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By Richard M. Winar

In September of 1988, the United States Environmental Protection Agency (US EPA) published final regulations that effect owners and operators of underground storage tanks (USTs). These regulation include stringent technical requirements for owners and operators of USTs containing petroleum and certain hazardous substances. The technical requirements effect both new (those installed after December 1988) and existing USTs and include notification; performance and operating standards; release detection, reporting, investigation; release response and corrective action; and closure requirements for out-of-service USTs. The USEPA also published final regulations regarding financial responsibility requirements for owners or operators of petroleum USTs in October of 1988. These regulations require owners or operators of petroleum USTs to demonstrate financial responsibility for the costs of corrective action and compensating third parties for bodily injury or property damage caused by releases from petroleum USTs. Like thousands of other industries, the mining industry must become thoroughly familiar with these new regulations in order to properly manage their UST systems, ensure compliance, and understand other ramifications imposed by these regulations.

Jan 1, 1990

By J. P. Rider, T. W. Beck, J. A. Organiscak

"The National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Mining Research Division (PMRD) has recently developed a series of models utilizing computational fluid dynamics (CFD) to study gas distribution around a continuous mining machine (CM) with various fan-powered flooded bed scrubber discharge configurations in an exhaust curtain working face. CFD models utilizing species transport model without reactions in FLUENT were constructed to evaluate the redirection of scrubber discharge toward the mining face rather than behind the return curtain. The study illustrates the gas distribution in the slab (second) cut. The following scenarios are considered in this study: 100% of the discharge redirected back towards the face on the off-curtain side of the continuous miner; 100% of the discharge redirected back towards the face, but divided equally to both sides of the machine; and 15% of the discharge redirected toward the face on the off-curtain side of the machine, with 85% directed into the return. These models were compared to a model with a conventional scrubber discharge where air is directed away from the face into the return. The models were validated based on experimental data and accurately predicted sulfur hexafluoride (SF6) gas levels at four gas monitoring locations. One additional prediction model was simulated to consider a different scrubber discharge angle for the 100% redirected, equally divided case. This paper describes the validation of the models based on experimental data and the gas distribution results of the CFD models. INTRODUCTION To control methane gas and dust concentrations on continuous mining operations, it is important to provide sufficient quantity of fresh air from the end of tubing or curtain to the active face. In this region, operating machine-mounted water sprays and fan-powered, flooded bed scrubbers can significantly increase the amount of airflow reaching the face [1-3]. However, when an entry is large, especially in higher coal seams, the water spray configuration and the scrubber capacity may not be adequate to maintain sufficient forward airflow to confine the dust cloud to the face. Dust can roll back toward the machine operator as a result [4]. Currently, several coal mines having negligible gas emissions in higher coal seams (2.4 to 3.0 m [8 to 10ft.] thick) with exhaust tubing systems redirect the scrubber discharge back towards the face in an effort to improve dust collection. NIOSH studied the redirected scrubber exhaust techniques for dust and gas control in a full scale CM gallery to assess its potential for industry-wide application [5]. For a CM face, previous studies have concluded that the behavior of gas and respirable dust is a complex process. The distribution/dispersion and transport of gas and respirable dust after they are liberated at the face are governed mainly by their mass properties, the movement pattern of the ventilation air, and interactions with the mining machine and water droplets created by the external water spray systems used for dust control [6-10]. To understand these behaviors in a complex CM environment and to evaluate the effectiveness of various control techniques, numerical modeling has become a necessary supplement to laboratory experiments and field studies."

Jan 1, 2017

By Taghi Sherizadeh, Kutay E. Karadeniz, Dogukan Guner, Samuel Nowak

Despite numerous empirical, numerical, and hybrid studies conducted on the subject of coal rib failure, mine fatalities related to rib failures continue to occur. To date, a standard methodology for the design of coal rib supports which can accommodate the wide range of unique conditions found in coal mines in the United States has not been developed, forcing mine operators to rely on the trial-and-error process or industrial legacy practices for the design of coal rib support patterns. In this study, to better understand the support systems' effectiveness, various rib bolting scenarios were investigated. A coal mass constitutive model is implemented to a Distinct Element Method (DEM) code, and rib factor-of-safety (RibFOS) analyses were performed by applying strength-reduction for the coal matrix and explicitly modeled face cleats. Ribs with various overburden depths and mining heights composed of solid Banded-Bright coal (BBC) were considered in this study. Different scenarios such as rib bolt length and the number of bolts were studied to quantify the effect of primary rib support density on rib stability. As a result of the unsupported coal rib analyses, a decaying exponential relationship is proposed between the mining height-overburden depth and the RibFOS. The developed coal rib models clearly show the coal-rib bolt interaction and illustrate how the addition of rib bolts restricts the movement of the coal rib. The results reveal that the strength reduction method in DEM code with the coal-mass model can aid in the development of practical primary support charts, which can help in the rib support design process. In addition, an unsupported RibFOS cutoff

By S. S. Peng

Since the introduction of shield supports in 1975, safety and production in U.S. longwall faces have improved considerably. Many faces had a daily production of up to 8,000-10,000 tons and productivity reached 600 tons per man-shift. But a longwall face requires expensive equipment which slows down the growth rate of the percentage of longwall production. Among the equipment cost the shield support accounts for 50-60%, a very large proportion. Therefore, it will have a large impact on the growth of longwall mining in the U.S. to reduce the support cost (1).

Jan 1, 1988

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices' specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and 'struck by' falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved 'strikes' by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By E. L. Gillenwater

Coal mine planning has three distinct decision levels: (1) decisions made at the upper level of management that determine production requirements for specific mines, (2) decisions made at the middle management level, the mine level, that determine how the mine satisfies its production requirements, and (3) decisions made at the operating level of management, the sectional level, that determine the sequencing of machine movements at the face. In this paper, the underground production scheduling problem is formulated in multi-stages using the Hierarchical Production Planning (HPP) approach. Two linear programming HPP models for accomplishing decisions at the first two levels are developed. Both models address the aggregate production planning process. The Mine Production Scheduling (MPS) model permits determination of monthly production and inventory levels. The Section Production Scheduling (SPS) model incorporates a quality blending requirement into the determination of weekly sectional production and inventory levels.

Jan 1, 1995

By J. E. Sepulveda

Tar sand deposits in the. Stare of Utah contain more than 25 billion barrels of in-place bitumen, Although thirty times smaller than the well-known Athabasca tar sands, Utah tar sands do represent a significant domestic energy resource comparable to the national crude oil reserves (31.3 billion barrels), The term "tar sand" refers to a consolidated mixture of bitumen and sand, Based upon a detailed analysis of the physical and chemical properties of both the bitumen and the sand, a hot-water extraction process for Utah tar sands is currently being developed in our laboratories at the University of Utah. This process involves intense agitation of the tar sand in a hot caustic solution and subsequent separation of the bitumen by a modified froth flotation technique., Experimental results indicated that, for a given tar sand sample, percent solids and caustic concentration were the two most important variables controlling the performance of the digestion stage, These variables were identified by means of an experimental, factorial design, in which coefficients of separation greater than 0.90 were realized. Although preliminary in nature, the experimental evidence gathered in this investigation seems to indicate that a hot-water extraction process for Utah tar sands would allow for the efficient utilization of this important energy resource.

Jan 1, 1978

By H. Walqui

The Empire concentrator currently processes magnetite ore from the Empire pit which is approaching the end of its mine life. Over the past two years plant wide tests have been run to access the plant capability of processing ore from the neighboring sister Magnetite pit. The neighboring pit presents challenges to the Empire plant due to its higher feed grade, finer liberation size and higher percent silica which require additional secondary grinding and concentrating capacity to achieve concentrate grade. The following paper summarizes the results from the second full scale test, run during 2009 under a challenging market environment

Jan 1, 2010

By I. S. Parrish

As noted by McKinstry (1948), "How to deal with erratic high samples is one of the knottiest problems in ore estimation." He described a variety of approaches and methods to alleviate the problem through the averaging and weighing of assay data and by "cutting" or "clipping" of the erratic assays. Bird (1991) elaborated on ways of dealing with the nugget effect (erratics). He listed several procedures for cutting. Bird's list has been reproduced and/or added to by several authors including Vallee (1992) and Stone and Dunn (1993). Missing from each of these otherwise excellent reviews is a definition, description or guideline to recognition of an "erratic," or just where or when cutting is warranted or justified.

Jan 1, 1997

By L. L. Probert

A three-year project focusing on the depth profiling of respirable coal mine dust particles using scanning electron microscopy coupled with energy dispersive x-ray analysis resulted in the development of statistically classified particle-behavior models. The models depict the changing silicon fraction for 5-, 7-, 9-, and 11-keV accelerating voltages, thereby giving a clear indication of the nature of various types of particles. This paper will focus on the results for respirable quartz particles that occur in underground coal mines across the U.S. The relationship between the amount of surface-available quartz and coal seam rank is well established for Pennsylvania

Jan 1, 1994

By L. A. Rapoport

This paper outlines the methodology and results pertaining to the simulation of the U.S. coal industry in the context of an integrated, time-dynamic model of Long-Range Energy Developments and Supplies (LORENDAS). By way of background we will present first an overview of the overall LORENDAS model structure. We will, then, outline the characteristics of the formulation and data used for the representation of the U.S. coal industry. Finally, a series of model results will be reviewed, with particular emphasis on the long term projections of coal production and distribution under different basic economic and/or political scenarios.

Jan 1, 1977

By Neely H. Bostick

The primary and secondary organic matter in sedimentary rocks changes markedly when it becomes heated as a consequence of thick sediment accumulation and normal geothermal gradients, high geothermal gradients, hydrothermal fluids, or igneous events. Contact zones adjacent to small igneous dikes allow sampling of a single sedimentary bed that has been exposed to temperatures in a range as great as 30-500°C. Dike-heated rocks are, therefore, good sites for research on thermal effects because variation of the original organic matter is minimal, the range of heating is great, and the history of rock heating in a dike zone is relatively well established. In rare cases, such as at the dike near Wolcott, Colorado, described in this paper, the rock samples are sufficiently unweathered to allow analysis of secondary bitumens generated by the heating. We show by optical microscopy, scanning and transmission electron microscopy, and geochemistry the kinds of solid organic constituents in sedimen¬tary rocks and in ore deposits and the effects of heating. The primary phytoclasts that make up kerogen In sedimentary rock and in some ore deposits (pollen, resin, charcoal, humified woody remains, etc.) change in ways that are familiar to coal specialists. Progressive loss of moisture and other volatile matter yields residual kerogen that Is relatively higher in carbon and lower In oxygen and hydrogen, and has greater molecular or polymer ordering. In our contact zone examples, atomic H/C ratios of kerogen decrease from 0.8 (one case from 1.4) to 0.4, the hydrogen index from Rock-Eval pyrolysis decreases from 70 to 5, and the transformation ratio from Rock-Eval pyrolysis increases from 0.06 to 0.5 with increased heating. The visible effects of heating seen by scanning electron microscopy and light microscopy include formation of vesicles and flow structure, the development of optical anisotropy and mesophase, change of color, and increase in the refractive index, absorption, and reflectance. Some of these changes In primary organic matter are observed also in secondary organic matter, for example, in solid bitumens in sedimentary gold ores or In lead-zinc deposits. Quantification of color or absorption changes is difficult because it Is hard to make thin sections of organic matter with uniform thickness. However, natural objects with uniform thickness can be used. We show the colors of pollen In rocks from the Salton geothermal field at temperatures ranging from 100 to 300°C, and of pollen heated in laboratory bombs from 150 to 400°C. Measurement of vitrinite reflectance is a practical way to quantify thermal changes of kerogen and minerals of most sedimentary rocks because vitrinite is fairly widespread In sedimentary rocks of Devonian age and younger. For example, the reflectance of vitrinite rises from the regional level, as low as 0.4 % Ro, to more than 5% Ro adjacent to dikes that cut sedimentary rock. This is equivalent to the rank change from subbituminous coal to extreme meta-anthracite. The rise is most abrupt near thin dikes, but when the distance from the igneous contact is expressed as a ratio of dike thickness the reflectance rise is similar for all dikes. Controlled heating of vitrinite-bearing rocks In the laboratory provides a means to assign known temperatures to vitrinite-reflectance values. The derived plots of temperature In dike contact zones have the same shape and level as those predicted by heat theory, and the temperature lines extend toward dike contact temperatures that are about half those of most magmas, as expected if the intrusions did not flow past the sedimentary rock for a long time. The globular submicroscopic structure of vitrinite and Its heterogeneity, which we show by means of transmission electron microscopy, pose some problems In selecting a uniform type of vitrinite for measurement of thermal changes. At present, routine reflectance values of vitrinite dispersed in rocks have a relative uncertainty as great as 10%. Thermally caused chemical changes in the primary kerogen, especially decrease in the oxygen and hydrogen contents, lead to formation of secondary products in the contact zone. The measured changes of secondary organic products in these rocks include formation of peaks In concentration of total extractable organic matter and in hydrocarbons, and in the saturated/aromatic hydrocarbon ratio over the interval where primary organic hydrogen decreases most rapidly--then destruction or expulsion of these products at high temperatures. The odd carbon-number predominance of the C25-C31 n-alkanes In samples far from the dike disappears by about 1.3% Ro whereas the 13C: 12C ratios of the saturated and aromatic hydrocarbons increase over most of the thermal range. The saturated aromatic hydrocarbon ratio at first increases, corresponding to thermal generation of saturated hydrocarbons, then decreases in the most altered samples. One use of organic petrography, in addition to identifying organic fades or thermal alteration, is recognition of sample contamination such as the solid organic additives used in some drilling muds which may contaminate geochemical samples.

Jan 1, 2013

By C. A. Rowland

When studying the feasibility of using autogenous or semi-autogenous grinding, operating and engineering companies frequently request assistance from the mill manufacturers in three areas - 1. Recommendations for studying the suitability of autogenous or semi-autogenous grinding for the specific ore. 2. Mill sizing, to accomplish required grind. 3. Recommendations on mill design and operation. Rod mills and ball mills are selected by the use of bench scale standard grindability tests and established calculation procedures, of which the most frequently used are the Bond Grindability tests and the Bond Work Index equation(1). The accuracy of these have been reviewed by C. A. Rowland in his papers, showing the correlation of work indices calculated from plant operating data and laboratory grindability tests(2)(3). The selection of autogenous and semi-autogenous primary mills has not reached the stage where mill size selection can be made from bench scale tests. Considerably more extensive testing will be necessary than required for selecting rod mills and ball mills.

Jan 1, 1974

By David A. Zegeer

Since 1980, more than 40 miners have lost their lives by being trapped and suffocating in stock-piles, bins, silos, and similar facilities. Mine Safety and Health Administration (MSHA) is concerned about these accidents, most of which could be prevented if management and miners appreciated the hazard. In February 1986, the coal industry had a suffocation accident that drew national attention. A group of engineers were standing on a coal stockpile. It gave way under their feet. Five of the engineers were buried in the coal and suffocated. As with all fatal accidents, MSHA investigated the cause of this tragedy. I also asked my staff to look into the history of similar accidents. I found out it is rare to have multiple fatalities due to suffocation in material. But, over the years, there has been a long series of individual fatalities from this type of accident. Here are some examples. In Oct. 1985, a coal mine employee in Virginia entered a hopper to collect a coal sample. The feeders were started while he was standing in the hopper. He sank into the coal and suffocated. In Sept. 1985, an employee at a Texas sand company climbed onto a surge pile, to check a problem with sand flow. When other employees found him, the sand had swallowed him. Only his face and one hand were showing. Though the victim's face was clear of the sand, he suffocated before rescuers could free him. Many people do not realize you can suffocate in this kind of situation, though your head is free. Every time you exhale, the material has room to shift and press harder against your chest. Then you have less room to inhale. If this goes on long enough, it gets to the point where you cannot inhale at all. That is what happened to the victim in this accident. In May 1985, a plant operator at a Wisconsin sand and gravel operation purposely slid into the drawhole of a surge pile. He tried to break up large rocks that were clogging the drawpoint. Material piled on the side of the surge pile slid down and buried him. The operator was wearing a safety belt and line. But he was covered by 3 to 3.7 m (10 to 12 ft) of material and could not be pulled free. This incident illustrates another point. A safety line can help. But it is not enough. Workers should never put themselves in a position where loose material can fall on them from above. In Jan. 1985, a young Utah coal miner was waiting for a ride home from a co-worker whose job was at the tipple. A jam developed in the coal bin. The men could not clear the jam from the outside. So the young miner jumped into the bin. He started digging with a shovel at the toe of the coal pile. It caved in on him and he suffocated. The young miner had only three months of mining experience. He had not been trained for work around the coal bin. Since 1980, 16 coal miners and 26 nonmetal miners have lost their lives in loose-material suffocation accidents. Suffocation accidents occur in many areas around mines - bins, hoppers, stockpiles, surge piles, and silos, among others. One recent accident occurred at a surface coal mine in Illinois. The victim was walking over an area that had been drilled and blasted the week before. He fell into a void that had been crusted over with earth. He too suffocated. And many materials have been involved in suffocation accidentscoal, crushed stone, sand, salt, different kinds of ore, ammonium nitrate pellets used in blasting, and a powdery byproduct of iron ore milling called calcine. In addition to fatal accidents, there are near misses. We do not know how many. Most of them are not reported. A person who gets trapped in material, but is rescued, is usually unhurt. MSHA sometimes hears about these accidents by chance, or because they are reported in the local media. We do know that near misses are happening far more often

Jan 1, 1987

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By F. J. Susko

The Bureau of Mines applied a flotation model to the hydrodynamics of the Bureau-developed rapid flotation system. Rapid flotation is a process which separates the bubble-particle attachment and froth-pulp separation stages of the flotation sequence into two discrete unit operations allowing the operations to be individually optimized. Both an air-injected hydrocyclone and an in-line mixer were used as the bubble-particle attachment unit and a shallow-depth froth separator was used to separate the ore-laden froth from the pulp. A 10-L/min batch system was tested on a coarse coal feed sample. The system was tested using two different bubble-particle attachment units at various aeration levels in order to test the flotation model. The results of the flotation model reflect general trends observed in the experimental data.

Jan 1, 1993

By S. Veeramasuneni

The interaction of soluble salt particles in their saturated brines is of interest from both a fundamental and practical perspective. The state of dispersion/aggregation of alkali halide particles as determined by optical microscopy can now be described based on non-equilibrium electrokinetic measurements for alkali halides. These results indicate that the sign of the surface charge has a significant effect on the extent of particle dispersion/aggregation. Atomic force microscopy (AFM) is now being used to directly measure the interaction forces responsible for the stability of these alkali halide aggregates observed in optical microscopy studies. These experimental results can be explained based on the interfacial polar interaction theory. In this regard the nature of such particle interaction forces at high ionic strengths are examined to improve our understanding of these systems.

Jan 1, 1997

By A. D. Hammond

Vanadium is a transition metal, ductile, silver-gray metal. It is harder than most metals and steels and it has a good resistance to corrosion, it is stable against alkalis, sulfuric and hydrochloric acids. Vanadium is oxidized in air about 660 °C, although an oxide layer forms even at room temperature. The chemistry of vanadium is noteworthy for the accessibility of the four adjacent oxidation states 2-5. In aqueous solution the colors are lilac V2+, green V3+, blue VO2+ and, at high pH, yellow VO42-. Considered to be as abundant as zinc, metallic vanadium is not found in nature, but it is known to exist in about 65 different minerals. Economically significant examples include patronite (VS4), vanadinite (Pb5(VO4)3Cl), associated to the uranium mineral carnotite (K2(UO2)2(VO4)2?3H2O) and in a vanadium mica called, roscoelite K(V3+,Al,Mg)[(Si,Al)4O10](OH)2. Other important vanadium ore oxides are montroseite (V,Fe)O(OH) and corvusite (Na,Ca,K)V8O2.4H2O. Much of the world's vanadium mine production comes from magnetite-hosted deposit and to a lesser extent, in sedimentary-hosted sandstone. Vanadium is also present in bauxite and in fossil fuel deposits such as crude oil, coal, oil shale and tar sands. In crude oil, concentrations up to 1,200 ppm have been reported. A significant vanadium production comes from the processing of petroleum residues, fly ash and recycling spent catalysts. In 2011, metallurgical applications accounted for 93% of the use of vanadium. Here, an alloy called ferrovanadium is utilized to strengthen steel to be used as rebar in the construction industry, in the manufacturing of axles, crankshafts, gears and chassis in the automotive and transportation industries and in the manufacturing of high-speed tools. In the aerospace industry, vanadium is utilized as a titanium alloy in the manufacturing of airframes and jet engines. The reason for the wide use of vanadium as alloy is its high strength-to-weight ratio, the highest of any engineered material on earth. The major non-metallurgical use of vanadium is in the production of catalysts for the chemical industry. The most commercially important compound is vanadium pentoxide (V2O5), which is used as catalyst in the production of sulfuric acid.

Feb 23, 2014