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By George K. Schenck

This Bureau of Mines publication describes and evaluates a Standard Industrial Classification (SIC) based end-use data system for minerals that is designed to link the demand for minerals to economic activity in commercial uses. The new data system is two-dimensional. The first dimension of technical product classes (TPC's) measures the consumption of minerals in major primary intermediate products. The second dimension measures the consumption of these TPC's in SIC end-use sectors in sufficient detail to identify the major determinants of demand for each TPC. Intercomodity comparisons are achieved at the level of SIC two-digit major groups. The methodology was developed so that all available sources of information from Government and industry can be utilized at minimum cost. This information will be supplemented with additional data to be obtained by modifying existing Bureau of Mines canvasses so as to elicit additional quantitative responses from mineral producers. The new data system is expected to improve significantly the ability to forecast demand for minerals, and provide an information base that would, for the first time ever, permit resolution of such questions as trends in substitution among materials and the intensity-of-use of materials in end-use sectors. For the 42 TPC?s identified in the report, data are available for the new system from current sources or, except for three TPC?s that require new canvasses, can be obtained by modification of existing Bureau canvasses.

Jan 1, 1982

By Douglas W. Huber

During the past decade, a general awareness of the need for and subsequent development of increased safety in United States coal mines has taken place. Government and industry alike were stimulated into accelerating their actions through unacceptable accident statistics and costs (in terms of human misery dollar values). Those of us who work in the United States coal industry are aware of the development of governmental controls on nine safety. Both Federal and State regulations have placed increased responsibility upon mine owners and operators to insure a safe working environment and to provide training for miners. No other industrial group of workers has received such attention or, in our opinion, deserves it more.

Jan 1, 1982

By Milton E. Jennings

On May 11, 1977, the SAMARCO 508-mm (20-in.), 396-km pipeline was placed in service. This system represents the latest thinking in slurry pipeline design and is one of the main contributors to the viability of the SAMARCO project. This paper presents the background, description, and latest information on the system's operation. Special attention is given to the handling limitations of the slurry and to the unique hardware requirements.

Jan 1, 1982

By Warren D. MacEvoy

INTRODUCTION Good ventilation bulkheads (stoppings) can be obtained only by conscientiously adhering to four basic principles: carefully selecting the site for the bulkhead, adequately preparing the site, properly constructing and sealing the bulkhead, and checking the results with a smoke tube before leaving the area. If all these guidelines are followed during each ventilation project, a poor job will result only from the use of improper materials and cannot be blamed on carelessness or oversight. SITE SELECTION A perfectly constructed bulkhead in the wrong location can be completely worthless. Conceivably, a bulk-head in the wrong location could cause dangerous ventilation conditions that are difficult to identify and isolate. If the suggestions herein are followed, a considerable amount of work, time, and material may be saved. Perhaps the single most important step is always to select the exact site for a bulkhead before starting construction. Although the latitude in selecting a site may be limited by instructions or operating conditions, it is common to find poor bulkheads built within a few meters (feet) of a site that would have been far superior but was overlooked during the planning process. The ideal site for a bulkhead should have as many of the following features and conditions as possible: 1) To provide a solid bulkhead, the selected site should be in firm and unbroken ground. 2) To minimize the amount of work and material required and to minimize the danger of later damage to the bulkhead, the site should have as small a cross section as possible. 3) The bulkhead site should not be encumbered by interferences such as pipes, wires, ditches, wire roof supports, trash, muck piles, etc. 4) The walls and timbers of the site should be free from oil, grease, or tar that would inhibit adhesion of the sealant used on the bulkhead. 5) Unless special precautions are taken, the site should not have water seepage, standing water on the floor, or water-carrying ditches. 6) To expedite construction of the bulkhead, the site should have reasonable access to transportation, supplies, communication facilities, and compressed-air lines. 7) The site should have a reasonably level floor, allowing direction of the bulkhead door swing to be reversed at a later time if so desired. In a steeply sloping location, it may be possible to open the door in one direction only. SITE PREPARATION Once a suitable site has been selected for the bulk¬head, it must be prepared properly. All site preparation work should be completed before starting any construction work on the stopping itself. Unfavorable site conditions can be identified during the course of thepreliminary site preparation, saving time, effort, and materials that otherwise might be wasted. The follow¬ing preparation steps help assure the construction of a good stopping: 1) If wire mesh has been used at the site, a strip 457 mm (18 in.) wide should be cut and removed from the walls and roof at the selected location. 2) The exposed strip of rock should be barred down thoroughly to provide a smooth surface. 3) All obstructing materials should be removed, including old timbers, pipes, rockbolts, wires, etc. If a conduit must cross the bulkhead area, it should be located or relocated away from the floor, walls, and roof to allow a 6.28-rad (360°) seal around the juncture between the conduit and the bulkhead structure. 4) All loose muck should be cleaned from the site and a trench about 152 mm (6 in.) deep should be dug from wall to wall in the floor. The trench must be wide enough to accommodate the entire bulkhead, including the posts. 5) The rock surfaces of the walls and roof should be cleaned with a wire brush to remove as much loose surface material as possible. Thorough cleaning promotes adhesion of the bulkhead sealant to the surfaces of the walls and roof, thus promoting an airtight seal. 6) It is quite difficult to seal landing mats that cross a bulkhead. If such a crossing cannot be avoided, the bulkhead should be placed between the end and the first hole in the steel so at least one side of the bulk-head can be sealed easily and completely. That may have to be done prior to actual construction if the end is located on the opposite face from the seal coating of the bulkhead. BULKHEAD CONSTRUCTION The three principal considerations in bulkhead con¬struction are the type of bulkhead, materials to be used, and construction method to be employed. For many projects, the bulkheads are specified by the requesting agency, with no latitude for independent choice of the type, materials, or construction method. In such cases, any deviations from the specifications, for any reason, must be approved in advance by the proper department or by the project senior ventilation engineer. Types and Materials When a choice of type, materials, or method is allowed, consideration should be given to factors such as cost, required useful life, proximity to blasting concus¬sions, availability of materials, direction of permissible air leakage, and degree of airtightness required. Other factors to be considered include the amount of time available, accessibility of transportation, potential for interference with operations or production, ambient water conditions, availability of connections to com¬pressed-air lines, etc. Despite the multitude of factors to be considered, most stoppings can be analyzed easily and the proper choices can be made without much difficulty. The four common classes of bulkheads utilized in

Jan 1, 1982

By L. L. Oden

The Bureau of Mines analyzed vapor-aluminum-diffused stainless steels prior to conducting corrosion measurements in severely sulfidizing environments. These studies were conducted in response to the national need to better control the environmental and economic costs of corrosion. Specimens of a ferritic alloy and austenitic alloys types 316 and 310 were treated by a proprietary vapor-aluminum diffusion process and a dual vapor-aluminum diffusion process. Weights and dimensions were recorded before and after treatment, and comprehensive metallographic, X-ray, and electron micro-probe analyses were conducted for both single-and dual-vapor-aluminum-diffused specimens. Surface aluminum concentrations ranged from 26 to 34 wt-pct for all alloys. Dual treatment increased the quantity of aluminum per unit area by a factor of 2 and increased the depth of penetration by a factor of 3. The aluminum-diffused region of the ferritic alloy was single-phase aFe solid solution that was ordered on the surface, whereas two distinct diffusion layers were ob-served on the austenitic alloys. The outer layer comprised aFe solid solution that was ordered on the surface, and the inner layer was a two-phase mixture of aFe solid solution and the interrnetallic compound AINi.

Jan 1, 1982

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By P. M. Daling

This report, sponsored by the U.S. Bureau of Mines, presents the results of an examination of the safety of the impactor shaft sinking system in comparison to conventional shaft sinking technology. Detailed study of the impactor system was also performed to identify and recommend design changes for optimum safety conditions. A comparison of the impactor machine and its support systems with conventional shaft sinking was made on the basis of relative risk levels. Risk incorporates measures of both the frequency and the potential consequence of accidents. A safety evaluation of the component operations of the impactor shaft sinking system was performed using a Failure Modes and Effects Analysis (FMEA) methodology. The results of this analysis show generally that the impactor shaft sinking system is safer than conventional shaft sinking. The potential risk of conventional shaft sinking operations could be reduced by an average of over 35 percent with use of the impactor shaft sinking system. Design change recommendations to further improve the safety of the impactor system include a revised fire suppression system, improved access to and from the operator cabs, reliable noise attenuation measures, and improved communication systems. An operational safety manual was also developed for the impactor shaft sinking system.

Jan 1, 1982

By Pablo Velasco

After becoming an oil exporter in 1972, Ecuador experienced a remarkable economic upswing until 1978. However, in 1979 and 1980, a combination of only slight improvements in agricultural output and stagnating petroleum output lowered rates of real gross domestic product (GDP) expansion to 5.3% and 4.3%, respectively. In current prices, the GDP in 1980 was estimated at $11.3 billion.2 The slowdown in growth is attributable in large part to the low level of petroleum exploration in the 1972-79 period. Ecuador's production of petroleum declined 4.5% in 1980 compared with that of 1979. The consortium of Corporación Eatetal Petrolera Ecuatoriana (CEPS), the State oil company, and Texaco, Inc., operating in Ecuador's eastern jungle region was responsible for nearly all of 1980 production, The volume of total crude petroleum exports from Ecuador dropped from 44.7 million barrels in 1979 to 40 million barrels in 1980, although crude petroleum export revenues rose from $1,032 million in 1979 to $1,294 million in 1980- The latter figure represented 54% of Ecuador's 1980 merchandise exports. Crude oil exports accounted far about 46% of the value of 1979 merchandise exports. The total nonfuel mineral exports in 1980 were valued at less than $1 million.

Jan 1, 1982

By Gary W. Kramer

The Bureau of Mines has developed a rapid analytical method for the determination of anions by ion chromatography (IC) in phosphate rock mining materials and cement kiln dust (CKD). The procedure employs an Na2C03 fusion of the sample, followed by a water extraction that allows direct injection of the sample into an ion chromatograph without separator column poisoning by the transition metal ions. Fluoride (F-) and phosphate (P043-) were determined in 24 phosphate rock processing and waste material samples by the IC method. The results were compared with a pyrohydrolytic distillation-titration method for F-and a colorimetric procedure for P043-. Excellent analytical agreement was obtained for F-and p043-values determined by IC and the comparable wet chemical methods. Fluoride values ranged from 0.3 wt-pct in the filter cake to 3.6 wt-pct in phosphate rock concentrate. Phosphate values ranged from 1.2 wt-pct in the filter cake to 33.8 wt-pct in the rock concentrate. Fluoride, chloride, phosphate, nitrate, and sulfate in several CKD samples were determined by the IC method and standard wet chemical techniques. Excellent agreement was obtained between the two methods for all anions. A total of 113 CKD samples from all areas of the United States was analyzed by the IC method. Fluoride values ranged from <0.01 to 0.36 wt-pct. Chloride values ranged from <0.01 to 12.5 wt-pct. Sulfate values ranged from 0.41 to 31.6 wt-pct. Phosphate and nitrate were detected in only about 25 pct of the samples at >0.02 wt-pct. Maximum phosphate and nitrate values were 1.0 and 1.6 wt-pct, respectively.

Jan 1, 1982

By Kadri Dagdelen

Design of the ultimate pit limit to direct long range mine planning has been a celebrated problem in the mining industry, and in mining academia, since 1965. Although many papers have been presented on this topic, even today, there is not a single solution technique which is widely accepted and considered universally best. Many philosophical differences exist among the pit planners concerning the use of different methods and it appears that resistance is very strong towards accepting other techniques if you are a believer in one. Whatever the case might be, the pit parameterization technique for the ultimate pit limit design is the latest and takes a different approach than its predecessors. There are already a number of papers which discuss the concepts involved (Francois-Bongarcon, 1976; 1978; 1980; 1981); however, the aim of this paper is to fill in some of the gaps which are thought to be missing in the previous discussions and, in addition, to present the recent research done on the subject at the Centre de Geostatistique de l?Ecole de Mines de Paris in Fontainbleau, France. This paper is divided into two parts. The first part will discuss practical aspects and interpretations of pit parameterization concepts with- out using mathematical or theoretical arguments; the reader who is interested in the theory should refer to Francois-Bongarcon (1978). The second part will present an approach to obtain a complete double parameterization of reserves using an inexpensive (computer costs) single parameterization algorithm. A case study of this approach is also included.

Jan 1, 1982

By C. R. Edwards, J. R. Smith

"INTRODUCTIONThe purpose of this paper is to give an overview of the ALSANDS PROJECT, with emphasis on the approach taken in the design of the Primary Extraction and Froth Treatment Plants.The Alsands Consortium (Table 1) is a Joint Venture of 9 companies formed in 1978, with its prime objective to obtain governmental approval and concomitant suitable commercial terms to build and operate a third oil sands mining and upgrading facility in the Athabasca Oil Sands Deposit.Alsands Energy Limited is the Consortium Manager.The Alsands Project will provide about 10 percent of Canada's forecast crude oil production by 1990, which will be 20,700 cubic metres per day (130,000 barrels/day) of synthetic crude oil plus 1,100 cubic metres per day (7,000 barrels/day) of liquid petroleum gases. To construct Alsands the time required by construction labour, field non-manual and engineering support personnel as well as those trades involved in project-related urban development will total approximately 100,000,000 manhours or 50,000 man years. During regular operations, 4,800 people will be employed by Alsands.Approximately $13 billion will be spent to design, construct and start up the plant by the scheduled 1988 completion date."

Jan 1, 1982

By Jack V. Hill

A unique type of grinding mill liner has been developed by Waagner-Biro A.G. of Austria. This liner system, termed an angular spiral lining (.ESL), was, successfully tested by Magma Copper Company&apos;s San Manuel concentrator in both their 10 Ft x 10 Ft (3.05m x 3.05m) and 12 Ft, 6 In x 14 Ft (3.81m x 4.27m) grate discharge wet grinding ball mills. Magma Copper&apos;s savings in power and grinding balls justified conversion of more than half of their twenty-six ball mills to the ESL system.1,2 The ASL system was provided by the Capitol Castings Division of Midland Ross Corporation, the exclusive licensee for such conversions. Capitol Castings proposed the following as expected benefits from installation of an ASL system in an 18 Ft x 21 Ft (5.49m x 6.40m) overflow discharge wet grinding ball mill at the Pinto Valley Operation of Cities Service Company: "(a) Reduction of power consumption per ton of ore milled = 12%, (b) Increased throughput = 6%, (c) Reduction of ball consumption = 6%."

Jan 1, 1982

By Joseph W. Leonard

Partition (Tromp) curves, as used in diagnosing problems in coal cleaning equipment, are discussed and classified using five arbitrarily chosen types to represent, on a scale from one to five, system performance. Causes of unsatisfactory Tromp curves for various cleaning units, ranging from operational inadequacies to problems best corrected by the equipment builder, are analyzed.

Jan 1, 1982

By David C. Oyler

Three long horizontal holes were drilled from a directionally drilled surface hole at the Emerald Mine near Waynesburg, Pa. The purpose was to adapt the technique of directional drilling for use in draining methane from coalbeds. A 504-m (1,652-ft) long, 76-mm (3-in) diameter, circular arc pilot hole was drilled, using a downhole mud motor, to enter the Pittsburgh coalbed at a vertical depth of 305 m (1,000 ft). The hole was reamed to 222 mm (8-3/4 in) and was cased to a measured depth of 486 m (1,595 ft). Three 76-mm (3-in) diameter horizontal drainage holes were then drilled, totaling 2,909 m (9,544 ft) of horizontal hole. Improvements in drilling methods increased the average drilling rate from 20.4 ill (67 ft) to 64.9 m (213 ft) per day. The cost of drilling the directional and horizontal holes (using a Government-owned drill rig) was $1,169,530, a figure inflated by inexperience and by delays caused by lost-circulation problems and fishing operations. An estimated total drilling cost of $960,000 (including rental of a drill rig) for an improved system, was determined by a detailed cost analysis. Initial gas and water production from November 1979 through May 1980 was low because of caving of the horizontal holes drilled in shale near the bottom of the casing.

Jan 1, 1982

By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

By Edgar S. Burkhardt

In mining, the process of comminution starts at the pit and ends at the classifiers. The first machine that breaks or comminutes ore is the primary crusher. There are many types of crushers, but the one most commonly used in large tonnage plants is the gyratory so discussion is restricted to this type of machine. See Figure 1. The primary crusher is the most important piece of comminution machinery in the plant because every ton of ore processed by the concentrator must pass through it, as shown in Figure 2. Most plants can operate if a secondary crusher or a mill is shut down, but if the primary crusher stops, the whole concentrator goes down. It is, therefore, important that the machine be able to produce sufficient ore to keep the concentrator running without hesitation. The plant designer first considers the required daily capacity and an estimate of the size of the ore from the pit. To this he adds an availability factor to compensate for delays due to shovels, trucks and rou- [ ] tine maintenance at the crushing station. This factor is based on experience and can be on the order of 75% (1). With these figures, he determines the hourly capacity required of the primary crushing station to keep the concentrator supplied. See Table I.

Jan 1, 1982

By Kelvin K. Wu, Daniel S. Mazzei

As a result of the 1972 Buffalo Creek disaster, the Mining Enforcement and Safety Administration (MESA), now the Mine Safety and Health Administration (MSHA) initiated the draft of new federal regulations to deal with the mine waste impoundment and refuse pile problems. Public hearings were held and finally promulgated in October, 1976. As stated in 30 Code of Federal Regulations 77.214-216, the complete engineering design plan which is mandated by the regulations has to be approved by a coal mine health and safety district manager. During the past 18 months MSHA has gained considerable experience through the submission, review, and approval process. This report indicates the procedures, general design criteria and inspection requirements for submission of engineering plans to be approved. In particular, problem areas encountered most frequently by both operators and MSHA will be discussed. Only through diligent efforts by the operators, consultants and government, can we hope to accomplish the common goal to improve the safety of operations.

Jan 1, 1982

By James R. Walpole

This paper will discuss the federal laws and regulations that are related to water quality. Both surface and underground water quality issues will be covered as well as point and nonpoint source requirements. Some of the subjects to be covered include the Federal Water Pollution Control Act (e.g., effluent guidelines, nonpoint source requirements, etc.), the Safe Drinking Water Act (e.g., underground injection control requirements; pits, ponds, and lagoons study, etc.), the Resource Conservation and Recovery Act (e.g., mining wastes study, etc.). The importance of related court decisions will also be discussed

Jan 1, 1982

The Government of New South Wales has granted approval for Hunter Valley Aluminium Pty., Ltd" and Alcan Australia Ltd. to proceed with the building of an aluminum smelter at Lochinvar and the expansion of a smelter at Kurri Kurri. Construction of the smelters is expected to start as soon as the consortium is finalized in early 1982. The smelters are part of a plan to increase Australia&apos;s aluminum capacity fivefold to 1.5 million tons by the mid-1980&apos;s. CHILE Construction of a $61 million lithium carbonate plant in northern Chile will be started in early 1983 by Sociedad Chilena del Litio Ltdao Production is scheduled to begin in 1984 with an initial capacity of 6,350 tons of lithium carbonate produced from a brine deposit. The firm is owned 55% by Foote Mineral Company, of Exton, Pennsylvania, and 45% by Corporacion de Fomento de la Produccion-CORFO.

Jan 1, 1982

By McCarthy PL

Widespread interest in reopening Victorian gold mines, regenerated by increased gold prices, has highlighted the technical problems prevailing. A combination of historical lease legislation, minimal recent mining activity, forms of mineralisation, high insurance rates and tortuous approval requirements for mining mean the potential investor assessing an old mine must address a wide range of technical problems. The requirements for establishment of ore reserves and attraction of capital for mine development are fundamental. Underground reef mining was historically generally on a small scale, and as mining on many fields ceased about the turn of the century, past geological and production records cannot be correlated to personal experience and comment. The impact of modern mining technology on future reef mining will be to reduce operating costs and significantly improve historic metallurgical recovery. Roth these are commensurate with increased capital costs. Workers compensation rates currently applicable in Victoria are many times higher than in other states, adding significantly to operating costs. Establishment of a pilot plant testing facility in central Victoria would provide an invaluable service in treating bulk samples essential for grade and recovery assessments of Victorian ores.

Jan 1, 1982