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The Federal Coal Mine Health and Safety Act of 1969 limits to 2 milligrams per cubic meter (for dust containing a maximum of 5% quartz) the average respirable dust concentration to which individual miners may be exposed. Section 205 of the Act directs that an appropriate exposure formula for dusts with quartz content in excess of 5% be established. This formula, entered in F. R. Document 70-12351 filed September 16, 1970, is 10 milligrams per cubic meter divided by % quartz. Thus, the total respirable dust exposure permitted limits the quartz to a maximum of 100 micrograms. For example, if % quartz is 5%, 10 mg/5 = 2 milligrams, and (.05) (2 mg) = 100 micrograms. Similarly, if % quartz is 10%, 10/10 = 1, and (.10) (1 mg) = 100 micrograms, etc. If the % quartz is less than 5%, the 2 mg/M3 limit prevails.

Jan 1, 1974

Two different techniques were used to make measurements of the absolute value of electromagnetic noise in an operating coal mine, Robena No.4, located near Waynesburg, Pennsylvania. One technique measures noise over the entire electromagnetic spectrum of interest for brief time periods. With present instrumentation, the spectrum can be covered from 40 Hz to 400 kHz. It is recorded using broad-band analog magnetic tape, and the noise data are later transformed to give spectral plots. The other technique records noise envelopes at several discrete frequencies for a sufficient amount of tine to provide amplitude probability distributions. The specific measured results are given in a number of spectral plots and amplitude probability distribution plots. The general results are that at frequencies below 10 kHz, power line noise within the mine is severe. Impulsive noise is severe near arcing trolleys, and at lower frequencies near any transmission line. Carrier trolley phone signals and harmonics are strong throughout the mine whenever the trolley phone is in operation. Additional information beyond that included in this report may be obtained from the authors, who are with the Electromagnetics Division of the National Bureau of Standards.

Jan 1, 1974

By Jose M. S. Ramos

This paper describes the slope stabilization procedure at Cauê Mine, Companhia Vale do Rio Doce - CVRD, Itabira, State of Minas Gerais, Brazil. Cauê Mine, is a large open pit, to produce about 41 million tons of iron ore during 1974, and an intensive program of investiga Lion was carried out to provide the basic data for stabilizing its slopes, The ultimate rock slope in the W area will be about 400 m high with dipping angles decreasing gradually from top to bottom of the pit. The investigations started in 1969 and since then many geologists and engineers, from both CVRD and Geotécnica S/A., assisted by international consultants, have given their contribution to the achievement of solutions. It is hoped that this paper may contribute to the slope stability of other mines in tropical climate.

Jan 1, 1974

An outline of the circumstances giving rise to the current Commonwealth-States con- flict in relation to offshore mineral explor- ation and production is set out, together with an analysis of the Seas and Submerged Lands Act 1973. The bases of the claims by the Commonwealth and the States in this regard are discussed with relation to the territorial sea and the continental shelf.

Jan 1, 1974

By Robert D. Coale

This paper will describe the copper concentrate slurry pipeline and pumping system between the concentrator and the port at Bougainville Copper Ltd. The operations of Bougainville Copper Ltd. are located on the island of Bougainville in the South Pacific. Geographically, it is the northernmost island of the Solomon Group but, politically, it is part of Papua New Guinea. Papua New Guinea is a United Nations Trust territory administered by Australia, having recently been given self-government, with independence scheduled for December 1974. (Fig. 1). The island is approximately 130 miles long by 30 miles wide, with the Crown Prince mountain range forming a spine along its length. The mountains in place are over 8,000 feet high with the mine located on the western flank at an elevation of 2,200 feet. Exploration on Bougainville for copper by Conzinc Riotinto of Australia (CRA) began in 1963 and by 1968, through geophysical, geochemical and diamond drilling methods, an orebody containing nearly one billion tons of ore grading approximately 0.48% Cu. and sane gold had been outlined. Bougainville Copper Ltd., the company operating the mine, is 53.6% wined by CRA, 20% mined by the government of Papua New Guinea with the public Folding the remainder of the shares. The first copper concentrate was produced in early 1972, with production for the first full year, 1973 being:

Jan 1, 1974

By P. T. Brooks

A process for extracting 69 percent of the vanadium from weathered Nevada dolomitic shale containing about 1 percent V205 was developed on a laboratory scale by the Bureau of Mines. The process essentially is comprised of salt roasting, leaching with dilute sulfuric acid, solvent extraction, and precipitation of vanadium as ammonium metavanadate. Carbonaceous unweathered shale was less amenable to treatment. A preliminary cost estimate for mining and processing 1,500 tons per day of weathered shale indicated a production cost of about 1 dollar per lb of V205. Considering recent sales at nearly $1. 75 per lb of contained V205, larger scale testing to obtain more definitive process design and evaluation data may be warranted.

Jan 1, 1974

By D. D. Porter, H. G. Carlevato

Blasting to prepare orebodies for in situ leaching usually involves extraordinary conditions which sometimes require special and innovative techniques. To date several unusually large blasts have been made for leaching with results that are very encouraging. Among the factors considered when designing these large blasts are orebody shape and depth, rock properties, vibration and damage potential, loading logistics, and safety. All of these factors, and others, are discussed with an effort made to put each of them into its proper perspective.

Jan 1, 1974

By O. E. Wheeler

The term "finite element analysis" is heard and seen in print a great deal these days; a few words on what precisely it is are in order. In the broadest sense, finite element analysis is an application of an idea developed by a nineteenth century mathematician named Rayleigh. His basic idea was to decompose any problem in mathematical physics into a large number of simpler problems, solve the simpler problems, and then put those solutions together to form an approximation to the solution of the original problem. The scheme works nicely because of two features: first, much of the difficulty in solving practical problems of mathematical physics stems from irregularities in the boundaries and the simpler problems can be chosen with regular boundaries and second the smaller problems are solved for small physical areas where simple approximations of the solution do not differ much from the true solution. In the narrower sense, in which we are now interested, finite element analysis is a technique of using imaginary simple structural elements to construct analytical models of real structures. A direct parallel with the Rayleigh approach exists since the "finite elements" have simple boundaries and are restricted to simple behavior within a single element. Although the correspondence between Rayleigh's ideas and finite element analysis is now widely recognized, the computational tool, as it has developed over the last thirty years, comes not from the discipline of physics but rather from that of structural engineering. It is interesting that it came not from the traditional research community, the universities, but instead from working

Jan 1, 1974

By G. L. Hundley

An anhydrous chlorination process has been developed by the Bureau of Mines for the production of copper, ferric oxide, and sulfur from chalcopyrite. The first part of the process is a chlorination step in which elemental sulfur and a mixture of iron-copper chlorides are produced. This paper describes the second step of the process in which the copper and iron chlorides are reacted with oxygen to selectively convert the iron chloride to ferric oxide and chlorine leaving the copper in the chloride form. The chlorides are pulverized and dropped countercurrent to a flow of oxygen in a vertical 20-cm-diam. reactor at 800° C. By leaching with brine, the copper chloride may be separated from the insoluble iron oxide and reduced to copper powder by electrowinning. It has been demonstrated that 70 to 80 pct of the iron can be converted to oxide in a single-stage operation with a chlorine content in the offgas of 80 pct. After removal of the oxygen, this chlorine could be recycled to the chlorination step. A two-stage operation will convert 98 to 99 pct of the iron in the mixture to insoluble oxides. The offgas from the second stage contains 70 pct chlorine.

Jan 1, 1974

By James A. Doolittle

In the underground type of mining operation, the finer portion of the waste material commonly is pumped as a hydraulic slurry to a tailings disposal pond. (1)* The tailings slurry is deposited via a pipe manifold system located around the perimeter of the pond. This method of deposition produces zoning because the coarser material tends to settle out immediately while the finer particles tend to remain in suspension and are deposited progressively nearer the center of the pond. The manifold system must be raised periodically as the pond fills with the tailings material. Over a period of time, this procedure will produce a zoned embankment known as a tailings dam. A typical pond and dam are shown in Figure 1. In plan view, the dimensions of tailings dams can be on the order of miles, with the height of the dam reaching 100 to 200 feet. Due to such factors as inadequate compaction, inherently low strength of the tailings material, and the possibility of high phreatic surface, many dams at some point become unstable, i.e., the factor of safety against a slope failure approaches 1.0. Remedial measures are required to improve the safety factor in order to prevent massive failure of the dam. In most cases, stability can be improved by effectively flattening the slope of the dam as shown in Figure 2, but this is not possible in every case. Occasionally, a tailings dam is constructed adjacent to a natural or legal boundary so that it is not possible to construct a conventional berm or to reduce the slope angle with earth moving equipment. Even a terracing process would require removal of the material and this may not be practical. In such a situation, an alternate scheme to improve stability must be found.

Jan 1, 1974

By J. T. Crawford

The purpose of this paper is two fold. It will: (1) examine leasing analysis as a project evaluation criterion, and (2) compare the cost of leasing with other methods of financing. The use of leasing has grown extensively in recent years as a method of obtaining equipment. The computer industry has been highly dependent upon leasing from its earliest days to dispurse its equipment to the customer, while airlines and railroads acquire large segments of their aircraft and rolling stock by leasing. Over the last few years the mining industry has leased haulage trucks, bulldozers and other equipment of similar nature.

Jan 1, 1974

By F. M. Carlson

Comparisons are made between the results of the Buckley-Leverett and Kern solutions, which describe immiscible displacement of oil by gas in a linear system and the results obtained from a less restrictive numerical model. This model takes into account the effect of compressibility of the oil and gas and the effect of allowing changes in the amount of gas in solution in the oil. These effects are best exemplified by imposing large pressure drops across the system. With these large pressure drops, comparison of the model solution with the Buckley-Leverett solution reveals that the Buckley-Leverett solution provides estimates of recoveries at breakthrough of the gas that are too low. In contrast to the model, the Kern solution predicts breakthrough recoveries that are too high. A modification to the Kern solution is presented to facilitate more accurate predictions of recovery at gas breakthrough for gas injection at pressures higher than the saturation pressure of the oil.

Jan 1, 1974

The observation of lower-than-predicted Fe/Cu ratios in the cementation of copper on iron in tests at elevated temperatures indicated that some other reaction was contributing to the reduction of cupric sulphate in addition to the reduction by elemental iron. Experiments showed that the copper sponge, which initially formed on the iron, subsequently reduced the cupric sulphate giving the products cuprous oxide and sulphuric acid.Tests were undertaken to determine the rate and equilibrium conditions of the reduction reaction with variations in the types of the reducing agent and solution parameters. The reaction first became apparent at temperatures above 50°C and as it ceased at pH 1·8 it was possible to induce further reaction by alkali addition.

Jan 1, 1974

1. Domestic Production and Use: Twelve companies operated 31 primary aluminum reduction plants, three firms accounting for 66% of production. Washington, Oregon, and Montana accounted for 32% of the total production; Gulf Coast States, 25%; Northern States, 29%; and other Southern States, 14%. Output of primary metal in 1973 was valued at $2.2 billion. About 80% of the secondary recovery was from new aluminum base scrap and 20% was from old scrap. Aluminum consumption by 25,000 firms was centered in the East Central United States. Of aluminum consuming industries in 1972, building accounted for 28%; transportation 19%; electrical, 13%; packaging, 16%; and consumer durab1es, 10%.

Jan 1, 1974

By Robert Stefanko

Test procedures for evaluating underground coal mine trailing cable splices are formulated. The tests are utilized to evaluate several trailing cable splices. The laboratory results are compared with actual field observations and experience. The results are utilized to define quantitatively the generalized terms of the 1969 Health and Safety Act. Recommended specifications and procedures for testing by MESA Department of Approval and Testing personnel are included. Failure points in splices are detailed and presented for use by a MESA training program and manufacturers attempting to improve product quality.

Jan 1, 1974

This three-part report provides the results of an industrial engineering study that assessed the safety and efficiency of underground ore handling in U.S. metal mines. One of the primary findings of the study was the improvement needed in controlling muck size in the are handling stream. Oversize ore is causing much additional labor, creating work stoppages, and is a principal cause of accidents in metal mines. The specific objectives of the Present Systems Study were to determine and evaluate the productivity of existing ore handling systems in a representative group of underground metal mines, to relate the available historical accident record statistics of these selected mines to the specific operations in the overall ore handling operations, and to highlight the potential areas of improvement for each mine on the basis of productivity and associated hazardous index for each operation.

Jan 1, 1974

By M. D. Aldridge

The results of a four-year research effort aimed at developing a practical and effective electronic system for monitoring safety conditions throughout a coal mine by displaying, at a central location, remotely measured parameters are described. This document presents the general philosophy of an experimental system in an operating coal mine. The conclusions drawn from this experience are presented and a typical design for a large coal mine is described.

Jan 1, 1974

By Roger J. Morrell

Bureau of Mines laboratory experiments with reduced scale cutter and cutter forces were performed to (1) define the fundamental relationships governing disk cutter performance and (2) to develop a method of predicting disk cutter performance in soft, medium, and hard rocks. The rock cutting experiments were performed in four igneous and metamorphic rocks that ranged in compressive strength from 38,000 to 67,000 psi. A special testing machine called a linear cutter apparatus (LCA) was used to load and traverse a free rolling disk cutter across a smooth rock surface. Each pass of the disk produced a single crater, and runs were spaced to avoid breakage between craters. During each run the normal and tangential cutter forces were measured and recorded. The soft and medium rock data were obtained from previous Bureau experiments with disk cutters. Crater depth, tangential cutter force, crater volume, specific energy versus normal cutter force, and crater width versus depth are defined within this study. With the use of standard physical properties, performance prediction equations were developed using stepwise linear regression analysis techniques. A method of using these prediction equations to calculate the boring rate, the cutter spacing, and the torque and energy requirements of a full-scale boring machine was also developed.

Jan 1, 1974

By R. L. Cosgriff

A commercial, battery-powered mine vehicle has been equipped for remote operation in coal mines. This vehicle is controlled by an operator from the remote station. Command signals originating at the remote station as well as video and data signals originating on board the vehicle are relayed between the remote station and vehicle using a single coaxial line. This coaxial line is normally 2,000 feet in length and is reeled in and out from a drum mounted aboard the vehicle The vehicle has a maximum speed of 6 ft/sec and a minimum speed of 4 ft/sec using 36-volt excitation of the traction motor. If this traction motor is excited by two of the three batteries, 24-volt excitation, the vehicle will travel at approximately 2 ft/sec. The speed can be reduced by the operator to 1 ft/sec. This vehicular system has been tested in coal mines. Presently, the major limitation of this system is associated with fatigue failure of the splices in the coaxial line and lack of mobility caused by its 3-wheel design.

Jan 1, 1974

This report provides the results of an industrial engineering study that assessed the safety and efficiency of underground ore handling in U.S. metal mines. One of the primary findings of the study was the improvement needed in controlling muck size in the ore handling stream. Oversize ore causes much additional labor, creating work stoppages and is a principal cause of accidents in metal mines. This report includes data on block caving, back cut and fill room and pillar, shrinkage stoping, and sublevel stoping.

Jan 1, 1974