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By P. K. Hopke, A. Hubbard, K. H. Leong, J. J. Stukel, K. Nourmohammadi

INTRODUCTION In order to understand the transport and deposition of radon daughters in mine atmospheres, it is necessary to know the variation in the attachment of the daughter atoms to particles as a function of particle size, composition, number density, relative humidity, temperature, and radon concentration, the free gaseous diffusion coefficients of the daughters, and the variation in the mass transfer of the activity, both free and attached to particles, to mine surfaces as a function of particle size distribution, surface roughness of the mine walls, and the flow conditions. If all of these parameters are known in a model system, it should be possible to understand the transport and fate of the airborne radioactivity in real mines under certain well-defined flow conditions. There have been a number of recent investigations of the attachment of radon decay products to particles 1-4, but there are still a number of unanswered questions regarding the process. However, it is clear that for most real mine atmospheres, the vast majority of the activity is attached to particles. The size distributions for the activity-bearing airborne particles have been studied 5,6, and it has been found that most of the activity resides on particles with diameters in the range of 0.05 µm to 0.3 µm with an average mass median diameter between 0.1 and 0.2 µm. The behavior of the unattached radon daughter species has also been recently studied[ 7], and many of the previous problems regarding the value of the diffusion coefficient for Po-218 have been resolved. A major problem in the understanding of the airborne transport of radioactivity in mines is the lack of detailed knowledge of mass tranfer to and fluid flow over rough walls under fully developed turbulent flow conditions. This paper will report the progress on a project that is designed to obtained that information. MATHEMATICAL MODEL DEVELOPMENT Deposition of particles on smooth surfaces in turbulent flow has been extensively studied. A comprehensive review of these results has been prepared by Sehmel 8. There has not been such a comprehensive study of particle deposition on rough walls under such flow conditions. In recent years, only a single model has been proposed to explain such deposition 9,10 and in both of these papers the flow structure in the rough walled pipe was not taken fully into account. As part of the work being conducted on this project, a more complete model was outlined in a previous report [11]. The basic theory will be reviewed to provide a context for the flow measurements to be reported. The flux of particle deposited on the walls of a pipe in a turbulent flow is derived from the one dimensional form of Fick's law as given by [N = Dpdpp/dr (1) where N is the flux of particles deposited per unit area per unit time, D is the total eddy diffusivity of the particles, p is the airborne concentration of particles, and pr is the distance measured from the center of the pipe. The rate of deposition is best expressed by a deposition velocity Vp = NIP pb (2) where P b is the mean particle concentration in the sulk flow. The shear radius, V/ut and the shear velocity, u , are used to calculate a nondimensional distance, and velocity, respectively, where v is the kinematic viscosity of the fluid. The nondimensional form of equation 1 is given by Vd = DP dpp(3) V dr+ where Pp = Pp/ Ppb(4) By integrating equation 3 from the rough wall stopping distance, S , to the center of the pipe, the deposition velocity can be obtained. In order to make this calculation, it is necessary to have accurate descriptions for the particle eddy diffusivity, stopping distance, and shear velocity in order to insure that the influence of the flow structure has been properly accounted for. The shear velocity can be determined experimentally from the shear stress evaluated at the wall, Tw, and the fluid density, ut =VT w/p = ub V f/2 (5) where ub is the mean bulk axial velocity. The wall shear stress for a given pressure drop, dP/dL, and hydraulic diameter, Dh, is]

Jan 1, 1981

By F. A. Molloy

As a result of feasibility studies made in 1975 a 3-0 seismic reflection survey was shot over part of the Veendam salt pillow during 1978-1979. The Zechstein salt pillow, located in the northeast of the Netherlands near the town of veendam, occurs at a depth of 1400 m and contains three carnallite rich horizons over an interval of 300 m below top salt. MgCl2 brines are produced from these layers as feedstock for a magnesia plant, which starts up in 1981. Prior to the acquisition of the seismic data four wells had been drilled. The thickness of the carnallitite layers encountered in these wells varied between 50 and 125 meters. In view of these variations it was necessary to obtain a reliable isopach of the different layers. The 3-0 seismic survey covered an area of ± 4x4 km with a sampling interval of 20 m in all directions. The data was recorded it an area array using a 120 channel instrument. Final presentation of the data was in the form of 3-0 migrated acoustic impedance cross sections. The quality and resolution of the seismic data allowed reliable structural and isopach mans to be made. It was shown that the carnallitite layers are continuous over the area surveyed, whilst local thickening is "fault" related. Having this information brine production locations could be choosen in the thicker parts, allowing maximisation of production potential whilst avoiding faults and fulfilling mining regulations. A subsequent benefit of this approach is the possibility of using the seismic data to predict ore grade. Using the available wells as calibration points the acoustic impedance response of the seismic at the well was matched to the ore grade found. From these points a synthetic section was made with varying ore grades. Pre¬liminary results are encouraging and the liability of this predictive tool will be evaluated in the light of future drilling.

Jan 1, 1981

By J. R. Odekirk

The nature of mineral distributions and locking characteristics in mill products are important factors in understanding mill performance. Such data are obtained by microscopic point counting methods. Reduction 'of point-count data from volume to weight percentages, and reduction of data for each size fraction back to a head sample requires lengthy hand calculations. The use of computer processing for such data reduction results in obtaining information on a realistic time basis. An example of computer processed point-count data is presented.

Jan 1, 1981

By L. B. Hales

December 3, 1972, James Wm. White presented the results of an instrumentation and process control survey which was sponsored by the Anaconda Company at the annual Tucson Subsection AIME meeting.(1) Twenty-six responses were obtained from questionnaires sent to 48 operating plants throughout the world. The survey results were discouraging in that there were more failures and problems than examples of success. The following quotes are taken from the text of the paper. "We run our mill by the seat of our pants." 'If these blue sky Ph.D.'s (in the instruments and controls area within research) had some operating experience they wouldn't make a lot of the suggestions they do."

Jan 1, 1981

By Russell G. Shomperlen

Grinding mills in excess of 5 meter diameter and 3500 kilowatts per mesh are putting new demands on big gears. This paper returns to the basics of ring gear design beginning with the fundamental power formulas specified by AGMA. Specific emphasis is given to those elements in the rating equations which require closer attention as gear diameters and transmitted power increase. Strict process control in the manufacture of the gear blank is also stressed. Finally, the writer offers a guide to proper installation, lubrication and maintenance.

Jan 1, 1981

By William L. Barnes, Maurice Leiser

INTRODUCTION The literature on project management considerations and techniques for large rail rapid transit projects in mid-1979 was fairly large and getting larger. UMTA was establishing special organizational elements to deal with the subject as it related to those properties requesting capital assistance, and comprehensive project management plans were being requested from those properties as an early step to build confidence in UMTA that the particular project could be planned, designed, and executed efficiently and effectively, thereby using critically short capital assistance funds to best advantage. In the late 60's and early 70's, the choices of models for heavy rail construction would have been quite limited.., and perhaps much simpler than was the case in 1979. One of the writers was with an agency in 1979 in the process of finalizing its organization for project implementation after several years of study and had the opportunity to review the experiences of others at length, presumably to build on the success and to try to avoid the pitfalls and failures. The properties that were surveyed all reported more or less success stories. One could only conclude that there are many ways to organize for a major transit (construction) project and that there is a great variety of viable management structures, but it seemed that chances for success were best where the executing property had prepared a comprehensive project management plan, then resolved to follow it. Perhaps more important, a commitment or prior approval had been sought from UMTA as to the validity of the project management plan, and the property was allowed to proceed without detailed prior UMTA review or approval of all specific intermediate steps. Most important, however, an early overall UMTA funding commitment had been obtained that permitted some reasonable expectation of the rate of funding, perhaps by commitment

Jan 1, 1981

By W. Fletcher

Startup of a new plant is a very interesting time. It generally is also a very demanding and often frustrating period. Startup problems can become serious ii they impede the early achievement of full plant production. Raw materials are wasted, manpower is unproductive, field modifications can be extremely expensive and market commitments may not be met. This paper outlines a well proven and highly successful approach to eliminating many of the problems and frustrations often experienced in plant startup. The procedures and methods summarized in the following paragraphs have been developed over a period of 25 years during which some 90 plants were started up.

Jan 1, 1981

One of the best known of the early mining districts of Nevada was that at Eureka, where gold was first produced in 1864. While the most productive period for the district was during the early 1900s, production was maintained until World War II. During this time it is said that the Eureka district produced as much as $200 million from over 3,200 km (2,000 miles) of tunnels and workings. The Windfall Mine was discovered in 1902; and a 100 ton cyanide plant was operated there during the periods 1907 to 1912 and 1932 to 1940. Since 1974, however, Windfall Venture has been operating a unique heap leaching facility on the site of the old Windfall Mine. Gold at the Windfall occurs primarily in the Cambrian Hamburg dolomite which has been "sanded" and in which the original cementing material has been destroyed. The gold is micron sized and is accompanied by anomalous amounts of arsenic and mercury; however, the ore is low in base metals which distinguishes it from the typical Eureka district ores. Mining the ore causes it to break down into very fine sizes, approximately 50% minus 150 microns (100 mesh). Conventional heap leaching by spraying proved to be impossible; and extensive test work on alternate methods, aided by the United States Bureau of Mines, was conducted using 9 m (30 ft) high columns and then large tonnage heaps. The leaching procedure adopted was to build 10.6 m (35 ft) high heaps on a 50 mm (2 in.) asphalt pad covered with a layer of gravel for drainage assisted by perforated pipe drains and gravel-packed sumps. A berm is constructed around the top of each pad and cyanide solution is allowed to “pond” on top of the heap. The resulting percolation is at the rate of 24 liters per sq m per hr (0.6 gal per sq ft per hr), a figure about 2.4 times the average used in other heap leaching operations. The dolomite furnishes its own protective alkalinity and no lime or caustic is used in the solution which contains 0.4 kg per mt (0.8 lb per st) of NaCN. Pregnant solution from the heaps averages about 0.7 g Au per mt (0.02 oz per st) with a cyanide content of about 0. 2 kg per mt (0.4 lb per st) and a pH of plus 9. Addition of NaCN brings the pH of the leach solution up to 11 to 12. Carbon adsorption of the gold was originally done in six enclosed columns mounted in a trailer but five conventional columns have been substituted as the size of the operations has increased. Carbon stripping is with hot caustic, cyanide and 10% ethanol solution. The

Jan 1, 1981

By Dan Crackel, G. G. Ramos, Mark Heisel

INTRODUCTION Feasibility studies indicated that a cratering type stoping method was best suited to the character of the first ore block at Carr Fork. This method known as vertical crater retreat (VCR) has since been modified to a combination cratering and deck blasting using VCR drop raises for initial undercut development . The modified form retains many of the advantages of VCR while allowing high tonnage production blasts which have offset many of the disadvantages. FIRST CONCEPTS AND PREMISES Geologic Considerations The Steep Highland Boy (SHB) skarn orebody, the first of several blocks being developed, is a steeply dipping 70° to 90° limb of a syncline, with a width averaging 25 m (82 ft). The bed has a consistent east-west strike length of over 400 m (1300 ft) and extends from the surface to depths of over 1000 m (3300 ft). The footwall is highly fractured quartzite and the hanging wall grades from garneti te to limes tone and hornfe 1s. Because of the limestone in the hanging wall, the hanging wall appears more competent than the footwall. The ore is garnetite with lenses of massive sulphide and magnetite. The quartzite footwall is highly fractured not only because of the bedding but also due to the folding and intrusion of the quartz monzonite. This irregular footwall outline of some stopes in- creases the likelihood of excessive dilution. The principal ore mineral is chalcopyrite with some bornite. The garnetite ore is quite variable in terms of compressive strength and drillability. In addition to the sulphide and magnetite lenses, it is transected by clay- altered dikes and major and minor north-east faults. TABLE (1) gives the range and mean strengths of the rocks. Investigating Mining Methods Initial consideration of a suitable mining method centered around maximum safety and efficiency with minimum potential dilution. The following major methods were considered and investigated. Cut and fill mining was eliminated because of low productivity, the cost and problems associated with providing fill, and the relatively unsafe nature of the method. Sublevel caving was rejected due to the unpredictable cavability of the waste and the geometry of the ore body. Conventional sublevel or blasthole methods require excessive development, as fan and ring drills are limited to efficient drilling in the 18 to 25 m range (60 to 80 ft). Mining downdip with equipment drilling small diameter holes would have required eight separate extraction/drill horizons with crosscuts and associated excavation in the main ore block. The use of down the hole (DTH) machines cap- able of drilling 165 m (6g in) diameter holes

Jan 1, 1981

By W. Clark Ashby

Trees are an essential ingredient of sound reclamation in nearly all surface mining regions. There are, however, possible problems in current reclamation regulations and practices which need to be evaluated and remedied. Recognition and accomodation of growth requirements for trees and other plants will greatly increase reclamation success. Surface mining encompasses a series of processes from a pre-mining baseline study to bond release following revegetation. Land owners, and society, have a great interest in subsequent plant growth on reclaimed areas and in future land values.

Jan 1, 1981

By G. A. Savanick

The Bureau of Mines has designed, fabricated, and field tested a water jet cutting device capable of perforating nonmetallic well casings for the purpose of completing or stimulating in situ uranium leaching wells. The device is lowered into the wellbore, where it issues a high-velocity water jet that penetrates either a polyvinyl chloride or fiberglass casing, cement, and from 1 to 4 feet into the surrounding uraniferous sandstone. The resulting perforations allow leaching solution to pass between the sandstone and the wellbore. This report describes the water jet perforator, tells how it is used, presents cost data, and discusses the results of laboratory and field tests. The tests were designed to (1) determine the optimum operating conditions for water jet perforation, (2) assess the effects of changes in these conditions, (3) evaluate the performance of the water jet perforator under actual conditions, and (4) compare this method with other methods of well completion. These tests showed that water jet perforation is superior to conventional well completion methods for selectively achieving communication between the well-bore and adjacent sand while maintaining sand control, for restoring permeability lost through wellbore damage, for stimulating poorly performing wells for completing wells with casings that have inside diameters as small as 2 inches, and for extending the effective well diameter.

Jan 1, 1981

By C. L. Pillar

INTRODUCTION The location, size, character of the ore de- posit and its adjacent formations will deter- mine the mining system best used for its ex- traction. The mining system chosen will have to be engineered to suit the ore deposit and the people that will extract it. What can be done in the Philippines cannot always be accomplished in the upper Northwest Territory of Canada. Mining operations developed under a free enterprise system are planned to achieve a satisfactory return on the capital required to place them into production. Hence, the choices made in planning and developing the mining system must be dictated by sound economics. The mining system planned must favor: 1. Maximum ore extraction of the deposit with minimal grade dilution. 2. The amount of capital expenditure used to develop and equip the mine must be in harmony with the financial policy of the group yet be sufficient to provide a reliable mining system. Excessive frills and extras that are not required for efficient, safe production should not be provided. 3. Mine production costs are determined by the amount of men, material, equipment and energy used in the production process. These must be kept to minimal limits in order to insure dependable, low cost production. Extras that later are needed in the production operation can be added as required by the operator. Basically block caving is a mining system that is normally used to extract deep seated, massive, low grade deposits of copper, molybdenum and iron where stripping ratios for open cast mining are economically prohibited. The caving system is limited in its selectivity. It is a mass production type of underground ore extraction that under favorable conditions is one of the most efficient, low cost mining systems in the world today. Usually block cave mining systems are large. They are capable of producing ore at a rate of 10,000 to 100,000 tons/day from large deposits at costs that are comparable with many open cast mining systems burdened by high waste to ore ratios. Block cave mining systems differ basically in the plan of the area layout, the method of undercutting and the method of draw. The balance of the mining system is the mechanical transportation of the drawn ore to the surface, the mine supply, drainage and ventilation systems. This paper deals mainly with the methods of draw. The method of draw to be used in a block caving system dictates the development plan of the mining area, is highly influential in the method of undercutting to be employed and the supply and ventilation system to be used. There are three basic methods of draw commonly used in today's operations. They are: Gravity draw Slusher draw LHD draw The type of draw planned will have a major affect on: Capital expenditure and time required to bring a deposit into production. Productivity, efficiency and cost of the production operation. Percent recovery of the ore deposit in both tonnage and grade. Degree of production reliability Fundamentally the type-of draw used must be compatible with the physical characteristics of the deposit to be extracted and the financial position of the mining operation. The type of draw chosen must work to advantage with the size, configuration, depth, strength and fracture pattern of the deposit and its adjacent formations. It must favor the financial protection of the planned mining operation. A Gravity Method of Draw requires a finely fragmented, easy flowing product from the caving action of the deposit. To obtain the desired economic advantage the ore deposit must also have sufficient area and height of column to permit minimal development cost on a per ton basis. The gravity system re- quires the maximum amount of development openings and uses the minimum amount of mechanized equipment of all the caving systems. It is unskilled, labor-intensive and is highly sensitive to size of ore fragmen-

Jan 1, 1981

By Robert P. Stevens

It is important for a refractory supplier to make systematic petro-graphic examinations of products which have been in service in the furnaces and kilns of customers. The petrographic examination of refractory service samples involves the microscopic study of polished and/or thin sections, supplemented by electron probe microanalysis, as well as the bulk chemical and XRD analysis of zones in the samples. The evolution of the A. P. Green Refractories Company petrographic laboratory is described, and the application of petrographic methods is illustrated in the examination of service refractories from a Portland cement rotary kiln and an AOD stainless steel vessel. The objectives of such studies are to assist the customer in the solution of his refractory problems, and to accumulate data which are of use in the design of improved refractory products.

Jan 1, 1981

By Slavoljub D. Maksimovic

At some coal mines, methane gas is released during raise drilling. At the request of the Mine Safety and Health Administration (MSHA), U.S. Department of Labor, the Bureau of Mines conducted ventilation surveys at different raise drill sites to assess various methods of reducing methane concentration. The resulting data indicate that raise drill shaft cavities are ventilated by air compressors, vacuum pumps, auxiliary blower fans, and free intake air from the surface. The compressors, pumps, and fans operate simultaneously or separately. At the bottom of the shaft cavities, ventilation is controlled by regulators and airlocks. Methane is monitored, measured, and periodically sampled on the surface, at the bottom, and at different levels in the shaft cavities. Maximum concentration of 3.5 percent was measured in one shaft in West Virginia. Air velocity down the raise drill shafts is always very low, Maximum recorded air down the drill stem was 3,000 cfm, which can be increased by adding air compressors. The best way to ventilate shaft cavities is through the simultaneous use of air compressors and vacuum pumps. If the methane concentration cannot be reduced with additional compressors, then the inside diameter of the drill rod should be increased, or larger pilot hole drill bits should be used.

Jan 1, 1981

By Claude A. Goode

Large quantities of high grade coal exist in thick seams in the United States. Many of these thick seams are too deep to be surface mined and do not lend themselves readily to underground extraction. To develop the new mining technology required to effectively mine thick seams, the Department of Energy has entered into a Cooperative Agreement with Mid- Continent Resources, Inc. to demonstrate the multilift longwall mining technique in a 28 foot-thick seam at their L.S. Wood No. 3 mine near Redatone, Colorado. The seam will be extracted in two lifts, with a 5-foot natural partition of coal left between the upper and lower lifts. Proper selection of face supports is critical to the success of this project. Specifications were defined for the roof supports, and it was concluded that the supports selected must provide full coverage, yield at the legs, and have a 60-inch push- out cantilever that can be set at any position under full load. A Hemscheidt two-leg shield was judged as the best qualified candidate and was selected for the multi-lift trial.

Jan 1, 1981

By Thomas A. Palmer

Free cyanide has been successfully determined in ground and surface water by isolation using a microdiffusion technique followed by a pyridine-barbituric acid photometric finish. The determinable free cyanide includes hydrocyanic acid, ionic cyanide, and the cyanide of weak complexes such as cadmium, lead, manganese, and zinc cyanide. Complexes of copper, iron, nickel, and silver are sufficiently strong so the dissociation into free cyanide does not occur at the microdiffusion pH 7. Cyanate, sulfide, and thiocyanate do not interfere. Samples containing as little as 0.005 mg/L are successfully analyzed. The relative standard deviation is <10%. Samples stored at pH >I2 in the dark are stable at least three weeks.

Jan 1, 1981

By A. L. Paverd

The petrology of a number of pyrometasomatic mineral deposits in north Queensland, Australia, were studied to evaluate the use of mineralogy and major element chemistry in the exploration for copper, lead and zinc-bearing skarns. Results indicate a direct correlation between the character of the mineralization and the composition and nature of the host rock. The study of garnet compositions proved the most useful. Economic mineralization was limited to limestone-derived skarns containing garnet of And60-90, whereas skarns containing garnet of less than contained less than 1And55 contained less than 1% total sulfide. Skarns containing garnet of And60-8S host the best mineralization; above And8S&apos; copper, lead and zinc sulfides decrease and magnetite and pyrrhotite become the dominant ore minerals. Similar correlations were found for pyroxene, amphibole and &apos;the bulk composition of the skarns.

Jan 1, 1981

By Thomas M. Lynberg

WIDCO is a surface coal mine located near Centralia, Washington. The WIDCO preparation plant processes over 7 million tons of run- of-mine coal each year with a product yield averaging 72%. In the jig type preparation plant the fine refuse amounts to approximately 130 dry tons per hour of 28mesh by 0. This fine refuse material is thickened in a 160ft. diameter static thickener with the use of lime as a coagulant and organic polymer as a flocculant. Once thickened the fine refuse, which is predominately bentonitic clays, has been dewatered and disposed of using several different methods. These methods will be discussed and include vacuum disc filters, ponds, solid bowl centrifuges, belt filter presses and experiments with electrokinetic dewatering.

Jan 1, 1981

By Aladar B. Dory

INTRODUCTION Anyone familiar with the problems of hardrock mining will agree that the majority of the serious dangers present in mining are quite visible and obvious to any person reasonably familiar with the profession. Having unsecured, unscaled back over ones head, gives one a very good chance of ending up under a caved in mass of rock. Staying too close to a blast gives one almost a certainty of being hit by a flying rock. Too little oxygen in the air will very quickly lead to loss of consciousness and death. One walks only so much over deep, unsecured openings before he falls into them. It is because of this clear visibility of the conventional health and safety hazards that mining regulations in almost all jurisdictions world-wide are a more or less comprehensive collection of "shalls" and "must nots" of good common sense. When basic rules of common sense safe working practices are at stake, there is little room for dialogue and compromise. The mine inspector is then observing, during his inspection, how well the mine follows these common sense rules. RADIATION AS A HIDDEN DANGER Radiation in mines is a risk, the impact of which does not demonstrate itself immediately. It is first of all a potential risk. Two individuals exposed to identical radiation will almost certainly be effected differently, if at all. This is certainly true of exposures and doses one might encounter in the mines today. We hear very often the phrase: "there is very little known about the effects of radiation". It is one of the most misused and misunderstood half-true statements. I would doubt that there is any other carcinogen whose effects have been studied as extensively as the health effects of radiation. Where the statement is correct is regarding the knowledge of the quantitative assessment of the risk of low level radiation exposures. The reason for this uncertainty is that the magnitude of their health effect is very close to the health effects of natural radiation, cosmic radiation and the effects of other carcinogens such as industrial pollution, hydrocarbons from cars and other chemicals we have grown accustomed to using. As far as lung cancer is concerned, the effects of wide use of tobacco probably outperforms any other single substance. All this having been said, the bottom line is still unchanged. Radiation exposure, in most cases mainly radon daughter exposure, was and still is one of the health hazards of uranium mining and as such has to be controlled to the best of our ability. Various jurisdictions have adopted different approaches to the control of radiation exposures of uranium minemill workers. The following sections of this presentation will attempt to explain the regulatory approach taken in Canada. THE CANADIAN REGULATORY PHILOSOPHY As indicated earlier, the health effects of low level radiation are quantitatively not yet defined and no proven threshold of radiation exposure exists. The Atomic Energy Control Board&apos;s (the Board&apos;s) regulatory system is based on the basic assumption that there is no absolutely safe limit of radiation exposure below which there are no health effects. Theoretically we should therefore strive to reach zero exposure. It is obvious that this objective cannot be reached in real life. The objective of the regulatory process therefore has to be to achieve radiation exposures of the workers that are as low as reasonably achievable, social and economic factors taken into account. This, of course, is the internationally acclaimed ALARA principle put forward by the International Commission on Radiological Protection (ICRP). To avoid any misunderstanding it is worth emphasizing that the ALARA principle is applied to achieve exposures below the regulatory limits which must not be exceeded in normal operation of any nuclear facility including uranium mines and mills. The present regulatory limits for radiation exposures of atomic radiation workers are based on the recommendations of the ICRP and they are almost universally accepted. They should ensure that the risk from radiation exposure is not greater than the risk associated with working in a comparatively safe industry. Basically, there could be two extreme approaches to the regulation of uranium mining and milling. One extreme approach is to develop very extensive and detailed regulations and requirements covering all aspects of radiation protection. This is a somewhat autocratic approach to the regulatory process. This approach has two very serious shortcomings. If detailed requirements are set in regulations, due to the great variations of actual conditions at various mine-mill facilities, they have to be set as a compromise between the desirable requirements and those which could be met by practically all facilities. This approach takes away from the management of the facilities the initiative to strive for improved conditions. Requirements are spelled out in clear, understandable targets and the only worry of the management is to comply with these targets. One of the basic duties of management is to manage the operations in the most effective way with the maximum health and safety of the workers in mind.

Jan 1, 1981

By William G. Kegel

Apart from a usable product and good 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

Jan 1, 1981