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By Richard E. Hucko

The bureau of Mines conducted a laboratory scale investigation of the process of selective flocculation with the objective of developing a method to effect separations on suspensions consisting predominately of ultra-fine-size (minus 400-mesh) particles of coal and refuse. Selective flocculation of coal from coal-refuse slurries could be achieved with a wide variety of reagents. Typically as much as 15 to 20 weight-percent of high-ash solids (70 to 80 percent ash) could be retained in suspension. The flocculated sediment remained a high-ash product, however, and would require further treatment to be regarded as clean coal. No combination of reagents and conditions led to selective flocculation of the refuse component. The dispersing agent was found to be the critical factor; dosages of 3 to 8 lb/ton2 were necessary to stabilize the suspension for subsequent selective flocculation. In addition, most of the dispersants that worked were most effective at substantially elevated pH (11 or so). It was concluded that this process would be prohibitively expensive and, because of yield and quality constraints, it would be of questionable utility in today's coal preparation industry.

Jan 1, 1977

By Edward A. Mihok

Laboratory batch-flow tests conducted on a highly ferruginous acid mine water with an activated carbon showed that ferrous iron is rapidly oxidized. This indicates that air oxidation of ferruginous acid in mine waters is catalyzed by activated carbon. The ferrous iron content of an acid mine water flowing through an aspirated activated carbon column was reduced from about 700 to about 10 parts per million in less than 1 minute. Highly efficient ferrous iron oxidation was achieved by conditioning the activated carbon by repeated treatment with acid mine water and air. The reaction proceeds rapidly in the acid environment (pH about 2.5) engendered by ferrous iron oxidation. The activated carbon was used repeatedly with little loss in efficiency, no discoloration, or solids deposition. In these batch tests, about one-third of the iron was adsorbed by the carbon, but with continuous flow, little iron adsorption is expected. Incorporating a catalytic oxidation step in the treatment of acid mine waters containing ferrous iron would greatly facilitate neutralization--precise control of pH, lower reagent costs, elimination of aeration equipment, and smaller holding ponds.

Jan 1, 1969

By S. M. Shelton, W. A. Calhoun, M. M. Fine, T. L. Johnston

"This report is () expansive test program by the Bureau of Mines to determine methods from various and ferrous manganese-grade product () widespread () and manfniferrous iron ores. Among the samples received for testing was one from the pits of the Barytes Mining Co., 1-1/ 2 miles northeast of Cartersville, Ga.. The property is approximately in the center of the Cartersville manganese district and shares a common corner with properties of the New Riverside Ochre Co. and the Etowah Development Co., both producers of manganifereus ores. The deposit is typical of the district and occurs in residual clays resulting from weathering of the Shady limestone and Weisner quartzite. 7/The property is developed by a number of open pits ranging to 60 feet in depth. In 1918 several carloads of manganiferous iron ore was mined from an alit on the property. Since then mining operations have been negligible. The ore is known locally as calico ore.A 4-ton sample representing -re as it occurs in the -pen pits vas furnished for testing. The sample consisted of a mixture of red iron oxides, wad, and clay, with small pebbles of manganese oxides and siliceous fragments scattered throughout."

Mar 1, 1943

By Peter L. Swanson

NIOSH has developed an automated PC-based seismic monitoring system for use in mine ground control studies. The distribution and coordination of data acquisition and processing functions across a local area network (LAN) is a central feature of the system. Design requirements include (1) the use of lowcost, off-the-shelf, data acquisition and computer hardware, (2) the capability of automatically merging waveform data sets from different types of monitoring systems, (3) delivery of raw and processed data underground and on the surface, as well as to remote sites, and (4) automated data collection and processing over long periods of time without the need for human intervention. Current mine-wide installations utilize multiple, autonomous seismic networks located both underground and on the surface and provide 20 to 40 stations within a radius of several kilometers. Event locations, magnitude estimates, and other processed data are distributed for display and analysis on network nodes both underground and in the offices of mine engineers and managers. This paper describes the system's hardware and software components, examples of installations in hard-rock and coal mines, and a few observations from western U.S. longwall coal mines.

By Charles E. Munroe

"It has been recently learned that certain foreign manufacturers are offering far importation into this country detonators at greatly reduced prices, some of these detonators being of such low grade (No.4), that their use would greatly increase the dangers in blasting.Detonators and electric detonators are used in initiating the explosion of the various high explosives that are now so extensively employed in a great variety of mining operations. Originally a large number of different grades (or ""strengths"") of detonators and electric detonators were marketed, the grade number indicating the weight of the charge of fulminating material contained in the copper case of the detonator. As late as 1914, a report of the Bureau of Mines (Bulletin 80, A primer an explosives for metal miners) shows that grades Nos. 3, 4, 5, 6, 7, and 8, containing from 0.54 to 2.00 grams each, were being offered for use. However, the writer bad observed as early as 1896, when the more insensitive types of explosives like nitro-substitution explosives began to be introduced for use in mining and other blasting operations, that use of a high-grade detonator was essential to insure certainty of discharge."

Mar 1, 1921

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 Eric Zahl, John Dunford, Hamid Maleki

Coal bumps are defined as sudden failures of rock and coal near entries that are of such a magnitude that they expel large amounts of material into a mine opening. Coal bumps are influenced by geologic conditions, the geometric design of coal mine excavations, and the sequence and rate of extraction. Researchers from private industry and the National Institute for Occupational Safety and Health have studied mechanisms of violent failure and have identified individual factors that contribute to coal bumps. In an attempt to develop tools for assessing coal bump potential, the authors initiated a comprehensive study using information from 25 case studies undertaken in U.S. mines. Multiple linear regression and numerical modeling analyses were used to identify the most significant variables contributing to coal bumps (excluding bumps related to fault-slip). Twenty-five geological, geometrical, and geomechanical variables were considered initially. The most important of these variables were then identified as (1) energy as calculated using the mechanical properties of the strata, depth of overburden, and joint density, (2) mining method, (3) pillar factor of safety, and (4) stress gradient and yield characteristics.

By M. E. Poad, M. M. McDonald, J. C. Johnson, W. G. Pariseau

A U.S. Bureau of Mines case study of pillar recovery in high-grade ore near the Ross shaft at the Homestake Mine, Lead, SD, has demonstrated the usefulness of the finite-element method for evaluating shaft pillar mining plans and shaft stability. In this study, borehole extensometers and other instruments were installed to provide data for model verification and to monitor shaft displacement. Results of a recalibrated two-dimensional model (UTAH2)confirmed the premining stability evaluation. However, after mining began, concern developed because of cracks and other signs of ground motion that appeared at considerable distances from the area of active pillar mining. An intense three-dimensional modeling effort using the computer modeling program UTAH3 was initiated. The results again showed the observed effects were within expectations and that the shaft would remain safe. Three-dimensional analyses of alternate pillar mining scenarios indicated that more of the shaft pillar ore reserve could be recovered than previously thought on the basis of two-dimensional analyses. This report focuses on startup and implementation of the three-dimensional model, calibration of the three-dimensional model, and application of the model to shaft pillar mining.

Jan 1, 1996

By M. L. Davis

The principal causes of many methane-gas explosions that have occurred in anthracite mines, especially in thick pitching veins, during recent years can be traced to lack of a satisfactory method of testing for such gas in areas adjacent to those in which blasting is to be done or other sources of ignition introduced. As coal-mine officials and inspectors are required to make examinations for methane in all accessible open workings, including high places, and permissible flame safety lamps have been used almost universally for making these examinations, considerable difficulty has been experienced in proper testing of atmospheres at elevations beyond the normal reach of the examiner, and resultant explosions have often indicated perfunctory tests. The accepted practice of testing for gas in voids or high places generally consists of raising a flame safety lamp, suspended from the end of a long drill, stick, or pipe, into the atmosphere to be tested. Disadvantages noted are: Action of the flame within the lamp cannot be seen readily when the lamp is more then 18 inches overhead, the flame will not be extinguished when less than about 4-1/2 percent of methane is present, and indications of small percentages of methane within the lamp may disappear entirely while the lamp is being lowered slowly to view; the examiner must usually be exposed unnecessarily to falling material; and the lamp may fall and be broken. Thus, the need for a more reliable, safe method of testing for gas in out-of-reach places is apparent. Several electrical methane detectors, tested and approved by the Bureau of Mines, have been used satisfactorily, though infrequently, for making the aforementioned gas tests. Some of these detectors employ a rubber aspirator bulb and rubber tubing of suitable length to draw the atmosphere from high places through and over energized parts of the detector. However, the original expense, susceptibility to damage, and awkwardness of tarrying these instruments in steep-pitch workings has precluded their general use in everyday examinations. Although not as efficient as the electrical detector for determining small percentages of methane or as reliable because of the limitations imposed by oxygen-deficient atmospheres, the modern per¬missible flame safety lamp can, it is believed, be used safely with the rubber aspirator bulb and tubing to greater advantage and with comparative reliability in testing for gas in out-of-reach places. Slight reference to this method of introducing air into a safety lamp is made in a Bureau of Mines publication of l921,2/ and it is quite possible that the method has been utilized in some manner by individuals in American mines during the intervening years. A similar method was used in connection with a nonpermissible Wolf flame safety lamp in at least one coal mine in Japan before World War II. A major change in the construction of this lamp would bar its approval by the Bureau of Mines. An American adaptation of the Wolf lamp, developed

Jan 1, 1954

It was decided to demonstrate the use of latex as a soil sealant by irrigating three small plots at Lanse coded 03A on the west of Brown's run, 3C on the east of the stream, and SB, an area on the spoil bank. Earlier plans to sprinkle latex on twenty acres of the side were abandoned because of easement difficulties. On the basis of results from laboratory soil columns and small plot experiments at Lanse, Naugatex J-3471 was the latex of choice. The irrigation system consisted of plastic pipe (two inch size) for the main line from the water meter to the irrigation area because of lower price and ease of handling. Because the municipal water supply was limited to 20 gpm, low-volume 14V-LA-TNT Rainbow sprinklers were specified with low angle 20° 3/32 inch nozzles to reduce wind effects. At the expected nozzle pressure of 50 psig, each sprinkler covers a 56 foot diameter area at a delivery rate of 1.64 gpm. Each of the three test areas was covered with twelve sprinklers, six used with water alone as a control area and six used with latex and subsequent water spraying. One-quarterturn inlet valves were positioned on the main lines leading to each area. Flow to the treated and untreated areas was set at 10 gpm by Dole flow control valves in the' main entry lateral to each of the six sprinklers. To avoid the use of dilution tanks, a Moyno pump operated by a 1/2 HP, 1140 rpm 5:1 reducing drive was used to deliver 1/2 gpm 48% latex for mixing with 10 gpm water. A 3500 watt gasoline generator provided power for pump and light. Latex was applied at night to take advantage of the lower evaporation and higher humidity. Neutron gauge access tubes were placed in orthogonal arrangement in the areas that were irrigated, with four tubes in the section that were sprinkled with latex and four tubes in the untreated irrigation section in each of the three locations. Each tube, five and one-half feet long and sealed at one end with a plug welded in place, was inserted five feet into the ground in a cored hole. In each irrigation area, the left section of six sprinklers spread the dilute latex while the right section sprinkled the same amount of water to serve as a control plot. As shown in Figure 7 each sprinkler has a 56-foot sprinkling diameter at 50 psig design water pressure. The sprinklers are 30 feet apart on the laterals and 35 feet apart between the laterals. The access tubes (coded AT on Figure 7) were placed in comparable spots on the two sections of each area, so that four soil moisture measurements on the two sections could be compared. Bouyoucos resistance blocks (BB) were also placed in areas 03A and 3C at one, two and three foot depths, but not on the spoil bank area.

Jan 1, 1972

Objective To characterize the roof fall risk associated with underground limestone mining. Background During the first 5 years of this decade, more than one-fourth of the underground mining fatalities occurred as the result of fall of ground. The large openings in underground stone mines present limitations in observing ground conditions because of floor-to-roof heights ranging from 20 to 60 ft or more. Generally, the status of ground conditions in under-ground stone mines is based on observation and experience obtained during mining development. Some mines use monitoring instruments to gain information on roof conditions, but this practice is typically focused and localized to address issues in a particular area or section of the mine. The Roof Fall Risk Index (RFRI) described here provides a method of systematically characterizing the risk of roof falls to enhance the safety of mine workers. Approach As in many industries, a defective product signals the potential for failure. Similarly, defective rock or strata in an underground stone mine presents a potential risk. Assessing ground conditions relative to the concentration of defects observed provides a way to index or rate these conditions and the subsequent risk. Through observations at approximately 50 underground limestone mines in the United States, strata defects were categorized to develop a quantitative method for determining the RFRI. This method is specifically geared to underground lime-stone mines because strata defects are difficult to see and limited by the experience and knowledge of the observer. Overall, 10 defect categories were deter-mined: 3 that relate to geology, 4 that result from mine development, 2 that are indicative of the roof profile, and 1 that addresses moisture or water flow (see RFRI Chart 1). The defect categories are also weighted to reflect the conditions indicative of roof or strata movement or separation (categories 4 through 8). These weights are multiplied by the assessment value to determine the total category value. The benefit of using monitoring techniques to gain information on roof conditions is also factored into the RFRI for overall assessment of an area in the mine. Each defect category is assessed with a value of 1 to 5 based on parameters as defined in each defect category, with 5 indicating the highest degree or level of defect. The assessment value of 3 is also used when information on a parameter is unknown. The parameters for the defect categories are illustrated in RFRI Chart 2 for use in determining the assessed value. How It Works The RFRI provides a means to categorize the risk for roof falls in underground stone mines. The concept of this calculation is aimed at providing an indication of the level or degree of defect relative to the area of the mine examined. The RFRI for the calculation produces a distribution where RFRI values approaching 0 represent a stable condition and those ap-proaching 100 represent an unstable condition.

Oct 1, 2006

Acid pollution from abandoned mines causes serious contamination of major streams in coal mining areas. In Pennsylvania alone it is estimated that 3000 miles of streams are devoid of fish and plant life because of coal mine effluent.1 The pollution arises because water and oxygen react with iron sulfide in the mine cavity to produce sulfuric acid and iron sulfate, which later find their way back into surface streams. Two general approaches to the problem are possible: (1) prevent the acid-forming reaction from taking place, (2) treat the acid mine water after it is formed so that it will not be harmful to the ecology when discharged into streams and other surface waters. The work described in this report belongs in the first category - it was based on the assumption that acid formation in the mine cavity can be prevented by keeping water (one of the reactants) out of the mine. This was to be accomplished by-forming a waterproof seal over the mine cavity which would prevent the seepage of surface water into the mine. It was planned to form such a seal at or near the surface of the ground by applying latex to the soil. Ideally a dilute latex would be applied to the soil and would penetrate to a depth of two to three feet below the surface where it would be coagulated by reacting with a constituent of the soil or by other means., A seal below the surface of the soil would have several advantages over a surface seal - it would not be subject to destruction by traffic, frost or vegetation and, more important, it would allow the use of the top soil for many agricultural purposes. A preliminary feasibility experiment was performed prior to application for the present grant. It showed that it was indeed possible to form a sub-surface seal by means of latex. A core of soil was obtained locally (Wayne, N. J.) by driving a 6.5 cm inside diameter steel pipe about 100 cm long into subsoil for a distance of 71 cm. The soil was retained within the pipe for this experiment. A part of the free space above the soil in the pipe was charged with water, and the seepage rate of the soil after it had become saturated was 15 ml/min under a total head (wet soil plus supernatant column of water) of 92 cm. After the supernatant water had drained completely, 100 ml of 5% rubber latex was added and allowed to seep into the soil. This was followed by additional water and the seepage rate was found to be 2 ml/min under the same head. This seepage rate remained unchanged when remeasured on the following day. Since there was no visible evidence of the latex in the collected drain water it is presumed to have been retained by the soil.

Jan 1, 1972

By George S. Rice, O. P. Hood

The need of protecting mines from the danger of inflow of natural gas from neighboring wells has become more apparent each year since it was found that oil and gas underlie the productive coal measures of Pennsylvania, West Virginia, Ohio, Indiana, Illinois, Kansas, Oklahoma, Colorado, and Wyoming, and, to some extent, the coal measures of other States. The proximity of the known gas and oil areas in the United States to the coal fields and the manner in which the fields overlap in certain States are shown in the map comprising the frontispiece (Pl. I). As the result of many requests received from various sources a conference to consider the protective measures necessary to insure the safety of mining operations in gas and oil fields was called by the Director of the Bureau of Mines to assemble in Pittsburgh, Pa., Feb- ruary 7 and 8, 1913. To this conference were invited a considerable number of gas-well, oil-well, and coal-mine operators within conven- ient distance from Pittsburgh, as well as State geologists and State mine inspectors from a distance. Gas and oil wells are also found near other kinds of mines, such as shale and limestone mines in western Pennsylvania, but these cases are so few that for this and other reasons it was deemed best to con- fine the present inquiry to the consideration of the questions involved in connection with coal mines. It was also not thought advisable to consider the protection of mines from other kinds of bore holes than those for gas and oil, such as coal-prospecting holes, drilled either with churn or diamond drills, or salt wells and artesian wells, because little trouble has been ex- perienced in mining operations from these classes of wells. There- fore the following papers and discussions concern almost entirely the following relations between coal mines and gas and oil wells: (1) Protection of coal mines froin gas and oil wells penetrating -through: (a) Open workings; (b) inaccessible spaces or goaves; (c) pillars, or coal which may be left as pillars. (2) The manner of drilling gas and oil wells passing through workable coal beds so as to provide protection for future mining. The preliminary inquiries of the engineers of the bureau had dis- closed that there was no uniformity in the methods of protecting mines against the leakage of gas from wells, that thousands of wells in coal fields are abandoned yearly without adequate plugging, and without surveys or records being kept to show the location of the holes, from which the casings are generally pulled. According to the State geologist of Pennsylvania, as many as 3,000 wells are being drilled annually in Pennsylvania, 2,000 of these holes being aban- doned within a single year. It is such unplugged, uncharted wells that are the greatest menace in mining. The quantity of gas that the wells produce, although not sufficient to be commercially avail-

Jan 1, 1913

By Fred N. Kissell, Jay F. Colinet

The two major lung diseases of miners are "blacklung" and silicosis. Blacklung, or pneumoconiosis, was first recognized as a disease of British coal miners in the 1600s. However, investigations into the cause of blacklung disease did not begin until the 1900s. The causative agent of pneumoconiosis in coal miners was thought to be silica until studies in the United Kingdom provided evidence that expo- sure to coal dust containing minimal silica could also cause pneumoconiosis. In the Federal Coal Mine Health and Safety Act of 1969, coal workers pneumoconiosis (CWP) was defined as "a chronic dust disease of the lung arising out of employment in an underground coal mine." Diagnosis of CWP is generally based on chest x-ray findings and a patient's history of working in coal mines, usually for 10 or more years. In its early stages, CWP is called "simple CWP." Miners with simple CWP are at increased risk of developing an advanced stage of the disease, called "progressive massive fibrosis" pMF) or "complicated CWP." PMF (complicated CWP) is associated with significant decreases in lung function and oxygen- diffusing capacity. PMF is also associated with breathlessness, chronic bronchitis, recurrent chest illness, and episodes of heart failure. Coal miners with silicotic lesions or PMF have an increased risk of tuberculosis and other mycobacterial infections. PMF may progress even in the absence of further dust exposure. This disease is also associated with increased mortality (National Institute of Occupational Safety and Health [NIOSH] 1995). The most recent study to determine the prevalence of CWP in coal workers was in the mid- 1990s. The average prevalence of CWP was 2.8%, but for those with 30 or more years in the industry it was 14%. For the 10-year period from 1987 to 1996, 18,245 deaths were attributed to CWP. The most frequently recorded occupation on the death certificate (70%) was mining machine operator (NIOSH 1999). Silicosis may develop when inhaled respirable crystalline silica (quartz) is deposited in the lungs. The clinical diagnosis of silicosis is based on (1) recognition by the physician that the level of silica exposure is adequate to cause the disease, (2) the presence of chest radiographic abnormalities consistent with silicosis, and (3) the absence of other illnesses (e.g., tuberculosis or pulmonary fungal infection) that may mimic silicosis. The radiographic patterns are often the same for CWP and silicosis; thus, these diseases are sometimes distinguishable only by work history or pathological examination. Chronic silicosis commonly involves 15 or more years of exposure to silica. The characteristic microscopic feature is the silicotic nodule. Chronic silicosis is often asymptomatic and may manifest itself as a radiographic abnormality with small, rounded opacities of less than 10 mm in diameter, predominantly in the upper lobes. Lung function may be normal or show mild restriction. Chronic silicosis is also associated with a predisposition to tuberculosis and other mycobacterial infections and with progression to complicated silicosis.

By Robert E. Hershey

1. The investigation shows that the Sewanee coal bed is the most important bed in Grundy County arid is the only bed that has had any sustained production. There is no production, at the present time, from any of the other coal beds in the county. The Richland (Lower Sewanee or Coke Oven) coal bed has been prospected but has never been developed extensively. The Angel coal bed has been worked to a very small extent near Beersheba Springs, Tenn., for local consumption. The Battle Creek coal bed has been worked sporadically near Monteagle, Tenn. 2. Known measured and indicated reserves in the Sewaneee coal bed, based on a minimum thickness of 14 inches and on 1,800 short tons per acre-foot of coal in place, are estimated to be 15 millon short tons as of January 1, 1953. Of this total, 10 million tons represents coal 28 inches and more thick. Areas in the bed were omitted from the estimate where available data relative to the bed characteristics are too meager for making an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1955

By James Paone

The Bureau of Mines conducted laboratory drillability studies on nine rock types using two mining-type percussive drills. Percussive-bit action was simulated by dropping a known weight on a rock sample to obtain a coefficient of rock strength for each rock drilled. While no single physical property of a rock is completely satisfactory as a predictor of penetration rate, coefficient of rock strength showed a better correlation with penetration rate than did other physical properties. Results also indicate that over the range of energies investigated, energy per unit volume in a given rock is independent of blow energy. A drillability index for the percussive system is proposed that utilizes the coefficient of rock strength test to determine energy per unit volume for a given rock which, along with the work rate of a given drill and the bit diameter, is used to predict the penetration rate as shown in the following equation: [R = 4P pD2 E, where R = penetration rate, in/min, P = work rate, in lb/min, D = bit diameter, in, and Ev = energy required to produce a given volume of rock, in lb/in3.] Laboratory tests indicate this method's potential as a simple and reasonably accurate method of predicting penetration rates with any percussive drill and bit combination.

Jan 1, 1969

By Steven J. Page

The increasing concern for dust control in metal and nonmetal mining prompted the Bureau of Mines to investigate the dust suppression technique of applying foam to face drilling and crushing operations in a gypsum mine. Previous attempts to achieve effective dust control on face drills using other techniques, such as dry collection and boom-mounted external water sprays, have not been successful. The largest problem encountered when using these techniques has been the lack of sustained effective dust capture at the face. Results for two faces studied show that injection of the foam through the drill steel provided an average dust reduction of 95 pct. However, the use of foam on the crusher provided no dust reduction at the operator's position (because of good local ventilation) and a 27-pct average dust reduction for two feeders at a location midway between the operator's position and the crusher. The conclusions obtained from this study are that (1) foam can be an extremely effective dust control technique when used on localized dust sources such as dust generated within a blasthole during drilling, and (2) foam is only marginally effective as a dust control technique when used on nonlocalized dust sources such as dust generated during the feeder-and-crusher operation.

Jan 1, 1981

By Iam Pou Ng

Combustion aerosols consist mainly of elemental and organic carbon (EC and OC). Since EC strongly absorbs light and thus affects atmospheric visibility and radiation balance, there is great interest in its measurement. To this end, the National Institute for Occupational Safety and Health (NIOSH) published a standard method to determine the mass of EC and OC on filter samples. Another common method of measuring carbon in aerosols is the aethalometer, which uses light extinction to measure "black carbon" or BC, which is considered to approximate EC. A third method sometimes used for estimating carbon in submicron combustion aerosols, is to measure particle size distributions using a scanning mobility particle sizer (SMPS) and calculate mass using the assumptions that the particles are spherical, carbonaceous and of known density. With an eye toward evaluating the use of these methods for measuring carbon in freshly emitted diesel aerosols, the objective of this paper is to compare the results among these methods and to investigate the role of OC when measuring the mass concentration of elemental carbon (or black carbon) in aerosols emitted by a medium-duty Diesel engine. Results indicate that the aethalometer response correlates well with the data from filter samples (R2 = 0.99), with a very slight positive interference at high levels of OC.

Jan 1, 2008

By Elizabeth A. Johnson

The effect of slag composition on lead losses to slags and on the behavior of associated accessory elements in lead blast furnace smelting has been studied by the Bureau of Mines. Equilibrations of 30 g of lead and synthetic slag doped with Cu, In, TI, and Sb at 1,200° C in CO-C02 atmospheres revealed the importance of the FeO:Si02 ratio and the basicity ratio, (CaO+MgO):SiO2, of the slag. Lead losses to slags increased with an increasing FeO:Si02 ratio and decreased with an in-creasing basicity ratio. Copper and indium distribution coefficients, [Lx =] increased from 15 at the lower basicity ratios to near 60 at higher basicity ratios. LTI values increased from 20 to 150 as slag basicity increased. Lcu, Lin, and LTI were relatively in-dependent of the FeO:Si02 ratio, but increased with greater amounts of accessory elements in the system. Antimony distribution alone was shown to be strongly affected by FeO:Si02 ratio, although data were scattered. LSb values ranged from 20 to 800 and decreased sharply with increasing FeO:Si02 ratios. The basicity ratio and the level of accessory element load did not have any noticeable effect on LSb.

Jan 1, 1983

By Tobias W. Goodman

The U.S. Bureau of Mines used a contact drilling strategy with short-collared assemblies in order to rotary drill long horizontal methane drainage holes in the Beckley Coalbed near Glen Daniel, WV. By decreasing the thrust, and increasing the rotation of the drill bit when in contact with the roof and floor rock, assemblies with 1-ft (0.3-m), 10-ft (3- m), and 144 (43-m) long collars were made in deflect and stay in the coalbed. Successful application of this procedure resulted in holes drilled full length in coal with the original assembly including bit. The test assemblies deflected away from the original bearings of the holes during drilling and arced. Four holes were drilled to depths ranging from 478 ft (145.7 m) before the strategy was developed, to 4,034 ft (1,230 m) when the strategy was applied. Total length of the holes in coal was 8,590 ft (2,618 m). After 577 days the total methane captured by the holes was 221,600,000 ft? (6,276,000 m3). The methane flow rate peaked at 805,000 ft3/d (22,798 m3/d), then decreased to stabilize at 670,000 ft?/d (18,974 m3/d) after 6 months.

Jan 1, 1989