early Proterozoic age of the Transvaal Supergroup. The fiber seams are developed in certain defined lithological units in the banded ironstone sequence. The lithological units have a remarkable lateral persistence with respect to their chemical nature and physical appearance. Vertically, the banded ironstone sequence displays a heterogeneous nature. Commercial Applications The features and properties that make amphibole asbestos commercially attractive are: 1) The long average length of the fibers is good for greater wet strength in composites. 2) The high tensile strength of crocidolite makes it superior to all other types of asbestos. 3) It is characterized by dispersion in most liquid media, particularly amosite in water. 4) Since the surface charge of amosite and crocidolite is negative, the fibers produce an open lattice work in water with a high bulk volume giving excellent filtration rates and rapid drainage in a wet asbestos cement mix. 5) The insulating properties of amphibole asbestos mats are still important, particularly in a corrosive environment, despite the introduction of many other synthetic insulating materials. 6) Corrosion resistance, particularly acid corrosion resistance. 7) Due to the high modulus of amphibole fibers, resins reinforced with these fibers exhibit

Jan 1, 1986

By Sathit Tandanand

Geological differences among various coalfields restrain the applicability of subsidence prediction using existing European methods. To modify these methods for domestic conditions, the Bureau of Mines, U.S. Department of the Interior, developed a method of assessment to evaluate the lithological effects on subsidence in the Northern Appalachian basin by examining data collected from 16 longwall panels in the coalfield. The assessment is partially completed, and the results to date show that the ratio of maximum subsidence to the extraction thickness, known as the subsidence factor, can be expressed in terms of the width-to-depth ratio by a simple exponential equation which has a coefficient tentatively considered as the subsidence index. This index varies with the lithology of a particular site and can be expressed in terms of the percent distribution of weak and strong rocks in the overburden.

Jan 1, 1982

By David B. Campbell

INTRODUCTION This submission was made to the Royal Commission of Inquiry into uranium mining in British Columbia. It discusses the stability of waste rock piles that might be developed in conjunction with uranium mining and the control of leachates that might be generated within these waste rock piles over the short and the long term. Short term refers to the interval during which a waste rock pile undergoes either continuous, or alternatively intermittent development, and in general can be considered to extend over the operating life of the mine. Long term refers to the postmining period, i.e. after the ore-body has been exhausted, and mining operations have ceased. Two major classes of waste material are generated in the course of mining and milling operations. Tails, which are the solid waste materials produced by the milling operations commonly exit the mill in slurry form and are stored in tailing retention ponds. Management and control of these slurried wastes are the subject of other papers to the Commission and this conference. In the course of the mining operations, it may be necessary to remove rock materials that are below cut-off grade. That is, ore concentrations within some of the mined rock may be lower than the concentrations which on the basis of economic considerations justifies milling the material. Rock which is removed in the course of the mining operations, but which is below cut-off grade is referred to as waste rock. This waste rock is commonly consigned to a specific area for disposal where it is placed in a mound or pile, i.e. a waste rock pile. Ore concentrations within the waste rock may vary from zero to just below cut-off grade. Since an increase in the market value of uranium could at some time in future result in reduction of the cut-off grade, it follows that waste rock containing ore concentrations slightly below today&apos;s cut-off grade could become ore grade rock at some time in future. For this reason, the mine operators may deem it advisable to separate waste rock according to grade, and to consign the waste rock to two or more individual waste rock piles. For purposes of clarity and simplicity, this submission treats the stability of waste rock piles and leachate control in a manner which assumes that these potential problems are isolated and separate from other aspects of the mining and milling operations. For many site specific cases, the problem of potential leachate control might be integrated with the tailings management aspect of the project, thereby realizing economic benefits while at the same time reducing the potential for adverse environmental impact.

Jan 1, 1980

By A. P. Schissler

2013 has seen metals, industrial minerals, and coal all reduce in price due to market and geo-political factors. Many companies are taking write-downs to recalibrate their reserves and have enacted focused cost reduction programs. The public domain includes statements where companies have operating cost reduction goals in the $500 million range. Zero-base staffing & budget analysis is a means to achieve optimal employment and cost levels for a given level of production. The method involves measuring the work to be done similar to Lean. Lean is a method to review each step of production and remove anything that doesn?t add value or is necessary to complete finished goods. Zero-base methodology has been applied to mines worldwide producing realistic and sustainable levels of staffing and costs based on the quantity of work. The objective of this paper is to provide the framework of zero-basing with examples of application.

Feb 23, 2014

By Serge Payant, Michael G. Li, Lionel J. J. Catalan

Alkaline solution rinsing for decommissioning of an acid-leached residue was evaluated by column experiments, neutralization tests and solid characterization by XRD and SEM. It was found that consumption of a strong alkali during residue neutralization was much greater than that required to neutralize the soluble acidity. The excess alkalinity was consumed by solid-phase conversions unnecessary for the success of decommissioning. Alkaline rinsing did not reduce the time requirement for decommissioning when compared with plain water rinsing. Since no benefit is gained to compensate for the additional costs associated with excess alkali consumption, we conclude that strong alkaline solution rinsing is in fact inferior to simple water rinsing.

Jan 1, 1998

By David Oliveira

"INTRODUCTIONOne of the key drivers for the selection of a mechanized tunneling method in good quality rock masses is a good excavation advance rate. In order to achieve satisfactory advance rates, a good geotechnical performance of the rock grippers in main beam open tunnel boring machines (TBM) is required, besides other factors related to ground support and the performance of the cutter head.One well known potential issue of main beam gripper TBM is associated with the bearing capacity of the tunnel side walls while crossing fault zones or poor rock masses in general. Such a mechanism is similar to that of footings on rock and perhaps the most common problem associated with gripper performance. Depending on the quality of the rock, the rock under the gripper contact may experience significant damage and crushing (Figure 1), eventually affecting the TBM advance rate due to a lack of tunnel wall reaction/support. Such a mechanism is primarily dependent on rock quality, applied loads and gripper size.In the case of large diameter TBMs (e.g., D > 12 m), a second potential failure mechanism may arise associated with the loss of confinement within a rock pillar between twin tunnels, even in good quality rock but particularly in bedded sedimentary strata. Punching of the rock pillar may develop due to excessive shear along bedding partings and sub-vertical joints (Figure 2), particularly under a low stress environment. As a result, this mechanism is dependent on both the rock pillar width between twin tunnels and stress conditions around the tunnels, in addition to the previous factors described for the first failure mechanism."

Jan 1, 2016

By James R. Arnold, John G. Mansanti

BACKGROUND Gold Fields Mining Corporation started their Chimney Creek operation in late 1987. The dump leaching operations commenced in October and the mill startup took place during the last half of November. The milling process at Chimney Creek is almost a conventional CIP plant with the major exception being the installation of a second thickener as a washing stage prior to CIP12 &apos; A simplified flow sheet is shown in FIGURE 1. Most of the gold in the mill is liberated by grinding in cyanide. After grinding in a conventional SAGball sill circuit, the cyclone overflow reports to the thickeners. The thickener overflow from the first thickener is pumped through 5 stages of carbon columns for gold adsorption. Slurry from the thickener underflow is leached, washed, re-thickened and pumped to CIP. After approximately 6 hours of retention time in CIP, the slurry is pumped to tailings for final deposition. The carbon from CIP is transferred to the third stage of carbon adsorption and then advanced as normal through the columns and strip circuit. Loaded carbon is stripped at approximately 1500C (3000 F) and 55 kPa (80 psi). After 4-5 hours of stripping, the grade of four tons of carbon is reduced from 5,140-6,860 g/t (150¬200 oz/ton) to typically 17 g/t (0.5 oz/ton). Zinc precipitation is used to recover the gold from the pregnant solution. After a water wash, the stripped carbon is transferred to acid wash or to the kiln feed tank for reactivation. All strip lots are acid washed in 3.0% HC1 and all stripped carbon passes through the kiln for thermal reactivation. Then the carbon is quenched, sized and stored in tanks or bins for transferral to CIP or to the carbon columns. Loaded carbon from the dump leach circuit is treated in a circuit consisting of five carbon columns which are parallel to the mill circuit. No distinction is made between the mill or dump leach carbon once the gold has been stripped from them. CARBON FINES HANDLING Once the carbon has been transferred from the carbon columns, every effort is made to minimize carbon attrition and to recover the carbon fines that have been generated in the process. Wemco Hidrostall pumps are used for transferring all carbon in

Jan 1, 1990

By William G. Kegel

Apart from a usable product and good &apos;mining conditions, the greatest asset for a profitable coal mining organization is an effective mine maintenance program. The first step in achieving this is to have good mechanics-or men who want to be mechanics and are interested in knowing more about the design and control of the machines in their charge. A good mechanic&apos;s knowledge and ability result from ambition, curiosity, and determination. Starting with these qualities, the aspiring mechanic can launch his career by learning all he can about mechanics, electricity, and hydraulics. Though the path is rough, the end result is rewarding. Mining machinery probably receives as much, if not more, abuse as any other heavy equipment in use today. Considering the conditions under which much of the repair work must be done, it is essential that the mine operator has the best men and tools to perform the work. Mining machines, from their inception, have utilized the electric motor; and down through the years of mechanization, this power source has undergone many changes to better adapt it to the job. The first motors used were direct-current (d-c) motors since d-c power was the most economical power source for mining at the time of early mechanization. These motors were large open-type machines which are seldom used in modern mining. Because of the mine environment, the d-c motor has been reduced in size to better suit the application in mining machines and is now manufactured with dust-tight or explosion-proof enclosures. The d-c motor is constructed with either (or both) series and shunt fields located on the stationary part of the motor (stator) and- an armature coil wound on the rotating part of the motor (rotor). Power is supplied to the armature by brushes which are held stationary and make sliding contact with the armature. The brushes requires regular and careful maintenance because worn or defective brushes will cause heat or arcing which can destroy the motor. D-c motors can be used as d-c generators when mechanical power is supplied to them by the rotating shaft. In the case of a locomotive, this principle is used in dynamic braking. The momentum of the locomotive drives the motor and generates d-c power. The power is regulated with resistors and provides opposition to the motion of the locomotive.

Jan 1, 1981