Abstract: Recent NIOSH published information has shown an increase of rapidly progressive coal workers’ pneumoconiosis (CWP) in the southern Appalachian coal region (SAR) of the U.S., despite the fact that compliance data indicates that most coal miners have been exposed to coal mine dust concentrations below the statutory limit of 2.0 mg/m3. While the exact cause of these elevated CWP levels in the SAR has not been established, several factors may be contributing to the increase in occupational lung disease among coal miners. The mining of high rank coal is known to lead to higher CWP rates, and this type of coal is mined in a portion of this region. Also, a high percentage of the mines in the region are on reduced dust standards because of the high silica content of the airborne dust resulting in miners possibly being exposed to excessive amounts of respirable silica dust. Exposure to excessive amounts of respirable silica dust can lead to silicosis, a disabling and potentially fatal lung disease. NIOSH's Respiratory Hazards Control Branch has been investigating the possible causes that would account for the observed higher trends in disease progression through literature review, data analysis, and inmine surveying. The investigation to date has revealed that underground mines are faced with cutting large amounts of rock in order to maintain haulage clearances. Cutting rock not only increases the potential for silica exposure, it increases the requirement for machine maintenance which was a concern observed during dust surveys conducted by NIOSH. Over half of the mines operating in the SAR are on reduced standards due to high silica content. Adequate face ventilation of the continuous miner and roof bolter and limited down-wind operations from the miner are also issues of concern and items which require the constant attention of miners operating in these conditions. Disclaimer: Mention of any company or product does not constitute endorsement by the National Institute for Occupational Safety and Health.

Abstract: A majority of continuous mining machines employ a water spray system and a machine mounted flooded-bed scrubber to suppress and capture dust during coal mining. These machine mounted dust control systems must be designed to function within the localized face ventilation system at the mining section to control both dust and methane. Spray systems can impede or improve the scrubber effectiveness in controlling dust or methane at the mining face. Laboratory experiments were conducted to examine the effect of spray type, spray pressure, machine body blocking sprays, and scrubber airflow on dust and gas levels while using a 12.2 m (40 ft) exhaust ventilation curtain setback from the face. These experiments were conducted with the mining machine positioned at the end of a simulated 6.1 m (20 ft) sump and slab cut. Results indicate that the hollow cone nozzles with blocking sprays best complemented the flooded-bed scrubber performance in an exhaust ventilation system. This external spray system notably reduced dust and gas levels on the off-curtain side of the mining machine for both the sump and slab cut as compared to the flat spray nozzles. Higher scrubber airflows reduced dust and gas levels on the curtain side and in the return of the continuous mining machine. The remote operator position, located on the off curtain side and parallel to the inlet end of the exhaust curtain, sustained the most stable and lowest dust levels around the mining machine.

Abstract: An assessment of U.S. western hardrock miners’ noise exposures was conducted as part of a multi-year National Institute for Occupational Safety and Health (NIOSH) survey of noise exposures in each sector of the mining industry. Noise from selected mining equipment and operator noise exposures were measured, analyzed and tabulated for dissemination to the participating sites and are being used to direct NIOSH research and interventions to address the greatest noise hazards. Eighty-two noise dosimeter measurements were obtained, along with time-motion studies as the miners operated hardrock mining machines. Of the operators, 96% had daily noise doses that exceeded the U.S. Mine Safety and Health Administration’s (MSHA) permissible exposure level. The average gold miner dosages while operating the following equipment were: haul trucks 261%, load-haul-dumps (LHDs) 235%, single boom drills 221%, bolters 214% and dual boom drills 163%. The worst exposure level was a silver miner with a daily dose of 873%. Time-motion data showed that this miner’s exposure accumulated most rapidly while operating a jackleg drill. These results will be used to help prioritize noise control development by NIOSH and other partners. Introduction Hearing loss and overexposure to noise continues to be a problem throughout the mining industry. Studies indicate that 70 to 90% of all miners have a noise-induced hearing loss (NIHL) severe enough to be classified as a hearing disability by the time they reach retirement age (Franks, 1996). To address this problem, as part of an ongoing strategic plan, the National Institute for Occupational Safety and Health (NIOSH) has been conducting a series of long-term noise surveillance and evaluation studies for the entire mining industry. Operator noise exposures were measured and tabulated for dissemination to the mine sites and analyzed to direct future noise control development. This report contains the description of studies conducted at U.S. western hardrock mines to determine the levels of miners’ noise exposure. Eighty-two noise dosimeter measurements were obtained from hardrock mining machine operators. Time-motion studies were performed as the operators used bolting machines, LHDs, haul trucks and drills. Of the operatiors, 96% had daily noise doses that exceeded the U.S. Mining Safety and Health Administration’s (MSHA) permissible exposure level (PEL). The only participants with exposures below the PEL were two bolter operators and one dual-boom drill operator. Approach to quantify noise exposure To evaluate the operators’ noise exposure, a sample population of each piece of equipment was monitored and the data analyzed using the following two research methods. 1. � Noise dosimetry was used to measure the machine operators’ total noise exposure during the course of their work day. 2. � Time-motion studies were used to identify work tasks and/or machines causing the higher doses that are in need of engineering noise controls or administrative controls. To complete these analyses, NIOSH researchers conducted noise surveys at eight underground metal mines located in the western U.S. to identify and quantify the noise exposure of hardrock underground machine operators. Noise dosimetry measurements The operators’ noise exposure was measured using a Quest Q-400 noise dosimeter. The dosimeter microphone was clipped to the midpoint of the operator’s shoulder with the diaphragm pointing up and worn for a full shift. The Quest Q-400 (Quest Technologies, 1997) is a single-microphone, dual-channel device that allows for independent user-configurable dose evaluation settings on each of the two channels. The two channels, referred to as dosimeter 1 and dosimeter 2, collect sound level measurements simultaneously. Dosimeter 1 was set according to the MSHA PEL (MSHA, 1999), and dosimeter 2 was set for wide-range data collection to measure and record all sound levels (Table 1). The MSHA PEL settings were used for consistency with the majority of noise dosimetry data reported for the U.S. mining industry. The wide-range data was collected for future analysis as part of a NIOSH equipment operators’ noise exposure database. Table 1

Abstract: To reduce fire hazards from spontaneous combustion of coal in longwall gob areas, a series of computational fluid dynamics (CFD) simulations were conducted by the National Institute for Occupational Safety and Health (NIOSH) to model the spontaneous heating of coal in longwall gob areas. The previous modeling results demonstrate that spontaneous heating of coal usually occurred behind the longwall shields and along the face with a bleeder ventilation system. Assuming a stationary longwall face, the spontaneous heating could turn to a spontaneous fire in several days for the most reactive coal under favorable conditions. When the longwall face advances, the spontaneous heating process will be significantly affected. In this study, the effect of longwall face advance on the spontaneous heating in the gob area is investigated using the CFD model developed in previous studies. One longwall panel with a bleeder ventilation system is simulated. The width of the panel is 300 m (984 ft), while the length of the panel is changed between 1,000 to 2,000 m (3,280 to 6,560 ft). The same permeability and porosity profiles are used for gobs with different lengths. The spontaneous heating first develops in the gob when the longwall face is stationary. Then, the face advances at a certain rate. The face advance is simulated as a series of discrete movements, and the effect of the face advance on the maximum temperature developed during the face stoppage is examined. & Glasser, 1986; Carras & Young, 1994; Edwards, 1990; Nordon, 1979; Krishnaswamy et al., 1996; Rosema et al., 2001), most of the research is mainly focused on small coal stockpiles without considering the effect of ventilation air flow. Saghafi et al., (1995, 1997) did numerical modeling of spontaneous combustion in underground coal mines with a back return U-ventilation system, but their work was limited to two dimensions. Balusu et al., (2002) conducted a computational fluid dynamics (CFD) study of gob gas flow mechanics to develop gas and spontaneous combustion control strategies for a highly gassy mine. These strategies include new gob hole designs to ensure that oxygen concentration in the holes was below 4 to 5%, immediately sealing of the crosscut behind the face, and reduction in air velocity on the intake side of the gob. In order to reduce the fire hazard from spontaneous combustion of coal in gob areas, the National Institute for Occupational Safety and Health (NIOSH) has conducted a series of CFD simulations using coal kinetic data from the former U.S. Bureau of Mines laboratory-scale experimental results (Smith and Lazzara, 1987). Previous CFD models were developed to simulate the spontaneous heating of coals in a two-panel gob area using a bleeder ventilation system with a stationary longwall face (Yuan and Smith, 2007). Parametric studies were then conducted to examine the effects of coal’s activation energy, coal surface area, heat of reaction, different ventilation conditions and gob permeability distributions on the spontaneous heating process (Yuan and Smith, 2008a). CFD simulations were also conducted to model the spontaneous heating in longwall gob area using a bleederless ventilation system with a stationary longwall face (Smith and Yuan, 2008). In this paper, CFD modeling of the effect of longwall face advance on the spontaneous heating of coals in a two-panel gob area using a bleeder ventilation system is presented. GOB layout and ventilation system In this study, a single longwall panel, 2,000-m(6560-ft-) long and 300-m(984-ft-) wide, with a three-entry bleeder ventilation system is simulated. The layout of the panel and the ventilation system is shown in Fig. 1. The original coal seam is 2-m(6.5-ft-) high, and the gob is assumed to be 10-m(33-ft-) high starting from the bottom of the coal seam. The ventilation airways are 2-m(6.5-ft-) high and 5-m(16-ft-) wide. This scheme and the panel dimensions are typical of longwall mines operating in the Pittsburgh coal seam in Northern Appalachian Basin. The bleeder entries at the back end of the gob are modeled as one entry connecting to the bleeder fan. Three regulators are located at the end of the second and third headgate entries and the tailgate entry, respectively, for controlling the bleeder ventilation. The longwall face is originally at location #1, 1,000 m (3,280-ft-) from the start line of the panel. It is assumed that the face advances in a rate of 20 m/d (65 ft/day) from location #1. The face advances 100 m (328 ft) in five days to reach locations #2, #3 and #4, respectively. The face advances 200 m (656 ft) in 10 days to reach location #5. FIGURE 1 Layout of longwall panel and ventilation system. Low temperature coal oxidation The chemical reaction between coal and oxygen at low temperatures is complex and still not well understood. Generally, three types of processes are believed to occur in low temperature coal oxidation (Carras and Young, 1994), including physical adsorption; chemical adsorption, which leads to the formation of coal-oxygen complexes and oxygenated carbon-species; and oxidation, in which the coal and oxygen react with the release of gaseous products, typically carbon monoxide (CO), carbon dioxide (CO2) and water vapor (H2O). The moisture content of coal can play an important role in the low temperature coal oxidation process. The interaction between water vapor and coal can be exothermic or endothermic, depending on whether the water condenses or evaporates. Sondreal and Ellman (1974) reported that for dried lignite, the rate of temperature increase due to the adsorption of water increased with the moisture content up to a value of 20% water (by mass) and then decreased with further increasing moisture content. Smith and Lazzara (1987) found that the effect of the moisture content of the air on self-heating process was also dependent on coal rank and temperature. In this study, the effect of water vapor is not considered, and the chemical reaction between coal and oxygen is simplified as: Coal + O2 → CO2 + 0.1CO + heat The detailed chemical structure of coal is not clear and is believed to vary with the rank and origin of coal. According to experimental data (Smith et al., 1991), one mole of coal reacting with one mole of oxygen generates one mole carbon dioxide and 0.1 mole carbon monoxide plus heat at the early stage of coal oxidation. The dependence of the rate of oxidation on temperature and oxygen concentration can be expressed in the form: Rate = A[O2] n exp(-E/RT) where the chemical reaction rate is defined as the rate of change in the concentrations of the reactants and products, A is the preexponential factor (in K/s), E is the apparent activation energy (in kJ/mol) that is the energy needed to initiate a chemical reaction, R is the gas constant, n is the apparent order of reaction, T is the absolute temperature (in °K), and [O2] is the oxygen concentration (in kmol/m3). The value of apparent activation energy, E, of different coals can very between 12 and 95 kJ/mol. The preexponential factor, A, depends more on coal rank and measurement method, and has a typical value between 1 and 7×105 /s. The value of the apparent order of the reaction, n, in low-temperature oxidation studies of coal and other carbonaceous materials has been shown to vary from ~0.5 to 1.0 (Carras and Young, 1994), and is about 0.61 for some U.S. coals (Schmidt and Elder, 1940). The coal source in the gob can be coal left from the mined coal seam or other overlying coal seams. In this study, the Pittsburgh coal seam was considered, with a 1-m(3.3-ft-) thick rider coal seam less than 1-m(3.3-ft-) above the 2-m(6.6-ft-) thick main coal seam. The coal source in the model is this rider coalbed that is assumed to cave into the gob after the main coal seam is completely mined out.

Abstract: The general geological situation,ore body properties,marks for ore detection and cause for formation of Dagushan iron ore deposit are explained.

Abstract: Steady increases in longwall production have required operators to apply greater quantities of ventilating air in an effort to control and dilute respirable dust. Significant increases in shearer speeds necessitate that longwall supports also be advanced at a faster rate. Both of these factors may contribute to overall respirable dust levels on the longwall face because as supports are lowered and advanced, broken material falling from the top of the canopy is entrained directly into the air stream.To address this issue, the Pittsburgh Research Laboratory collected respirable dust samples from four longwall faces to characterize shield-generated dust. This paper investigates the influence of air velocity and shield advance rates on respirable dust levels. Also discussed are engineering controls currently used to reduce shield dust and alternative controls being investigated by the National Institute for Occupational Safety and Health (NIOSH). Introduction Reducing respirable dust levels on longwalls through the development of engineering controls remains a primary health concern for the National Institute for Occupational Safety and Heatlh (NIOSH), industry, regulators and university researchers. NIOSH researchers from the Pittsburgh Research Laboratory conducted longwall dust surveys in an effort to better understand the influence of both air velocity and support advance on shield-generated respirable dust. Ultimately, this research will lead to lowering worker exposure to respirable coal and silica dust. Increased production has put a greater demand on longwall dust control systems as operators have had difficulty maintaining consistent compliance with federal dust standards. During the five-year period 2000-2004, analysis of mine operator and U.S. Mine Safety and Health Administration (MSHA) inspector samples shows that the percentage of these samples exceeding 2.1 mg/m3 (the level at which a citation is issued) was 14% and 15%, respectively (Niewiadomski, 2004). Dust control technology employs both air and water as a means to limit, direct and eliminate respirable dust. Historically, longwall dust control research focused on reducing dust generated by the two largest sources: the shearer and stageloader. Increasing air velocity on the face is a fundamental control used to dilute dust generated by these sources. A minimum longwall face air velocity of 2 to 2.3 m/s (6.5 to 7.5 ft/sec) is recommended for proper dilution of dust along the face (Foster-Miller Assc., 1982; Breuer, 1972). A research study conducted during the1980s found that face air velocities ranged from 0.6 to 3.3 m/s (2 to 11 ft/sec), while average production was 735 t/shift (810 st/shift) (Jankowski and Organiscak, 1983). The shearer and stageloader combined accounted for 82% of the total respirable dust generated. A second study during the 1990s showed that the range of air velocities increased to 1 to 7.6 m/s (3.3 to 25 ft/sec), while the average production increased nearly four-fold to 2,885 t/ shift (3,180 st/shift) (Niewiadomski, 2004). Despite increased production, the total dust contribution of the shearer and stageloader decreased to 68% as a result of improved dust control measures. Increased ventilation, improved water spray application at the shearer and enclosing the stageloader and adding water sprays were the primary modifications that led to this reduction (Colinet and Jankowski, 1997). Despite the above advances in dust control, due to faster shearer speeds or the presence of methane gas, higher air velocities are often required for dilution. Past studies (Hall, 1956; Hodkinson, 1960) have shown that, for low moisture coals, velocities in excess of about 2.3 m/s (7.5 ft/sec) can increase dust levels. More recent studies (Tomb et al., 1991; Breuer, 1972) have shown that with adequate moisture on the coal, such as that provided by sprays, airflow may be increased beyond 5.1 m/s (16.7 ft/ sec), without significantly increasing the dust levels on the face. In these studies, the shearer and stageloader were being shielded from the face airflow by physical barriers, and/or water spray systems were being used to increase moisture and promote particle agglomeration. Dust sampling on these longwalls clearly indicates that if moisture content is adequate, increasing air velocity does not promote dust generation from controlled sources, such as the shearer or stageloader. Studies conducted by the U.S. Bureau of Mines in the 1980s investigated support-generated dust as it related to geology, support design and cutting sequence (Organiscak et al., 1985). The studies found a correlation between support-generated dust and the geology of the immediate roof strata. Studies conducted during the early 1990s showed that shield dust was a significant contributor to the total dust generated from all sources (Colinet et al., 1997, Colinet and Jankowski, 1997). Dust attributed to shield movement almost doubled from 12% in the 1980s to 23% in the 1990s. NIOSH studies show that the primary factors responsible for this increase are most likely a combination of higher production and lack of effective control technology for shield dust (Rider and Colinet, 2006). Higher production rates have led to increased shearer speeds, which requires that shields be moved faster and in greater numbers. As shield supports are lowered and advanced, broken coal and/or rock falls from the top of the shield canopy and is being introduced directly into the air stream ventilating the longwall face. Also, the amount of air required to ventilate longwall faces continues to trend upward (Rider and Colinet, 2006).This has the potential to entrain greater quantities of respirable dust during shield advance under certain conditions. Further complicating the issue, most longwall operations utilize a bi-directional cutting sequence. When mining from headgate to tailgate, shield movement upwind of the shearer also contributes to dust exposure of the shearer operators. In mines where roof rock is the main component on the shield canopy, the entrainment of respirable silica dust becomes more of an issue, since the quartz component of rock is the primary contributor to silica dust generation (Ramani et al., 1987). In laboratory studies, NIOSH conducted tests using a wind tunnel to study the basic behavior of dust dropped into an airstream under varying air velocities (Listak et al. 2001; Chekan et al., 2001; Chekan et al., 2004). The studies were designed to determine dust concentration as it relates to increasing both the air velocity and the shield rate of advance. Results from these laboratory tests showed two fundamental relationships for low moisture coals (<1%): 1) increasing the air velocity increases airborne dust concentrations and 2) increasing the rate at which respirable dust is introduced into the airstream (faster shield advance rate) also increases airborne dust concentrations. Tests in the wind tunnel were conducted at air velocities of 2 m/s, 4.1 m/s, 6.1 m/s and 8.1 m/s (0.6 ft/s, 1.2 ft/s, 1.8 ft/s and 2.5 ft/s). Respirable dust concentrations increased as air velocity increased in a linear relationship, indicating that particle entrainment was greater than dilution effects. Respirable dust concentrations increased from 1.47 mg/m3 at 2.0 m/s (0.6 ft/s) to 19.84 mg/m3 at 8.1 m/s (2.5 ft/s). Statistical analysis of the concentrations measured at each velocity resulted in significant differences at a 95% confidence interval. Analysis of the size distribution of the sampled dust showed an inverse relationship between velocity and mass median diameter of the dust. As velocity increased, the mass median diameter of the entrained dust particles decreased in a near linear manner. The mass median diameter was found to be 10.8 microns at 2.0 m/s (400 fpm) and decreased to 7.7 microns at 8.1 m/s (1,600 fpm). Higher concentrations and finer particle size distributions suggest that at a moisture content of approximately 1, a portion of the dust particles were adhering to each other at the 2.0 m/s (400 fpm) velocity. As the velocity increases, the adhering forces are overcome by the increased energy supplied to the system, resulting in higher concentrations and smaller particle sizes in the airstream. Although the particle size decreased with increased velocity, dilution was not observed in any of the tests. In summary, the tests showed that adhesion of respirable dust to larger material and to other respirable particles plays an important role in determining the fraction of airborne dust that is measured as respirable. To further investigate the relationships from these laboratory studies, a research effort was conducted in the field to characterize shield generated dust at four longwall mines. The longwall surveys would help ascertain if shield dust generation may be influenced by face air velocity and/or the rate of shield advance. Figure 1 Percent weight distribution by size for coal samples collected from shield canopy at mines A and C.

Abstract: The National Institute for Occupational Safety and Health (NIOSH) conducted field tests to evaluate the effectiveness of a wet head continuous mining machine for reducing dust exposure for continuous miner operators. Wet head technology delivers water via sprays to the continuous miner's cutter head as opposed to traditional standard sprays located on the boom and body of the mining machine. The sprays, positioned directly behind each bit on the cutter head, deliver water at the point of attack, serving to cool the bits during mining to reduce the potential for frictional ignitions. The sprays also flood the coal with water to potentially suppress dust generation. Dust surveys were conducted at several mines to evaluate the wet head's effectiveness to control respirable dust exposure at the continuous miner operator location and in the immediate return. Results show that the wet head miner improved air quality at both locations to varying degrees in some cases and not in others when compared to a continuous miner with a standard spray system.

Abstract: Numerous fatal/nonfatal accidents involving underground workers struck by powered machinery occur yearly, with continuous miners and roof bolters involved in the majority of these accidents. In an attempt to reduce these accidents, researchers at the the National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Research Laboratory conducted studies of operator interactions with the motions of continuous mining machines and roof bolter boom arms. These operators generally perform their tasks in close proximity to the equipment in confined workspaces, often employing awkward postures. Since experiments with human subjects using the actual equipment were not feasible due to safety concerns, researchers opted to conduct controlled experiments in laboratory settings. Utilizing motion capture technologies and computer simulations, the studies collected data on equipment and operator movement in a range of seam heights and working postures. This article details the results of these studies to examine operating speeds based on usage and seam height. Introduction The U.S. Mine Safety and Health Administration (MSHA) reports that both fatal and nonfatal remotecontrol continuous miner accidents from 1999 to 2007 averaged 243 per year during routine mining activities, with the majority of accident victims working within the turning radius of moving continuous miners. The mining industry uses an educational aid called “red zones are no zones” (MSHA 2006) to help operators of continuous miners to understand which areas around the machine to avoid. However, fatalities and injuries continue to occur. During the same period, MSHA data shows an average of nearly 600 accidents per year related to roof bolting, accounting for nearly one-third of powered machinery accidents. There are currently no regulations or data on determining safe velocities for roof bolter boom arms operating in close proximity to workers in underground mines (MSHA, 1994 and Turin et al., 1995). Other industries, such as robotics, have conducted studies and implemented guidelines for safe machine velocities. The U.S. Department of Occupational Safety and Health Administration (OSHA, 1987) and the U.S. Department of Energy (DOE, 1998) recommend velocities be limited as low as 152 mm/sec (6 in./sec) for manufacturing robots during programming. Another study, (Etherton, 1987), recommends 254 mm/sec (10 in./ sec) as a safe velocity for robots operating in proximity to humans. To address these issues, the National Institute for Occupational Safety and Health (NIOSH) conducted studies to investigate safe operating speeds for these two machines. To minimize risk to the human subjects, the studies used motion capture technologies and computer simulations in a controlled laboratory setting. For the continuous mining machine study, the researchers attempted to determine how quickly subjects could escape from the perceived danger of a machine rotating towards them. Continuous mining machines move in a straight line relatively slowly, but their rotational speed can be appreciably faster when the tracks move in opposite directions. In that case, the machine pivots around the center point of its tracks. The front and back of the machine are far enough from this center to result in high velocities in the turning direction. Researchers conducted numerous motion capture sessions with human subjects in controlled laboratory environments. This data was used to create a virtual environment to analyze factors influencing struck-by accidents during the tramming of a continuous mining machine. The roof bolter boom arm motion studies had United Mine Workers of America (UMWA) volunteers operate a model of a Fletcher Ranger II left-hand roof bolter arm while recording the motions of the bolter boom arm and subjects. While this manufacturer has the majority of the underground roof bolter market, the use of this particular manufacturer’s design should not limit the application of the results of these studies because other manufacturers’ articulating boom arms use similar designs. NIOSH’s model is of the common 1,830 mm (72 in.) length arm and all testing was done with full sump extension. For the vertical boom speed research, a virtual environment with digital humans was created to run simulations. A level of randomness added to the motion of the digital operator during the drilling and bolting cycle enabled the simulation to realistically represent the operator’s motions while performing the bolting task. Volberg and Ambrose (2002) discuss in detail the development of these random motions. The simulator output produced data on the number of contacts between the operator and moving boom arm. This data was analyzed to show the effects of vertical speed on the risk of boom arm contact with the operator. For the horizontal boom speed study, it was only necessary to complete analysis of motion capture data from the human subjects testing to determine if an operator’s safety could be compromised by excessive speeds. Procedures The continuous mining machine investigation analyzed factors influencing struck-by accidents during tramming by using a digital human model (DHM) with simulations driven by captured human motion data with a variety of subjects, postures, facing orientations, environmental constraints and machine characteristics. The DHM used MSHA fatality information to validate the model parameters relating to operator position, which could pose a threat to operator safety. Some of these positions were in the MSHA red zone (MSHA, 2006). It should be noted that the results from this study make a case supporting or even expanding the red zone strategy. The human subjects recruited from local mines were asked to perform realistic movements in a laboratory setting (Fig. 1) that mimic escaping from the path of a moving machine. FIGURE 1 Motion capture for continuous miner speed study. � The motion data was obtained using various operator work postures and escape paths (directions) typical for tramming operations of continuous miners. Complete details of these tests, the development of the DHM and data analysis are contained in Bartels et al. (2008). The DHM, shown in Fig. 2, was developed using the motion capture data to provide the means to measure parameters that would be used to predict struck-by events when the operator tries to move out of the way of the moving machine. FIGURE 2 Digital human/continuous mining machine model. The digital human operator’s movements were constrained by using motion capture data of test subjects as discussed in (Bartels et al., 2007). To present a realistic operator response to the moving machine, researchers programmed the operator’s movement using a delayed start in accordance with reaction times reported by Drowatzky (1981). This delay ranged from 0.19 to 24 seconds. The DHM output parameters included: • The time when the machine first begins to move. • The time when the operator first begins to move. • The time when the operator is struck by an object. • Name of the object that struck the operator. • The operator’s distance from the start position when struck by an object. MSHA’s fatality reports (Dransite and Huntley, 2005) provided information to help validate the model regarding objects that struck the operator and the operator’s distance and location from the machine at the time of being struck.

Abstract: One of the major ignition sources of fires and explosions in underground U.S. coal mines is flame cutting and welding. On May 20, 2006, a flame-cutting operation at the Darby Mine No. 1 led to an explosion that resulted in five fatalities. On Jan. 22, 2003, another explosion caused by flame cutting at the McElroy Mine resulted in three fatalities and three serious injuries. On June 22, 1999, a flame-cutting and welding operation at the Loveridge Mine caused a fire that required sealing the mine for an extended period. On March 19, 1992, an explosion initiated by welding at the Blacksville No. 1 Mine resulted in four fatalities. These examples of mine incidents demonstrate the need to conduct research to develop best practices for safely conducting flame cutting and welding in underground coal mines. The National Institute for Occupational Safety and Health (NIOSH) conducted a study on fires and explosions in underground U.S. coal mines that were caused by flame-cutting and welding operations to determine the root causes of these types of incidents. The methodology included interviewing mining personnel who perform flame-cutting or welding operations in underground U.S. coal mines and visiting mines to observe these operations. In addition, MSHA reports of investigations and accidents statistics were analyzed. The findings were used to identify and compare differences between flame-cutting and welding practices and techniques in small and large mines, eastern and western mines, low-seam and high-seam mines, room-and-pillar and longwall mines, and between experienced miners and new miners. As a result of this study, best practices were developed to reduce the number of fires and explosions caused by flame cutting and welding. This paper provides a summary of this research. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of the National Institute for Occupational Safety and Health. Introduction Flame-cutting and welding activities in underground U.S. coal mines continue to be a major originator of fires and explosions. Seventeen reported fires or explosions in underground U.S. coal mines for the period 1990 1999 were initiated by flame-cutting and welding activities (DeRosa, 2004). In addition, from 2000 2007 six underground fires or explosions were reported that were attributed to flame-cutting and welding operations and caused injuries or fatalities (DeRosa, 2008). These fires occurred when hot sparks or hot molten metal came in contact with combustible materials, such as coal, oil, grease, clothing, rags, paper and aerosol cans. These explosions occurred when hot sparks or hot molten metal came in contact with flammable gases. The source of fuel for the majority of these incidents was methane gas. The Mine Safety and Health Administration’s policy in 30 CFR §75.1106 states that methane tests must be done continuously in all locations where methane is likely to exist (MSHA, 2007). In order to reduce the fire and explosion hazards caused by flame-cutting and welding operations in underground coal mines, a study to better understand the root causes of incidents and recommended best practices was completed by the National Institute for Occupational Safety and Health. The information from this study is presented here in the following format. In the first section of this paper, case histories are presented of major fires and explosions in underground U.S. coal mines that were caused by flamecutting or welding operations. The second section of the paper identifies and discusses the root causes of fire or explosions associated with flame cutting or welding operations. Results of flame-cutting experiments performed at NIOSH’s Lake Lynn Laboratory are presented in the third section of this paper. The paper concludes with recommended best practices and the conclusions of this research. Darby Mine explosion On May 20, 2006, an explosion occurred in a sealed section of the Darby Mine No. 1 in Kentucky, resulting in fatal injuries to five miners and injuries to one miner. A methane explosion occurred behind a block seal. MSHA determined that the explosion was caused by sparks that were generated by flame-cutting of a metal roof strap that passed above the seal (MSHA, 2006). Figure 1 shows the type of metal roof strap that was being flame-cut by an oxygen and acetylene torch at the time of the explosion.

Abstract: The geological conditions of Liguo Iron Ore Mine are explained in brief.The relation between mineralization mineral enrichment,strata,structures textures,magmatite and wall rock alteration is analyzed.

Abstract: The harm of dust and the standard of dust concentration in underground coal mine are explained.According to practical experiences,the methods such as water filling into coal bed,dedusting ventilation,wet type mining,mud water seal blasting,water spraying,putting on of breathing mask are discussed.Among these methods,water filling into coal bed and dedusting ventilation are the focus of discussion.

Abstract: According to deep hole blasting practice,the wide hole spacing small burden technology is studied because the technology can improve quality of blasting.To ensure blasting performance,the coefficient of blast-hole concentration should be determined by test;the holes should be in a triangular arrangement;and the hole rows for the first blasting should be controlled.

Abstract: The properties of China's tailings resources,their comprehensive utilization ways and status are analyzed.It is pointed out that the compressive utilization of tailing source is the apodictic choice for realizing sustainable development of mining industry.The deepgoing study and execution of comprehensive utilization of tailings resources will bring good economic social benefits.

Abstract: The goal of reasonable mining distribution is to guide and coordinate the mining production from the standpoint of whole mining process and overall benefit of the mine.The key point of mining distribution is the determining of a rational mining ratio from different stope,which is the guarantee for ore quality qualification,ore quality improvement and economic benefit increment.

Abstract: In China,the gold ore that is difficult to be treated takes near 1/4 of proven gold ore resource.The direct leaching of such ore demonstrates often low gold recovery rate.Three kinds of leaching method are introduced among which the method of pretreatment-cyaniding leaching is popular.

Abstract: This thesis make a study on the coal preparation plant dust screening pipeline,according to fluid dynamics and phase flow theory and the test data,set up the Simulation model of dust settlement movement discipline,and applicated FLUENT software set up simulation of different sizes dust settlement movement discipline of Coal Preparation plant's horizontal pipe section,Finally get the Dust settlement track in the three pipe sections,provided important theoretical basis to enhance dust collection efficiency.

Abstract: Development of the permanent drum and roll magnetic separation equipment,both at home and abroad,are briefly introduced in this articleAnd magnetic properties,mechanical structure,magnetic roller and magnetic drum design,process parameters and other aspects are discussed and analyzedAt the same time,while in the design and application problems,a few measures on the optimization of permanent magnetic device are put forward

Abstract: According to the properties of gentle-inclined ore body,FLAC3D software kit is used to simulate goaf's stability data of an iron ore mine before and after backfill.The results of simulation are helpful to treating of goaf,determining of pillar recovery scheme and realizing of safe operation.

Abstract: The geological background,geological properties and mineral distribution regularity of Karamansu Deposit are discussed in detail.The ore finding marks and ore formation regularity are summarized that provides reference for finding similar deposit.