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By R. V. Ramani, V. T. Burgos, J. A. McClay

The data for Test I describes an underground bituminous coal mine operation in a 48 inch seam, with an estimated production rate of 1,056,000 tons per year. The system includes four continuous mining sections, conveyor belt haulage, and a rail network for supply and personnel transportation. This section illustrates the input data sequence and the various report formats which may be selected for the printing of results. As explained earlier, at the present time, the elements of the Cost Data Library must be read in with the regular data input. Thus, in the input data listing, immediately following the required user input, are cards which contain the inputs of the cost library. Examination of the cost reports indicate the absence of cost values for some categories for which the system would normally be expected to incur costs. This situation has arisen due to the format of the cost data provided in the U. S. Bureau of Mines reports. An example of this is the charge of $20/ton for brattice cloth while no charge is indicated for stoppings. The test report data provided a single figure for ventilation cost without a supplemental breakdown. Thus the cost indicated for brattice cloth is actually the total ventilation cost. These explanations hold true for the data input and output for Test II, Appendix III.

Jan 1, 1975

By R. V. Ramani, C. B. Manula, A. Owili-Eger

PART 1 THE INPUT AND OUTPUT DATA AND THE MINE SCHEMATIC SHOWING THE FLOWS FROM COMPUTER SOLUTIONS FOR APPLICATION I

Jan 1, 1975

By R. V. Ramani, C. B. Manula, A. Owili-Eger

ANALYTICAL PRESSURE - VOLUME COMPUTATIONS FOR THE MINE PROTOTYPE A new coal mine is projected, with entrance by slope 1,500 feet long. Nine main entries will be standard for development, five as intake and four as return. One return shaft is planned near the slope bottom, and an intake shaft and portal are planned 5,000 feet north of the slope bottom. The slope will have an area of 150 square feet and a perimeter of 55 feet. A belt and supply track will be within the slope. The return shaft will be circular with a diameter of 18 feet and will be 500 feet deep with an escapeway. Two hundred feet of the shaft and 500 feet of the slope will be concrete lined. Select the fan to operate five units per shift and drive the 5,000 feet to a proposed new portal, using the slope for an intake. Two main splits are planned with the maximum of three units on any split. The amount of air desired at each unit is 25,000 c.f.m; a shop at the slope bottom will require 10,000 c.f.m. In the solution of this or any similar problem, the first step is to determine the resistance factors. CALCULATION OF R FACTORS Slope area = 150 sq. ft.; Perimeter = 55 feet; Length = 1,500 feet. The friction factor K must be established. Five hundred feet of the slope is smooth lined, moderately obstructed by belt; K = 35. The remaining 1000 feet is in sedimentary rock and moderately obstructed; K = 85.

Jan 1, 1975

By R. V. Ramani, V. T. Burgos, J. A. McClay

The data for Test I1 describes an underground bituminous coal mine operating in a 72 inch seam, with an estimated production rate of 4,257,000 tons per year. The system includes 15 continuous mining sections, conveyor belt haulage, and a rail network for supply and personnel transportation. The following section includes the input data sequence, annual cost report, and cash flow analysis "report for this data set.

Jan 1, 1975

By R. V. Ramani, V. T. Burgos, J. A. McClay

Cost Data Library The data retained in the Cost Data Library is of a semi¬permanent nature in that it is held in storage for immediate use with the program although its contents may require periodic evaluation to maintain it in a current status. The library is composed of six arrays containing both technical and cost data. The following description of the library input system provides the information necessary to maintain an active storage file. Array POWERA: FORMAT (8F10.0) - contains the step groupings of the power rate schedule for power cost computations. Four input data cards are required to fill this storage area. Card Column 1 1 - 10 Demand in Kilowatts for the First Step of the Demand Rate Schedule 11 - 20 Demand in Kilowatts for the Second Step of the Demand Rate Schedule y 41 - 50 Demand in Kilowatts for Fifth Step of the Demand Rate Schedule 51 - 60 Maximum Demand Level for this Rate Schedule 61 - 70 Percent Power Factor for this Rate Schedule 71 - 80 Unassigned 2 1 - 10 First Step Demand Charge, $/Kw 11 - 20 Second Step Demand Charge, $/Kw 4¬ 41 - 50 Fifth Step Demand Charge, $/Kw

Jan 1, 1975

By R. V. Ramani, C. B. Manula, A. Owili-Eger

All the input data required to operate the ventilation simulator except the title, are entered with I10 or F1O.O formats. The title is entered with (18A4) format. The floating points are designated by (FP) and the integers by (INT)

Jan 1, 1975

The current procedure for measuring the concentration of respirable coal mine dust is as follows. Each filter is preweighed by the filter manufacturer to ±0.1 mg. Following sampling with the Coal Mine Dust Personal Sample Unit (CPSU) at 2.0 L/min, the filter with coal mine dust is sent to MSHA for weighing. The current MSHA procedure for weighing respirable dust samples uses a Mettler Model AE163 analytical balance in conjunction with an automatic weighing system with a precision of ±0.02 mg [Tomb 19901. Each balance is calibrated twice per day. Quality control for the automatic weighing system includes the systematic weighing of one in eight filters on a second Mettler AE163 balance. Tolerance is set at ±0.1 mg between the two weighings of the same sample. Weights are truncated at the 0.1 mg level (e.g., 3.457 mg is truncated to 3.4 mg) [Bowman et al. 19851. The difference of the two truncated weights is then recorded as the weight of coal dust deposited. The respirable concentration (mg/m3) is computed by multiplying by a correction equal to 1.38 and dividing by the volume of air sampled (2.0 L/min x sampling time [min]).

Jan 1, 1997

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

MINE VENTILATION NETWORK ANALYSIS

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

1 14 THIS IS A DIGITAL COMPUTER SIMULATION OF THE VENTILATION SYSTEM OF A MINE LOCATED IN WASHINGTON COUNTY, PENNSYLVANIA. THE VENTILATION SURVEY ON WHICH THIS SIMULATION IS BASED WAS RUN FROM OCT. 27-30 AND ON NOV. 2, 1976 IN THE ACTIVE PORTION OF THE MINE INBY COMBS SHAFT AND THE INACTIVE PORTION AFOUND AND INBY DONAHOO SHAFT. FOLLOW-UP SURVEYS WERE THEN RUN ON FEB. 18 AND FROM MARCH 7-11, 1977 TO INCLUDE THE MAIN HAULAGE SYSTEM OUTBY THE ACTIVE AREAS OF THE MINE. ALL SURVEY DATA WAS NORMALIZED TO A STANDARD DENSITY OF 0.075 POUNDS/CU.FT. PRIOR TO THE CALCULATION OF INPUT PARAMETERS FOR THIS SIMULATION. IN THIS RUN, THE ORIGINAL VENTILATION NETWORK HAS BEEN MODIFIED TO INCLUDE NEW INTAKE AND RETURN AIRSHAFTS AT THE WEST MAINS. IN ADDITION A CONNECTION HAS BEEN MADE BETWEEN THE MINE UNDER STUDY AND A SHAFT IN A WORKED OUT MINE IN THE 1 FACE AREA. THIS SHAFT IS CONSIDERED AN INTAKE SHAFT FOR THIS RUN.

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

MINE VENTILATION NETWORK ANALYSIS

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

1 11 THIS IS A DIGITAL COMPUTER SIMULATION OF THE VENTILATION SYSTEM OF A MINE LOCATED IN WASHINGTON COUNTY, PENNSYLVANIA. THE VENTILATION SURVEY ON WHICH THIS SIMULATION IS BASED WAS RUN FROM OCT. 27-30 AND ON NOV. 2, 1976 IN THE ACTIVE PORTION OF THE MINE INBY COMBS SHAFT AND THE INACTIVE PORTION AFCUND AND INBY DONAHOO SHAFT. FOLLOW-UP SURVEYS WERE THEN RUN ON FEB. 18 AND FROM MARCH 7-11, 1977 TO INCLUDE THE MAIN HAULAGE SYSTEM OUTBY THE ACTIVE AREAS OF THE MINE. ALL SURVEY DATA WAS NORMALIZED TO A STANDARD DENSITY OF 0.075 POUNDS/CU.FT. PRIOR TO THE CALCULATION OF INPUT PARAMETERS FOR THIS SIMULATION.

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

MINE VENTLATION NETWORK ANALYSIS

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

Table

Jan 1, 1977

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

Le Système International d'Unités (SI) Equivalents of Common United States Units Length 1 ft (foot) = 0.30480 m Mass 1 lbm (pound mass) = 0.45359 kg Force 1 lb (pound) = 4.4482 N Volume Flow Rate 1 cfm = 0.00047195 m3/s Mass Flow Rate 1 lbm/min = 0.00756 kg/s Pressure (wg at 20° C) 1 inch wg (water gage) = 248.36 N/m2 1 inch wg = 248.36 Pa Power 1 hp (horsepower) = 0.74570 kW Energy 1 ft-lb = 1.3558 J Torque 1 lb-inch = 0.11298 N-m

Jan 1, 1977

By Richard H. Spencer, Warren G. Bender

Since the passage of the Coal Mine Health and Safety Act of 1969, the Bureau of Mines has made many advances in mine communications. These efforts have resulted in the following developments and demonstrations: • Surface-to-surface electromagnetic transmission of baseband voice signals through overburdens as deep as 600 feet. This transmission was achieved with equipment from the Interim Mine Rescue and Survival System Program, and demonstrated that baseband voice communications could be obtained through these overburdens in the absence of mine-generated noise; that is, when the mine power system was shut off. However, during mine operational conditions, the mine-generated noise proved to be so large that effective baseband voice communications to the desired depths could not be effected. • The development of beacon transmitters for locating trapped miners. This work was again the outgrowth of the Interim Mine Rescue and Survival System. Extensions of this work have led to reliable detection and location of trapped miners to depths of 1000 feet. The system, which makes use of a loop antenna deployed where the trapped miner is located, uses low-duty-cycle, tone-burst transmission. Portable tuned receivers and loop antennas are used on the surface to determine the location of the trapped miner. An extension of this work has led to a call-alert paging system concept, wherein a transmitter, similar to that used in the trapped miner system, is used to audibly and/or visually alert key mine personnel carrying pocket receivers that they are wanted at the nearest mine telephone. In addition, a roof-bolt paging system has been developed, wherein commercial paging and mine carrier phone equipment are used in conjunction with a long-wire "roof bolt" antenna to page individuals and relay to them short voice messages.

Jan 7, 1975

By J. Shawn Peterson, David S. Yantek, Adam K. Smith

Exposure to excessive noise over time can cause permanent hearing loss. Workers in the mining industry are frequently exposed to A-weighted sound levels in excess of 90 dB. The A-weighted sound level at the roof bolter operator’s location can exceed 100 dB while drilling. The National Institute for Occupational Safety and Health (NIOSH) measured the sound pressure at the roof bolter operator’s position and utilized a phased array of microphones with beamforming software to identify noise sources on a roof bolting machine while drilling. The test data indicates the sound level at the operator’s position is dominated by the 2000 Hz through 5000 Hz one-third­octave-band sound levels. The beamforming results indicate that the drilling noise is primarily from two areas: the portion of the drill steel just below the rock and the drill steel-chuck interface. This paper will discuss the methods used to identify the primary noise sources.

By V. A. Marple, K. L. Rubow

"IntroductionA particle dispersion system has been developed to aerosolize bulk powders or material deposited on surfaces such as filters. The size distribution of the aerosolized particles can then be obtained by sampling the particles with an aerosol sizing instrument such as a TSI aerodynamic particle sizer (APS) or a cascade impactor.Of particular interest is the applicability of using this technique to obtain the size distribution of particles collected on filters. This system has been used to determine the size distribution of coal dust particle collected on membrane filters. The use of this device in this application is quite intriguing in that, if this technique can be shown to accurately determine the size distribution of airborne particles, one can easily collect particle samples using either personal or area particle samplers, determine the mass concentration of the collected material through gravimetric techniques and then use the particle dispersion system coupled with the APS to determine the particle size distribution. With this technique one can determine the particle size of dust in environments, such as underground mines, where it would be inconvenient or impossible to use a particle sizing instrument.Particle Dispersion SystemA prototype particle dispersion system has been developed for aerosolizing particles. A schematic diagram of the system is shown in Figure 1. The system consists of mechanism for advancing the particle sample beneath a dust pickup tube, a venturi for deagglomerating the particles, and a instrument for analyzing the aerosolized particulate matter.The procedure to analyze a particle sample is as follows. The particles to be aerosolized are placed on a sample plate from where the particles are picked up by an air stream flowing upward in a small capillary tube (dust pickup tube). This tube operates as a small vacuum cleaner passing over the sample plate. The air is drawn up the capillary tube at a rate of 2 L/m by a venturi aspirator. This is accomplished by placing the outlet of the capillary tube near the throat of the venturi. The accelerated velocity of the air creates a low pressure in this region which draws flow up the capillary tube. From the outlet of the capillary tube particles are transported to the inlet of the APS where a small fraction are sampled and the aerodynamic diameter of the particles determined. The particles which are not measured by the APS are passed into a HEPA filter."

Mar 1, 1989

Demonstrate the technical feasibility of using nontraditional, low-rank, coal-based fuels and new coal-based technologies for mineral and metallurgical applications. Approach A cooperative research effort was initiated between the Bureau of Mines, U.S. Department of Energy and the minerals industry to extend the application of the fixed-bed gasification process to low-rank Western coals. The research effort enabled cooperators to gain operating experience in the use of hot, raw producer gas for iron ore pellet induration and other high-temperature mineral processes. Researchers were able to gather data regarding the fundamental engineering aspects of the fixed-bed process and gas cleanup for a wide range of solid fuel properties. Applying Fixed-Bed Gasification As a prerequisite to gathering the data needed for the study, the Bureau of Mines designed and installed, at one of its laboratories, a 6.5-ft-diam. single stage, fixed-bed, air-blown atmospheric pressure gasifier. The gasifier was coupled with a hot, raw gas combustion system and iron ore pellet in. duration kiln. The gasifier is rated at 1.5 ton/h capacity of bituminous coal.

Jan 1, 1987

By M. A. Trevits, M. R. Thibou, A. Ozment, A. C. Smith, J. B. Walsh

The Excel No. 3 Mine is a room-and-pillar mining operation owned by MC Mining, LLC, a subsidiary of Alliance Resource Partners, LP, which produces coal from the Pond Creek Coalbed in Pike County, Kentucky. Late on December 25, 2004, a fire was discovered near the bottom of the slope and, after an unsuccessful attempt to fight the fire underground, it was decided to address the fire remotely from the surface. The mine operator developed a plan to fight the fire that called for the installation of temporary seals at the slope and mine shafts to eliminate the inflow of oxygen. In addition, the plan called for drilling and completion of vertical boreholes into select areas of the mine workings to monitor the mine atmosphere and inject nitrogen gas and gas-enhanced foam. Over the next 9 days, nitrogen gas was injected almost continuously and gas-enhanced foam was injected periodically as deemed necessary. On January 7, 2005 it was determined that the fire was successfully suppressed. This paper presents a discussion of the fire-fighting approach and an analysis of the results of the application of gas-enhanced foam technology.

By William D. Monaghan, Michael A. Trevits

Over the period 2000 to 2003, roof falls have accounted for 4 to 14% of the fatalities in underground mining operations. The National Institute for Occupational Safety and Health (NIOSH) is conducting research to reduce the frequency, exposure, and risk of these events through an ongoing program of field and laboratory studies. One area of research involves the evaluation of polyurethane grouting technology that is commonly used to stabilize fractured mine roof strata. Concurrently, NIOSH is conducting work to evaluate the application of ground penetrating radar for advanced delineation of problematic mining areas. In this study, NIOSH partnered with Sub-Technical, Inc. who injected a polyurethane grout (also known as glue) into a roof area of NIOSH’s Safety Research Coal Mine. The mine roof area was scanned using ground penetrating radar technology before and after grout injection in an attempt to determine the extent of grout infiltration. A comparison of the pre-and post-grouting radar records showed a significant change at a mine roof depth of about four to five feet. The interpreted radar records were then compared with drill core information, borescope evaluation of roof-bolt holes in the study area, and underground observations. At this site, the interpretations of the radar records correlated with data obtained from the core holes, borescope evaluations and underground observations. This study showed that ground penetrating radar technology can be a useful tool for detecting changes in mine roof due to the injection of the grout.