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By A. J. Stewart, M. Guay, J. W. Critchfield, A. P. Hoddeson

The MetroWest Water Supply Tunnel project includes three main work shafts, excavated primarily in rock. Shaft L is 9.8 m (32 ft) diameter and 124.1 m (407 ft)deep. Shaft E is 7.0 by 12.8 m (23 by 42 ft) and 91.4 m (300 ft) deep. Shaft 5A is 7.9m (26 ft) diameter and 72.5 m (238 ft) deep. Construction methods are compared and contrasted. Lessons learned involve a differing site condition in soil overburden, contractor-proposed modifications, and remedial slurry wall work. Other aspects discussed include noise and vibration control, groundwater inflow, water treatment, muck handling and contract administration issues.

Jan 1, 1999

By Clay Haynes

While the Kansas River Tunnel is a relatively short, small diameter tunnel, the shaft excavation methods implemented were long and large on innovation and use of reverse circulation drilling equipment. ATS Drilling, L.P. used mechanical excavation techniques for both soil and rock segments of the 42.7 m (140 feet) and 38.1 m (125 feet) deep shafts that were 2.9 m (9.5 feet) and 4.1m (13.5 feet) diameter, respectively.

By Robert A. Jankowski, J. Drew Potts

The Bureau has identified several ventilation techniques and a water spray application for reducing auger mining dust levels. The dust control methods all work by improving air quantity or flow patterns at the face. Section airflow is maximized by reducing air leakage between entries. Sealing stoppings with brushed- on mortar, containing fiberglass additives, reduces leakage. Heading airflow is maximized by properly installing the brattice curtain and keeping the area between the curtain and rib free from obstructions. The brattice curtains should be hung at least 1.7 m (5.5 ft) from the ribs in 7.9 m (26 ft) wide entries and kept less than 4.6 m (15 ft) from the face. A tight curtain seal is created by using the mine air pressure to force the curtain against the timbers, roof and floor of the mine. Using both a blowing curtain and an exhausting curtain will isolate some face workers from dusty air. The Bureau designed, auger clearer spray system, can be used to force heading air across the active face. Laboratory tests indicate that the auger clearer spray system can increase face velocities by more than 0.5 m/sec (100 ft/min), while using less than 0.13 l/sec (2.0 gal/min) of water.

Jan 1, 1989

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

Canadian mining giant Teck Resources Ltd. will be the exclusive supplier of the metals used in the production of the more than 1,000 medals to be awarded at the Vancouver 2010 Olympic and Paralympic Winter Games. Teck, the diversified resource company headquartered in Vancouver, B.C., is providing 2.05 kg (4.5 lbs) of gold, 1,950 kg (4,300 lbs) of silver and 903 kg (2,000 lbs) of copper sourced from its operations around the world. In addition to the raw materials mines, this year?s medals will make Olympic history as they will be the first to contain metals recovered from processing the circuit boards from end-of-life electronics (e-waste) that would have otherwise been destined for a landfill. After mining or production through a variety of smelting and recovery processes, each metal has been refined to enhance the metal purity.

Jan 1, 2010

By Bruce Clements

The Bagdad mill, located near Bagdad, AZ, was built in 1977 with a designed mill capacity of 36.3 kt/d (40,000 stpd). An autogenous mill-ball mill crusher (ABC) grinding circuit was chosen. The original mill contained three independent grinding circuits. Each contained one 9.75m-diam x 4-m-long (32-ft-diam x 13-ft-long) fixed-speed autogenous mill, a4.7-m-diam x 6.7-m- long (15.5ft-diam x 22-ft-long) ball mill and one 2. 1-m-diam (7-ft-diam) short head cone crusher. Total installed power per grinding circuit was about 8.4 MW (11,350 hp), with 5.9 MW (8000 hp) in the autogenous mill, 2.2 MW (3000 hp) in the ball mill and 261 kW (350 hp) in the crusher (Fig. 1). The autogenous mills were designed to be able to operate semi-autogenously. However, they were never converted to semiautogenous grinding (SAG) mills. As a result of this design, only about 4.1 MW (5450 hp) of the designed power is used in the autogenous mills. The grinding circuits are closed, with 2.4 x 6 m (8 x 20 ft) double-deck vibrating screens. A series of conveyors re- turn screen oversize underneath the mill coarse ore stockpile to feed the crushers. Variable speed belts feed the crushers and control power draw. Crusher product reports directly to the tail end of the autogenous mill feed belt (Fig. 2). A magnetic sensor detects tramp iron in the autogenous mill-crusher circuit.

Jan 1, 1992

By Barnaby J. Watten, Sibrell. Philip L., Brian J. Vinci, Andrew E. Friedrich

We evaluated a new process for remediation of acid rock drainage (ARD). The process treats ARD with intermittently fluidized beds of granular limestone maintained within a continuous flow reactor pressurized with CO2. Tests were performed over a thirty day period at the Toby Creek mine drainage treatment plant, Elk County, Pennsylvania in cooperation with the Pennsylvania Department of Environmental Protection. Equipment performance was established at operating pressures of 0,34,82, and 117 kPa using an ARD flow of 227 L/min. The ARD had the following characteristics: pH, 3.1; temperature, 10 OC; dissolved oxygen, 6.4 mg/L; acidity, 260 mg/5; total iron, 21 mg/L; aluminum, 22 mg/L; manganese, 7.5 mg/L; and conductivity, 1400 uS/cm. In all cases tested, processed ARD was net alkaline with mean pH and alkalinities of 6.7 and 59 mg/L at a CO2 pressure of 0 kPa, 6.6 and 158 mg/L at 34 kPa, 7.4 and 240 mg/L at 82 kPa, and 7.4 and 290 mg/L at 117 kPa Processed ARD alkalinities were correlated to the settled bed depth (p<0.001) and CO2 pressure (p<0.001). Iron, aluminum, and manganese removal efficiencies of 96%, 99%, and 5%, respectively, were achieved with filtration following treatment. No indications of metal hydroxide precipitation or armoring of the limestone were observed. The surplus alkalinity established at 82 kPa was successful in treating an equivalent of 1136 L/min (fivefold dilution) of the combined three ARD streams entering the Toby Creek Plant. This side-stream capability provides savings in treatment unit scale as well as flexibility in treatment effect. The capability of the system to handle higher influent acidity was tested by elevating the acidity to 5000 mg/L with sulfuric acid. Net alkaline effluent was produced, indicating applicability of the process to highly acidic ARD.

Jan 1, 2000

By Niels Jonassen

INTRODUCTION The occurrence of the radioactive element radium, [2gg~a], often in trace amounts, in the crust of the earth gives rise to a flux of the daughter product radon, [&apos;ii~n], across the earth-air interface. This so-called exhalation process has been studied by many investigators [KO 45, Kra 64, Pe 66, Wi 741] over several decades. In the earlier investigations the exhalation process was studied in a purely geophysical context, but over the last 20 years it has been realized, that under certain conditions the exhalation may be intense enough to also have radiological implications. This is for example the case where relatively radium-rich wastes or by-products, like uranium mill tailings and phospho-gypsum, are stored in larger quantities. In order to be able to predict the radiological impact on the environnent it is necessary to know the rate of exhalation of radon from the surface of the exhaling material. Usually the determination of these exhalation rates has been done by measurements in situ directly upon the accumulated material [~le 74, Ha 791. Although such field measurements have the obvious merit of dealing directly with the full scale problem, the results are often lacking in generality, making predictions for other situations difficult or even impossible. The present paper describes some results of an investigation where it is attempted from laboratory measurements on samples of limited size to obtain material parameters, which are suited for extrapolating and predicting the radiational behaviour of larger accumulations of the material. THEORY Let us consider a sample of a radon exhaling material enclosed in a container with a dead space volume, Vd, and initially containing radon free air. If the total exhalation rate (number of atoms per unit time) is c, the radon concentration (number of atoms per unit volume) N will initially grow with the time t according to the equation [ ] where [R] is the decay constant of radon. As the concentration (or activity) in the container grows, the net exhalation rate will slow down because of back diffusion, and, furthermore, radon atoms may be removed from the container, not only by radioactive decay, but also by diffusion leakage [JO 81 ]. If, however,only the initial. part of the activity growth curve is considered, the total exhalation rate, c, is given by [c = r•Vd(2) where r is the initial activity growth rate in the container. Let the container be cylindrical with an inner cross-section S and filled to a height L with a radon exhaling material with the specific mass p, the porosity E, the radon pore production rate f and the diffusion length i. When the container is left open,i.e. the radon exhalation takes place into a virtually radon free room, the radon area exhalation rate, E, can be written [Jo 80] E = Ef?t tanh L.R (3) assuming the exhalation only takes place in one direction, towards the surface of the material. It appears that for values of L > 2•ithe area exhalation rate becomes E a = Efi (4) independently of the sample dimensions and the total exhalation rate c is then c = E •S (5) For L ;~ 0.5•i, tanh P _ k and equation (3) may be written E ~ EfL (6) then c = EfL•S = p•m= Em •m (7) where m is the mass of the sample and Em = P (8) is the mass exhalation rate. It thus appears that the radon exhalation rate from a given amount of material may often be conveniently expressed by c = E •Sfor L z 2•I (5) a orc = Em •mfor L0.5•i (7) where S is the total exhaling area and m the total exhaling mass, E and E the area and mass exhalation rates respectively defined by equations (4) and (8).]

Jan 1, 1981

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By Donald J. Olson, Thomas J. Lutzenberger, John Ramage

The North Shore Inline Interceptor, Phase IA tunnel is a 8,550-meter (28,000-ft) interceptor tunnel currently being constructed by Traylor Bros./Frontier-Kemper, Joint Venture, using a 9.8-meter (32-ft - 4-in) diameter Robbins tunnel boring machine (TBM). This tunnel is being driven in Silurian age dolomite approximately 91.5 m (300 ft) below ground surface. Significant water inflows were encountered when the tunnel drive penetrated the thinly-bedded Waubakee Formation. Flow increases approaching 0.75 cu m per min per 30 m of tunnel (200 gpm per 100 ft) have occurred in this formation. After driving approximately 2,285 m (7,500 ft) of tunnel (200 m full face in the Waubakee Formation), total flows exceeded 12 cu m per min (3,200 gpm) with about 5.7 cu m per min (1,500 gpm) entering the tun¬nel from the 200 m (600 ft) of exposed Waubakee Formation. Additional water control measures consisting of increased pumping capacity, increased water treatment capacity, surface grouting ahead of the tunnel, grouting ahead of the tunnel face and panning to improve the working environment for installation of tunnel support, were implemented to deal with the large water inflows. Currently (January 24, 1989), 3,400 m (11,000 ft) of tunnel has been driven, total water inflows are in the range of 11.7 cu m per min (3,100 gpm) with only 6.8 cu m per min (1,800 gpm) inflow from about 1,220 m (4,000 ft) of tunnel driven full face through the Waubakee Formation.

Jan 1, 1989

By D. R. Dolinar

The fall of small pieces of rock between roof bolts causes over 80 pct of the ground fall injuries in underground coal mines. Roof screen, if installed on the bolting cycle, can significantly reduce the number of these injuries. As the surface of the roof fails, the screen will have to retain the broken rock without failing or excessive deformation. Welded wire screen is the most common type of screen material used for roof surface control in underground coal mines. The load-deformation characteristics and the load capacity of welded wire screen were evaluated in the laboratory in a specially built test frame. In the frame, the square screen sections were held in place by bolts and bearing plates at the four corners while a center load was applied to the screen. The parameters evaluated included the wire gauge and configuration, the bearing plate loads and bolt spacing. The wire gauge of the tested screen was 4, 6, 8, 9 and 10. Maximum screen capacities for both load and displacement were developed for each of the wire gauge sizes along with the screen stiffness. This information can be used for screen selection and for the evaluation and determination of screen performance. DISCLAIMER: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009

By G. Danko

MULTIFLUX, a new thermal, hydrologic, and airflow model and software is being developed to solve the flow of heat, moisture, and air in and around an underground opening. MULTIFLUX couples two distinct domains that are solved separately first and re-coupled as an iterative task: (1) the rockmass, and (2) the airway. The airway is solved with a lumped-parameter Computational Fluid Dynamic (CFD) module which is an embedded part of the MULTIFLUX code. The CFD model includes convection, conduction, and radiation for heat, as well as convection and diffusion for moisture transport in an air-filled opening. The CFD model also solves for the airflow field. The surrounding rockmass model may be from any analytical solution, or from a complex thermal-hydrologic numerical model such as NUFT or TOUGH2. The rockmass model is interfaced to MULTIFLUX using a novel technique called Numerical Transport Code Functionalization (NTCF). The method provides fast convergence on a personal computer platform. The paper reports the results of MULTIFLUX simulations for a mine drift, comparing calculations with a known mine climate software, CLIMSIM, and with measured data. Through this validation example, the performance of MULTIFLUX is demonstrated. The results show excellent agreement due to the use of realistic modeling assumptions and boundary conditions.

Jan 1, 2009

By David van Cleave, David L. Rome

Whether you buy your minicomputer from a manufacturer or another vendor, the first thing you must do is define your needs. With good planning, the initial cost will be low and your return on investment will be maximized over the long term. Involving all members of your staff who will he using the new computer in the planning phase is also important. Their participation will alleviate any potential concerns they may have about new equipment, and their suggestions will provide valuable input for your specifications. After you define your proposed system, survey available hardware and software. The survey will help you in the next step, preparation of specifications. The specifications, along with a quote request, are then sent to various manufacturers. Consider the Manufacturer When selecting your minicomputer, one important item to consider is the manufacturer. Remember, when you buy a computer, you&apos;re also buying the services of a computer company. A firm that makes a commitment to its customers and provides a product that is field tested before releasing it for sale is preferable. Examine the company&apos;s maintenance and support packages, since every machine requires preventive maintenance. Does your manufacturer or original equipment manufacturer (OEM) support service sites in your area? Has their service record been commendable? Will they provide training for your personnel and a supply of circuit boards so you can retain a degree of self-service autonomy? Three methods of buying a computer involve the manufacturer, an OEM, and the systems house. Buying direct from the manufacturer means you go straight to the source, using no middleman. This gives you maximum flexibility; you can consider every alternative and many possible combinations of computers, peripherals, and software until you find exactly the right system for you. The OEM buys computers direct from the manufacturer and builds them into his own computerized, special purpose product. He usually sells a complete package, including software and hardware, and often noncompetitive equipment as well (a scientific or engineering instrument, for example), designed for the mining market. Because he offers a packaged product, the OEM can often charge less for the system than if you put the pieces together yourself. In many ways, the OEM also provides a unique system. And while the OEM sells a packaged solution, he also usually has the special market knowledge necessary to modify the system to meet your unique needs, if necessary.

Jan 11, 1981

By R. D. Williams

In spite of the increasingly polarized and politicized view of climate science and climate change in the United States, many of the principles behind the ?theory? of Anthropogenic Global Warming (AGW) or Global Climate Change (GCC) are in fact firmly established both in fundamental physics and detailed observations and are widely accepted elsewhere. The discovery of the earth?s ?greenhouse effect? dates back nearly 200 years to the work of French mathematician and physicist Joseph Fourier in 1824 and 1827. In 1872, John Tyndall an English physicist and mountaineer, described the capacity of various gasses to absorb radiant heat. He identified water vapor and carbon dioxide as the principal ?greenhouse? gasses. The work of Fourier and Tyndall was the foundation for the theory that increasing carbon dioxide (CO2) in the atmosphere might be responsible for an increase in global temperatures developed by the brilliant Swedish chemist Svante August Arrhenius in 1896. This was backed up by the additional work of Gilbert Callendar in the 1930s, and Gilbert Plass, Hans Suess and Roger Revelle in the 1950s. There is little scientific dispute about the physics by which this occurs: carbon dioxide in the atmosphere absorbs long wave radiation that would otherwise escape into space, thus warming the lower atmosphere and respective temperatures worldwide. The Intergovernmental Panel on Climate Change (IPCC), approximately 2500 climate scientists from around the world, has concluded that increases in carbon dioxide resulting principally from the burning of fossil fuels are responsible for most of the continuing increases in global temperatures (IPCC, 2007). NASA?s Goddard Institute for Space Studies (GISS) states; ?The ten warmest years all occur within the 12-year period 1997-2008? (GISS, 2009). 2009 was the second warmest year on record and 2010 was tied with 2005 for the warmest year on record (GISS, 2010, 2011). The science is in fact settled; the earth has not cooled since 1998 as some have claimed. It continues to warm as evidenced by not only surface temperature records but also ocean temperature records as well as changes in physical and biological systems (IPCC, 2007).

Feb 27, 2013

By Scott G. Britton

Microcomputeranalysis has long been a technique for engineering professionals to improve the theoretical and practical design of continuous mining systems. Such work has been typically composed of modeling observed events and then introducing a variety of known elements with predicted results to compare against the results of the model. Programs written in BASIC, Cobol, Fortran or other language dominated such efforts. These programs have taken years to develop and prove out. This article explores a new management tool used daily at Tanoma Mining Co. (TMC) to track and improve continuous mining systems through the use of microcomputers and Lotus l-2-3 software. Data provided by section foreman, shift foreman, and maintenance personnel are analyzed against predetermined goals. In this way, new production targets are developed, weaknesses in the current mining systems are identified, and the results of changes are measured. The current results of this system on the productive output of TMC systems is on the order of 10% improvement. Thin seam coal mining is an industry within the broader industry of coal mining. Thin seam mining faces the same cost pressures and engineering problems as its higher-seam competitors without the advantages of moving volume per foot of extracted material to stay at an economic parity with others requires additional creativity within the management teams to extract higher productivity gains from the existing workforce and mining systems. TMC operates a deep mine in Indiana bounty, PA, about one hour northeast of Pittsburgh.

Jan 1, 1991

By W. K. Hitchcock

Hydrogen is regarded as being a high temperature indicator of spontaneous combustion. It is currently used in underground coal mines to detect advanced heating?s. Previous work has mostly consisted of small scale tests with high air flow to mass ratios. A large-scale test rig has been used to investigate the gas indicators of coal self-heating. In particular, the relationship of gas evolution, the location of its evolution and the temperature at which it is evolved has been examined. It has been found that there are two distinct zones of hydrogen production: one at the hot spot and the other in an oxygen deficient moist zone downstream from the hot spot which is at a relatively low temperature of approximately 100°C. This has implications for the interpretation of mine atmospheres.

Jan 1, 2009

By Gordon Miner

GENERAL DESCRIPTION At the Star mine, Hecla Mining Co., located in Burke, Shoshone County, ID, the mining method is horizontal slicing (timbered or untimbered) cut-and-fill using hydraulic sand backfill. Back stoping as well as breast stoping is used; see Fig. 1. The annual produc¬tion from all cut-and-fill mining is 259 816 t (286,400 st). The average depth ranges from 366 to 2408 m (1200 to 7900 ft). A majority of the tonnage comes from an average depth of 2240 m (7350 ft). Average travel time to the stope is 1 hr. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size and Ground Conditions The ore body is 610 to 1219 m (2000 to 4000 ft) along the strike and 2408 m (7900 ft) on the dip, with an average vein width of 2.6 m (8.5 ft). The dip is essentially vertical. The ore body is heavily sheared and broken on the main fault structures; it is more competent on the wall veins but everywhere subject to rock bursting at depth. The hanging and footwalls are strongly fractured on the main fault structures, again more competent but still fractured on the wall veins. All host rock is Revett quartzite. Value and Production Requirements The ore value is $56 per ton. Production require¬ments are 1089 t/ d (1200 stpd), five days per week. Why Cut-and-Fill? The slusher extraction method is the most satisfac¬tory and practical for narrow vertical veins with sheared walls. The method must be quite selective to be com¬mercial. The veins are too narrow for rubber-tired ex¬traction and the grade is not high enough to make resuing practical. DETAILED DESCRIPTION Development Access to the main shaft is by the 3.2-km (2-mile) main haulage tunnel. The level intervals in the vertical timbered shafts are 61 m (200 ft). Vein development is by crosscut and laterals located 30.5 m (100 ft) into the vein wall. Vein crosscuts are on ±61 m (200 ft) centers. Access to the vein is by raises carried up with the stope. The raises are equipped with four wood¬cribbed chutes (two orepasses reinforced with angle irons, one manway, and one timber slide, Figs. 2-6). The initial cut is located 6.7 m (22 ft) above the main level (Fig. 1). Stoping Cycle Detail Sampling: Grab samples are taken from the mine cars. Channel sampling of the faces and backs is done when necessary. Drilling: For stope drilling, Ingersoll-Rand 300M percussion drills are used, with Ingersoll-Rand 300M and Gardner-Denver RB83M stopers, used for back stoping. Breasting down is done with Gardner-Denver S83FM jacklegs. The back stoping rounds are 0.9 to 1.8 m (3 to 6 ft) and the breast stoping rounds 1.5 to 3m (5 to 10 ft). Blasting: For stope blasting DuPont Tovex 100 watergel with fuses and igniter cord is used primarily. Approximately 60% of the blasting is done with am¬monium nitrate-fuel oil (ANFO). For development headings and shaft sinking, DuPont Drivex watergel with Ensign Bickford Nonel Primadet primers is is used. Mucking: Broken ore is removed entirely by slusher in the stopes. These are mostly 3.7-7.5-kW (5-10-hp) Ingersoll-Rand air-powered or 7.5-37.3-kW (10-50-hp) Ingersoll-Rand electric-powered. In the horizontal de¬velopment headings Eimco 12B and 21 and Atlas Copco LM-56 overshot mucking machines and a 0.76 (1 cu yd) hydroelectric Wagner HST-1 LHD (load-haul-dump) unit are used. Rockway for Extraction: The cribbed chutes (Fig. 7) are nearly vertical and hexagonal in shape and are made of spaced 152 x 203-mm (6 x 8-in.) timber, reinforced with varying weights and sizes of angle iron, depending on wear requirements. Rectangular cribbed chutes are used where ground pressure is not a problem. Cribbing is added in 2.4 or 3-m (8 or 10-ft) extensions accord¬ing to the floor height (whether timbered or untimbered). See Figs. 8 and 9. Backfilling: Backfilling is done with hydraulic sand consisting of classified mill tailings at 65% solids. The fill contains approximately 20% 325-mesh fines. The tailings are pumped 3353 m (11,000 ft) at 50-55% solids 22.7 dry t/h (25 dry stph) with a duplex piston pump. Cycloning increases the tailings density to 65% solids. Fill storage is in a 998-dry-t (1100 dry st) agi

Jan 1, 1982

As metal prices fluctuated over the years, so too have the fortunes of many of those mining in Canada?s Yukon province. With a rich mining history that dates back to the Klondike gold rush of 1896 to 1898 and continues to thrive today, Yukon has experienced several ?boom-and-bust? cycles. However, during the world?s most recent mining ?bust? Yukon has held strong and has re-mained one of the world?s hottest destinations for mineral exploration. In fact, the area even exceeded forecasts in 2009, more than doubling last year. And the near-future looks just as bright.

Jan 1, 2010

Grouting is mainly used to improve mechanical properties or decrease the permeability of rock mass around civil and mining projects. The purpose of this paper is to present the most effective factors on grout penetration depth in rock mass. Such factors are essential for reliable estimation of the grout penetration depth and the expense for cement take of any grouting project. In this paper, about 50 different types of the relations presented so far by various researchers to estimate the depth of penetration in fractured rock were studied. The effective factors are classified to rock mass properties, grout properties, and operational factors. Results show that firstly, the discontinuities aperture and Lugeon value are respectively the most investigated factors of rock mass properties influencing on penetration depth. Secondly, the yielding stress, viscosity, and density of grout are the most effective factors of grout properties. Finally, grouting pressure has an intense effect among other operational factors. The investigations were also verified in Siah Bishe dam case study. Keyword: grout penetration depth, rock mass properties, grout properties, operational factors.

Feb 27, 2013

By k. Osei-Boakye

This paper presents the calibration of a real-time aerosol monitor, the TSI DustTrak?, to the NIOSH 5040 method being used by the Mine Safety and Health Administration (MSHA) for compliance enforcement evaluation. The standard NIOSH 5040 instrument for compliance evaluation of diesel particulate matter (DPM) has an inherent time lag of up to two weeks for laboratory processing of sample results. Therefore, providing the underground mines with a means of quantifying DPM conditions in real-time, will enable them to efficiently and effectively manage DPM levels to ensure compliance with the regulated limit.

Jan 1, 2009