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By Oliver Bowles

"Underground operation.In its broader sense the term ""mining"" is applied to all processes whereby raw materials are taken from the earth to be used as sources of mineral products. In a more restricted sense ""mining"" is applied to underground operations as contrasted with open pit work designated as quarrying. The tern is used in the restricted tense in this paper; for it relates to the exploitation of limestone deposits by modern methods of underground operation rather than by open pit methods.Present tendency toward miningIt may seem to some readers that a paper on mining limestone will find little general interest among lime plant operators, for only a limited number of them have yet resorted to this means of procuring raw materials. It is noteworthy, however, that a growing tendency exists toward underground operation, for one by one quarry owners are faced with increasing difficulties in open pit operation. It seems desirable, therefore, for operators to keep closely in touch with this gradual change in process, to inquire into the underlying causes, to compare costs, and to study the effects of the change in method on subsequent steps in calcining the stone to produce lime."

Mar 1, 1924

By R. S. Dean

"Interest in the nitrogen content of steel has increased sharply in the last few years. Investigations have indicated that its presence in steel has an important bearing on high temperature brittleness3, magnetic properties4, room temperature hardening5, and shock resistance. The steel producer may therefore look forward with certaintly to the necessity of controlling this element in a considerable proportion of his product. The literature, however, reveals a great many contradictory statements concerning the percentage of nitrogen which must be present to produce given effects on the properties of iron and steel. Most of these arise from a misunderstanding of the nature of the iron-nitrogen diagram near the iron end and a lack of recognition of the effect of state of aggregation on physical properties although both have been fairly well worked out.My thanks are due to Mr. J. L. Gregg, Prof. R. H. Dowdell, Mr, R. F. Mehl and Mr. H. C. Vacher for suggestions in connection with the preparation of this paper, and to Mr. C. H. Hefty and Mr. G. T. Motok for furnishing a valuable bibliography and translations of articles by Must and Duhr.The iron-nitrogen diagram in the solid state, so far as it concerns the nitrogen content of commercial iron and steel, consists simply in a solubility curve increasing with temperature. The important thing about this solubility curve is that without cold wore the iron nitride when precipitated does not go into solution on heating for long periods below 300°C. This slowness of solution is probably due to the fact that iron nitride when present as a separate phase has at these temperatures an equilibrium pressure of several thousand atmospheres7 and that therefore each particle of iron nitride surrounds itself with a layer of nitrogen. The precipitation of such an unstable crystal presents points of interest thermodynamically but we can not go into these here. It is quite well established that it precipitate. It can not diffuse then unless this layer is broken through or the temperature and pressure are increased to a point where the nitrogen is absorbed at an appreciable rate in the surrounding iron. This rate will depend on the degree of unsaturation of the iron and the excess pressure of the nitrogen above equilibrium pressure. The degree of unsaturation is determined by the temperature alone so that the temperature at which the rate of solution of the nitride becomes appreciable will depend on the pressure of nitrogen developed. This, in turn, will depend on the balance between rate of decomposition of the nitride and the rate o: flow of the iron. In annealed iron, this temperature of appreciable solution rate is between 300 and 350°C.8 By cold work, however, the elastic limit of the iron is increased, and conditions for the nitride to go into solution are reached at a much lower temperature, giving rise to the so-called blue brittleness. This brittleness and hardening is due to iron nitride going into solid solution and is not precipitation hardening."

Mar 1, 1931

By W. H. Ryback

When a fire or explosion occurs in an underground coal mine, the atmosphere guickly becomes irrespirable. Highly irritant smoke and toxic levels of carbon monoxide and dioxide are produced, and consumption of oxygen by the fire or blast can reduce its concentration below that needed to support life. Work parties entering such a hazardous atmosphere for rescue or recovery operations are required to use, as their primary respiratory protection, a fully self-contained breathing apparatus of at least two hours duration. Apparatus used in this role has primarily been of the closed circuit rebreather type, using pressurized oxygen storage and chemical removal of expired carbon dioxide. Typically, commercial self-contained breathing devices for mine rescue have been heavy, bulky, and uncomfortable to use. The McCaa apparatus, long the industry standard, represented the best art available when it was First designed in 1926. It weighs 41 Ibs., protrudes about six inches above shoulder height, and provides breathing gas at temperatures up to 120° F. More recent designs, such as the Drager BG174A and Scott Rescue-Pak, using modern light weight materials and fabrication techniques have improved on all parameters of the McCaa to some extent. It was the intent of the Bureau of Mines, acting under the authority of the Coal Mine Health and Safety Act of 1969, to fund the development of a new self-contained breathing apparatus which would incorporate the best available technology and serve as a benchmark for future commercial designs, as the Bureau-sponsored McCaa had 45 years previously.

Jan 1, 1974

By S. D. Hill

To help insure the continued viability of the minerals and materials sector of the U.S. economy, the Bureau of Mines is conducting research and disseminating technology that will result in more efficient minerals extraction. As part of this effort, a process is being investigated that involves an aqueous chlorine-oxygen leaching procedure to treat complex sulfide concentrates to produce ZnCl2. Subsequent fused-salt electrolysis of the ZnCI2 produces high-purity zinc metal. Previous reports described preliminary, small-scale research on the chlorine-oxygen leaching system and fused-salt electrolysis of ZnCl2. This report discusses the details of design, construction, and longterm, continuous operation of both monopolar and bipolar fused-salt cells. The effects of temperature, electrode gap, current density, electrolyte composition, and electrode configuration on current efficiency and energy consumption were determined for a ZnCl2-KC1-NaCl electrolyte system. Zinc metal was produced at the rate of about 50 lb/day in a monopolar cell which was operated continuously for 60 days with current efficiencies from 85.8 to 98.6 pct and an average power consumption of 2.1 kwhr/lb. Bipolar cells with several different electrode configurations were operated for 90 days. A current efficiency of about 70 pct and power consumption of about 1.7 kwhr/lb of zinc were achieved using a three-electrode-pair, bipolar assembly with a chlorine gas lift pump to circulate electrolyte.

Jan 1, 1981

By J. J. S. Sebastian

The process development described in this report was carried out at the Morgantown Station of the U. S. Bureau of Mines in cooperation with West Virginia University. An experimental laboratory-scale pilot unit to gasify coal dust in entrainment was constructed on the Campus of the University in Morgantown. The enormously accelerated demand for liquid and gaseous fuels and utilization of our immense coal resources require that a low-cost process be developed for the gasification of a wide variety of coals. Such a process would be a major factor in lowering the cost of synthetic liquid fuels. None of the proc3sses developed in Germany prior to and during World War II, such as the Winkler process, the H. Koppers dust gasification process, the Wintershall-Schmalfoldt process, etc., meet all the requirements for a low-cost process under American conditions. Nor has any other process designed and experimentally operated at atmospheric pressure during the post-war period in this country appeared to fulfill all of the necessary conditions for a low-cost process.

Jan 1, 1950

By E. S. Chalpin

On 25 August 1971, the U.S. Bureau of Mines contracted with the Aerojet Liquid Rocket Company to perform an engineering research program to be completed on 25 May 1972. The program objective was to evaluate the advantages and disadvantages of the underground use of several power sources including but not limited to a closed cycle diesel, electric cable/motor, rechargeable battery/ motor and exhausting diesel for mine shuttle cars, with emphasis on health and safety. This report, Research in Advanced Power Systems for Mining Health and Safety consists of four (4) main parts: (1) Hazards and Safety Analysis of Mine Power Systems, (2) Evaluation of Closed Cycle Diesel Power System for Mine Applications, (3) Closed Cycle Diesel Power System Consumables Compatibility, and (4) Concluding Remarks and Recommendations. The hazards introduced to a mine by shuttle car power system failure modes were evaluated and reported. A "fault tree" analysis was generated for each power system and the probability of death or injury during an eight (8) hour shift and for a week of operation (105 hours) were estimated. Advantages and disadvantages were also reported for each system studied. Aerojet's assessment, based on a "forced decisions analysis," of the four shuttle car power systems as to their relative health and safety is that the exhausting or open cycle diesel is preferred, followed by the closed cycle diesel, the electric cable/motor and the rechargeable battery/motor. The assessment was based on ten parameters pertinent to the health and safety aspects of shuttle cars. Drawing layouts of two typical mine shuttle cars incorporating closed cycle diesel powerplant concepts are included in this report. The performance parameters established for the two concepts are tabulated. A study was made of the logistics involved in transporting and replenishing the consumables needed to operate a shuttle car equipped with a closed cycle diesel. The closed cycle diesel powerplant consumables -oxygen, potassium hydroxide, and diesel fuel - were analyzed as to their compatibility with materials found normally in a gassy coal mine. Spill tests simulating inadvertent leakage of liquid oxygen in a coal mine were performed. A critical evaluation of existing oxygen sensitivity data and the relationship to the use of oxygen on a shuttle car in a coal mine environment is presented.

Jan 1, 1972

By Robert L. Lagace

This final report documents the work done by Arthur D. Little, Inc. (ADL) on behalf of the U.S. Bureau of Mines, Pittsburgh Mining and Safety Research Center (PMSRC), on Contract H0122026 (which began in August of 1971). Under this contract ADL provided technical assistance to the Bureau on a task basis on virtually all aspects of the Bureau's programs related to present and planned emergency and operational communications and miner location systems for underground coal mines. The work consisted of independent investigations, analyses, experiments, breadboard and prototype hardware development, workshops and technology transfer seminars on mine communications, and on-going evaluations and guidance related to the Bureau's contracted programs on electromagnetic noise, mine communications systems, and trapped miner location. This final report documents the work in two volumes, Volume I, "Emergency and Operational Mine Commun¬ications," and Volume II, "Seismic Detection and Location of Isolated Miners." The Tables of Contents of both Volumes are included in each Volume.

Jan 1, 1974

By Mark Sanders

This report compares a proposed 10/60 "piggyback" 02 generating self-rescue unit to the standard CO self-rescue unit with respect to work task interference. A simulated low seam coal mine was constructed and subjects performed locomotion, shoveling, cart pushing, and timbering tasks with the 10/60 or self-rescuer attached to their belts. Task completion time was longer on all tasks when wearing the 10/60 unit, although only locomotion was statistically significant. The 10/60 unit was bumped over 100% more often than the standard self-rescuer. Implications for design of the proposed 10/60 container are: 1) unit?s container must be extra durable to withstand the added abuse it is likely to receive by low coal miners; and 2) extra consideration must be given to preventing the device from slipping on the belt while working, especially in low coal.

Jan 1, 1980

By William H. Schiffbauer

The Bureau of Mines has developed a passive encoder for range knobs, a device that automates the task of obtaining the range settings from devices or instruments to be input to microcomputers. It is easily constructed and installed and will not disrupt the device to which it is being attached. It consists of a custom knob and a small printed circuit board that has light-emitting diodes and magnetic reed switches attached. A ribbon cable connects it to a microcomputer. This encoder eliminates one source of human error in an automated process.

Jan 1, 1983

By B. T. Brady

Recent theoretical, experimental, and field studies conducted by the Bureau of Mines suggest that rock failures may be predicted and, in some instances, controlled. Theoretical studies that were later corroborated by laboratory and field studies predict that a seismic anomaly will precede rock bursts and/or roof falls in mines. The seismic anomaly is predicted to consist of an increase in the number of seismic events occurring in the immediate vicinity where the failure 1s nucleated. This increase will then be followed by a dramatic decrease of seismicity in a broad region that surrounds the location of the impending failure. The dimensions of this region allow a preliminary estimate of the time to failure (measured from the time the seismicity increase began) as well as an estimate of the energy that will be released at the instant of failure. Field observations indicating that this predicted behavior does precede rock bursts and a roof fall are presented in this report. Implications of results to the problem of failure control in mines that experience massive rock failures are discussed.

Jan 1, 1978

By James H. Rowland

The Mine Improvement and New Emergency Response Act of 2006 (MINER ACT) established a Technical Study Panel (The Panel) to provide recommendations on the utilization of belt air and new technology that may be available for increasing the fire resistance properties of conveyor belt used in underground coal mines. The Panel Report recommended use of the Belt Evaluation Laboratory Test (BELT) as the method for testing and approval of flame resistant conveyor belts used in underground coal mines. The research conducted to establish the correlation of the BELT with large-scale belt fire flammability tests was done using 36- to 42-in wide conveyor belt. Due to today?s coal haulage capacity, the mining industry is using 72-in and wider conveyor belts. The National Institute for Occupational Safety and Health (NIOSH) conducted a study to determine if the BELT will also qualify wider belts as fire resistant for use in underground coal mines. This paper describes the results of recent experiments comparing results from using the BELT and the large-scale tests for six different belts.

Jan 1, 2010

By M. A. Trevits, J. E. Urosek, A. C. Smith, M. P. Valoski

The National Institute for Occupational Safety and Health (NIOSH), in partnership with the Mine Safety and Health Administration (MSHA), is conducting research to test, evaluate, improve or modify coal mine fire-fighting strategies and methodologies through large-scale tests. Since wood is the second most abundant fuel available during a coal mine fire, it was decided that series of large-scale wood crib fire tests were needed to measure the products of combustion and to observe the capability of commonly available direct application fire suppression equipment, namely, fire extinguishers, water and gas-enhanced foam. This paper discusses the wood crib fire tests, provides insight into the products of combustion and describes observations made during the application of the fire suppression agents.

By Mark Colwell, Christopher Mark

This paper summarises the results of a research project whose goal was to provide the Australian coal industry with a rib support design methodology and software tool that could be utilised by suitably qualified colliery staff. The project was primarily funded by Australian Coal Association Research Program (ACARP) and further supported by several Australian longwall operations. The outcome of the project is a design methodology and software tool called, Analysis and Design of Rib Support (ADRS). ADRS is an empirical technique, which recognises that several geotechnical and design factors affect ribline performance and in addition that operational and safety issues essentially dictate the level of performance required. Therefore, the design recommendations associated with ADRS are specific to the Australian coal industry; however the procedure(s) for data collection and analysis could be applied to other countries' 'underground coal industries. Case history data were compiled from 34 longwall and two bord and pillar operations resulting in 204 case histories with each case history data set being defined by approximately 130 individual data fields. In addition monitoring (incorporating stress cells and extensometers) was undertaken at 10 collieries to assess and quantify the effectiveness of different rib support patterns and hardware on rib performance. The monitoring sites allowed for an improved understanding of the mechanisms of rib failure and degradation in terms of its interaction with the installed support and at the various stages of the mining cycle. The design methodology deals with both mains development and gateroad development specific to Australian longwall mines. This paper focuses on longwall gateroads subjected to abutment loading and in particular the travel road which becomes the tailgate of the subsequent panel. The statistical analyses associated with these cases suggested that the level of rib support should be based primarily on the development height and the pillar stress level. The ADRS software guides users through the design process, allowing them to develop rib support plans based on sound science and a broad base of in-mine experience To the best of the authors’ knowledge ADRS is the first systematic rib support design technique to be developed for any country’s underground coal industry. The development and use of empirical models in mining have substantially contributed to improving safety and productivity. ADRS further demonstrates that empirical techniques are particularly relevant and beneficial in dealing with the complexities of geotechnical design associated with underground coal mining.

Jan 6, 2008

By David Hill

The Australian underground coal mining industry has made extensive use of the Coal Mine Roof Rating (CMRR) classification system for a diverse range of purposes in recent years. These include mining method selection, and coal pillar and roof support design. This paper outlines a series of case histories, from large-scale feasibility studies to local support design investigations, that collectively illustrate the broad applicability, advantages, and usefulness of the methodology, as well as some of the current limitations. The key role of the CMRR in an overall hazard definition methodology is demonstrated for a major Australian project, and some ideas with regard to the future application of the CMRR, in the context of geotechnical risk management within a progressive, highly productive extractive industry, are put forward.

Jan 5, 2007

By Arnold O. Tanner

Tables based on 1985 data, which were not available in time for publication of the "1985 Minerals Yearbook," are included in this chapter; corresponding tables for 1986 are not yet available. Raw nonfuel minerals produced in the United States during 1986 had an estimated value of $23.5 billion, an increase of $0.25 billion over the 1985 value. This is the fourth consecutive year that the value has increased, and except for a de- crease in 1982, the value each year has increased since 1971, or 14 out of 15 years. Industrial minerals growth slowed from a 3% increase in value in 1985 to a 1.5% increase in 1986, reflecting the slow growth in the domestic economy. The value of industrial minerals, more than three times larger than the value of metals, easily offset the less than 0.5% decline in the value of metal mine production. Although the decline was small, this decrease makes 1986 the fourth year in the past 6 years in which a drop in the value of metal mining production occurred in the United States. U.S. production continued to lag behind the growing world production. Overall, the estimated raw nonfuel mineral production rose 1% over that of 1985. Domestic mine production of ferrous metals declined as follows, in decreasing order of production: nickel, iron ore, tungsten, and molybdenum. All remaining domestic primary nickel and tungsten mines had closed down by yearend. Nonferrous metals that similarly declined included titanium, rare-earth metals, lead, silver, zinc, and magnesium. For the second consecutive year, aluminum, produced predominantly from imported bauxite, declined in production, while imports reached record-high levels. Somewhat modest gains were made in beryllium and copper. As has been the case for a number of years, gold was a bright spot in the domestic and internation-al metal mining industry. Gold production was spurred on by the increasing usage and development of heapleaching techniques that make lower grade ores, which were previously uneconomical, profitable. Additionally, various mining companies, the U.S. Mint, individual States, and a number of foreign countries began production of gold bullion and commemorative coins. Gold output in the United States increased from 2.43 million troy ounces (revised) in 1985 to about 3.73 million troy ounces in 1986, a dramatic 54% rise.

Jan 1, 1988

This study of underground mine water inflow was initiated by the U. S. Bureau of Mines in March 1980. The objective of this study is to evaluate the hydrologic variables associated with surface water and ground water systems that affect the inflow of ground water into an underground mine. Data collection techniques are reviewed for their applicability for quantifying these hydrologic variables. Innovative techniques are also investigated. Existing analytical and numerical procedures for predicting underground mine water inflow are reviewed and evaluated. Over 300 computer models were reviewed for their applicability to modeling underground mine water inflow. The model UNSAT2 is utilized for modeling inflow to a shaft and tunnel or drift. This model and the techniques developed in this study are applied to two existing mines.

Jan 1, 1982

By Robert N. Roby

This report is one of a series based upon investigations made and being made within the Missouri River Basin in Montana by engineers of the Bureau of Mines , Albany Branch , Mining Division . These investigations are made primarily to provide preliminary information to Department of Interior agencies and others concerned with the planning of the proposed Federal power developments within the Missouri River Basin . The present report has been designed to permit public distribution of the factual data obtained . The field investigations in Meagher County were made during tually all 1948 ; Vir- known metallic mineral occurrences within the county were investigated ; some nonmetallic mineral deposits also were investigated . The investigations cannot be considered complete , however , because some long-inactive mines or prospects may have been missed . Most of the old mine and prospect openings are caved; many are completely inaccessible . Many of the claims located years ago have been abandoned ; others have been relocated under different names and ownerships . Because of such conditions , much of the information regarding their development and operation necessarily has been obtained from Federal and State publications , supplemented by information provided by owners and local residents .

Jan 1, 1950

By F. E. Cash

"Explosions in bituminous coal mines are preventable. The combination of adequate ventilation, thorough rock-dusting, and water on the cutter bars of mining machines and of empty and loaded cars will certainly reduce the hazard to a minimum, if not remove it entirely. This is the attitude taken by the Woodward Iron Co. in lessening or removing the dust and gas explosion hazard from its Dolomite and Mulga mines.This report was made possible through the willingness of Frank H. Crockard, president of the Woodward Iron Co. to make records available and to authorize their publication for the benefit of the coal-mining industry.The compilation of these data and preparation of maps was facilitated by the cooperation of M. Heard, general superintendent of coal mines, J. B. Whatley, assistant general super-intendent of coal, and R. P. Whetmore, chief mining engineer.The Dolomite and Mulga mines are about 16 miles southwest of Birmingham, Jefferson County, Ala., on a company owned and operated railroad which connects ore mines, coal mines, coke ovens, and furnaces.In the Dolomite property there is a 30°, 860-foot, double-tracked slope for hoisting, three manway and air slopes and one air shaft, all in rock.At Mulga there is a 220-foot hoisting shaft, an air shaft and two manways, and material and air slopes, all in rock to the coal bed.The coal in both mines comes from the Pratt bed, which is the next coal above the Mary Lee bed in the Warrior Basin. The coal is horizontal or slightly undulating, with thickness varying from 36 to 72 inches and averaging 50 inches.The method of mining is essentially room-and-pillar with modifications for mechanized sections. The mines produced approximately 1,100,000 tons with 2,328,309 man-hours exposure n 1929 and 900,000 tons with 1,697,436 man-hours exposure in 1930.Through the work of the safety department the entire personnel in the coal mines, as well as in other departments, makes safety an essential part of efficient operation.Both mines are gassy; Dolomite produces about 334,000 cubic feet and Mulga about 919,000 cubic feet of methane every 24 hours.The ventilation maps, two of Dolomite mines (figs. 1 and 2) and one of Mulga (fig. 3) give as much of the details of ventilation as is practicable to put on maps of such scale. As will be seen from these maps there are two main ventilating fans in Dolomite mine and two in Mulga mine and the general direction of the air travel is indicated by long arrows to fan 1 and short arrows to fan 2 in each mine. Each fan has a gas or steam auxiliary drive and each mine has a complete fan and power unit ready for operation in emergency.In Dolomite, recently taken data show that fan 1 exhausts 82,000 cubic feet of air per minute at 3.2-inch water-gage pressure, with an analyzed methane content of 0.15 per cent, and fan 2 exhausts 115,500 cubic feet of air at 3.8-inch pressure and 0.10 per cent of methane. In addition, a Capell fan at No. 2 air shaft is a standby or emergency unit.In Mulga recently taken data show that fan 1 exhausts 83,000 cubic feet of air per minute at 3.0-inch pressure and 0.18 per cent of methane, and fan 3 exhaust 119,000 cubic feet of air per minute at 5.4 inch pressure and 0.35 per cent of methane. Fan 2 at No. 2 shaft is a standby or emergency unit for the Mulga mine.The ventilation is on a split system in both mines. All part split and total returns are measured, sampled, and analyzed weekly, requiring one man two days each week. During December, 1930, 77 samples were collected and analyzed in Mulga and 78 in Dolomite. Two weekly reports covering part of December, 1930, are given here so that they can be studied to show control of ventilation in the Dolomite and Mulga mines. In the event that a split sample shows more than 0.50 per cent of methane, immediate steps are taken to lower the explosive-gas content. The reports follow:"

Mar 1, 1932

By R. F. Randolph, R. S. Fowkes, J. J. McClelland, G. H. Schnakenberg

Between 1986 and 1992, there were268 deaths and 54,642 severe injuries to workers in underground U. S. mining operations. Clearly, this level of health and safety risk is not acceptable. While most underground coal mining jobs are dangerous, some are much more dangerous than others. Many jobs in an underground coal mine are clustered around the working face—the place where machinery digs coal from the surrounding rock. The hazards in this area are numerous: When the coal is removed, the ground around it becomes unstable and can cave in on the workers. The mining machines used there are large, heavy, powerful, and move in many different directions in a very confined space, presenting danger to workers. Also, the machines churn up coal dust that causes black lung and other ailments. Figure1 shows a modern radio remote controlled continuous mining machine and its operator. In a typical mine, the jobs with the greatest exposure to hazards involve operation of machinery that cuts tunnel-like passages into the coal (the continuous mining machine) and machines that insert supportive steel bolts into the rock above the work area (roof bolting machines). The workers who endure these hazards are at almost twice the risk of death or disabling injury than co-workers in other safer areas of the mine. Of the statistics quoted above, 106 of the 268 deaths and 21,552 of the 54,642 severe injuries occurred in jobs involving equipment at the working face. The operators of roof bolting and continuous mining machines account for only 21% of the work force but suffered 40% of the deaths and 39% of the lost-time injuries during1986-1992. These death and injury rates are almost twice the rate for workers in other underground mining jobs (see figure 2). The estimated total cost of the 106 fatalities and21,552 lost-time injuries was $327 million from 1986-1992. On average, each fatality cost over $1 million and each lost-time injury costs in excess of $9,000.

Jan 1, 1994

By E. V. Deshayes, W. E. Young

"INTRODUCTION The Camp Bird mine was investigated by Bureau of Mines engineers 3/ September 22 and 24, 1942, and April 29 and 30, 1944. In a subsequent report and supplement, it was recommended that core drilling be done below the 14th. level to determine if commercial ore existed at greater depths, and particularly if ore existed in the vein or adjacent thereto where it passes into the sedimentary formations estimated to underlie the volcanics approximately 400 feet below the 14th level.Seven diamond-drill holes aggregating 3,956 feet were drilled. Four of these holes were under the Hematite ore body and three were under the Big Discovery ore body. Two holes under, the Hematite ore body wore directed at angles but little steeper than the angle of dip of the vein and penetrated the sedimentaries beneath the volcanics close to the vein. Enough low-grade ore was found at the base of the Telluride conglomerate to indicate the presence of a large body of low-grade zinc sulfide ore. Two holes were directed to crosscut the vein at a depth of approximately 125 feet below the 14th level. No commercial ore was cut, but the vein showed the average width and same general characteristics as at the 14th level. One hole under the Big Discovery ore body was directed to crosscut the vein and test the base of the Telluride conglomerate in the footwall block. No commercial ore was found in either the vein or the sedimentary horizon. One hole was directed to test the base of the Telluride conglomerate in close proximity to the vein in the hanging-wall block. This hole was lost approximately 100 feet short of its objective, and a third hole was drilled to accomplish this same thing. The third hole was stopped approximately 50 feet short of its objective when failure of equipment caused a recess of the"

May 1, 1948