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By Thomas Novak, Sampurna Arya

Controlling dust generation and keeping it below permissible limits to meet federal dust standards at the working face of a roomand-pillar coal mine is a challenge for a mine operator. With the recent changes in federal dust regulations requiring lower worker exposure, maintaining compliance has become increasingly difficult. The current most effective practice of dust control at a continuous miner face in an underground mine is the use of a flooded-bed scrubber. A study carried out by the National Institute for Occupational Safety and Health (NIOSH) indicated that a flooded-bed scrubber could achieve cleaning efficiencies between 58 and 90%. But, the operation of such a system is maintenance intensive. The flooded-bed scrubber screen becomes clogged with dust particles and requires frequent cleaning to maintain performance. However, the dust control issue is not solely a mining industry problem. Other industries face similar issues. The University of Kentucky collaborated with Toyota Motor Manufacturing on the development of a novel wet scrubber, called the vortecone scrubber, for capturing oversprayed paint particles in automotive paint booths. The vortecone scrubber achieved a cleaning efficiency of 99% and required minimal maintenance. This study aims to assess the ability of this vortecone scrubber to capture respirable dust in an underground coal mine. The paper presents the results of computational fluid dynamics (CFD) modeling of this vortecone scrubber. It discusses the effects of air quantity and dust particle size on the performance of the scrubber.

By R. Hogg

A standard procedure for determining size-specific gravity distributions of fine coal (including the -400 mesh fractions) by sieving and centrifugal sink-float separation is described. The significance of using a centrifugal sink-float technique and the effect of the magnitude of the centrifugal force on the separation of fine coal are confirmed in the present study. The relative merits of cumulative versus incremental sink-float are discussed. The procedure is evaluated quantitatively using data obtained for the products of froth flotation of -100 mesh coal.

Jan 1, 1994

By Robert A. Blake

Introduction Sulfide copper, zinc and iron concentrates are recovered in the 5000 ton per day flotation plant operated by Cities Service Company at Copperhill, Tennessee. The heavy-sulphide ore is first ground to 50 pct. minus 200 mesh and a bulk concentrate floated. The bulk concentrate, over 60 pct. of the ore weight, is reground to 72 pct. minus 200 mesh and then treated in a selective flotation circuit to recover copper and zinc concentrates. The tailing of selective float is the iron sulphide material source for the manufacture of sulfuric acid and iron sinter. The valuable minerals in the ore are chalcopyrite, marmatitic sphalerite, pyrite, and pyrrhotite. A typical mineralogical analysis will be found in table 1. A typical material flowsheet will be found in figure 1. Shown in the flowsheet are a scavenger float on the bulk rougher tailing and a copper-talc separation unit. At times a lead-copper separation circuit has been run.

Jan 1, 1971

By Gamal Rashed, Zorica Radakovic-Guzina, Khaled Mohamed

"Underground coal mines use mechanical bolts in addition to other types of bolts to control the rib deformation and to stabilize the yielded coal ribs. Limited research has been conducted to understand the performance of the mechanical bolts in coal ribs. Researchers from the National Institute for Occupational Safety and Health (NIOSH) conducted this work to understand the loading characteristics of mechanical bolts (stiffness and capacity) installed in coal ribs at five underground coal mines. Standard pull-out tests were performed in this study to define the loading characteristics of mechanical rib bolts. Different installation torques were applied to the tested bolts based on the strength of the coal seam. A typical tri-linear load-deformation response for mechanical bolts was obtained from these tests. It was found that the anchorage capacity depends mainly on the coal strength. Guidelines for modeling mechanical bolts have been developed using the tri-linear load-deformation response. The outcome of this research provides essential data for rib support design.INTRODUCTIONThe design of rib bolting in U.S. coal mines is based on a trial-and-error process, which is insufficient to eliminate injuries and fatalities caused by rib falls. In the last 10 years, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States (MSHA, 2018). Recently, NIOSH researchers surveyed 90 sites in 14 underground coal mines in the Eastern United States and found that 11 mines are bolting coal ribs, with 4 of those mines using the conventional mechanical bolts and the others using fully-grouted bolts. NIOSH researchers are developing an engineering design procedure for rib support. Pull-out test programs for mechanical and grouted rib bolts were conducted to understand the rib control mechanisms of rib bolts. Only the test results of mechanical bolts are presented in this paper.Mechanical bolt (or point-anchored bolt) is a tensioned bolt anchored down the borehole with a mechanical expansion shell or a resin-grouted medium (Peng and Tang, 1984). The mechanical anchor bolt consists of a plug, a shell anchor, a headed bar with the top end threaded, a bearing plate, a hardened washer, and occasionally a friction-reducing washer. The shell anchor can be either standard or bail type (Figure 1). The standard type shell is held in place by a nut, while the bail anchor uses a strap to hold the shell in place (Peng, 2008). The diameter of the bar is generally 16 mm (? in) for the #5 bolt or 19 mm (6/8 in) for the #6 bolt. The diameter of the hole that the bolt is inserted into must be carefully controlled, since an oversized hole can result in poor anchorage (Moss, 1971). The anchorage is obtained by applying torque to the bolt head, which in turn pulls the plug down in the shell and expands the serrated leaves against the sides of the hole-wall. In the resin-grouted anchor, the resin replaces the mechanical expansion shell at the end of the hole. The anchorage is achieved by bonding between the resin, bolt, and the hole-wall."

Jan 1, 2019

By Ewald H. Kurt

The story of America's progressing standard of living is the story of machines. The automobile, the airplane and the telephone for example have made our transportation and communication systems what they are today. Even they depend in turn on another machine -- the rock drill. By making it possible to economically mine and quarry minerals in large quantities, this versatile machine supports virtually every American industry relying on the earth for its buried raw materials. And, like all machines, the rock drill's mechanical evolution was stimulated by production demands placed upon it. All percussion drills are mechanized versions of the original drilling unit -- man, Figure - illustrates a double jack team at work before the turn of the century. The man with the sledge hammer is the impact device while the man sitting is the rotation mechanism and feed motor. The first drifter drills (Figure 2) performed the same function but with increased mechanical efficiency. A piston controlled by a valve produced impact. A built-in ratcheting mechanism provided rotation and a hand-cranked screw provided feed thrust. As the industrial revolution proceeded into the twentieth century, the demand for raw materials stimulated refinements of all recovery machinery, including rock drills. As drills became heavier and more powerful, a mounting (or carrier) had to be provided to move the drill from hole to hole. This resulted in the development of the wagon drill (Figure 3). Propelled by muscle power, the wagon drill was for years the standard tool of the surface drilling industry (Figure 4). As drills penetrated faster and as labor increased in cost, the time consumed in nondrilling peripheral operations had to be

Jan 1, 1977

By Weijie Wei, Jinwang Zhang, Shengli Yang

Due to the unbalanced top coal drawing process, the recovery ratio is low, and the equipment stability is poor in the longwall top coal caving mining (LTCC) with an inclined seam. Aimed at these problems, this paper proposes the dynamic group caving method (DGCM). The basic procedure is that when the drawn volume of top coal from a support reaches about half, the next support begins drawing top coal, and the drawing process between the two supports is half done. Taking the panel no. 8103 in Beixinyao mine as the research object, the characteristics of drawing body shape, particle movement trace, and variation of recovery ratio are investigated by using Particle Flow Code. Furthermore, the effect of mining-caving ratio on the top coal drawing law under DGCM is revealed by the simulation results, and compared with the single-opening sequence caving method (SSCM) and the doubleopenings sequence caving method, advantages of DGCM are determined in inclined seams. The results show that the drawing body can be divided into the left and the right under DGCM. Specifically, the drawn top coal particles in the left are from the upper part of this support, that in the right are from the lower part of next support, which verifies the correctness of the theoretical analysis. The characteristics of the drawing body shape are more significant when the mining-caving ratio is relatively small. In addition, with increasing mining-caving ratio, the recovery ratio of the panel both increases gradually underDGCMand SSCM.When the miningcaving ratio is less than 1:1, the recovery ratio of DGCM is obviously higher than that of SSCM, indicating that DGCM can greatly improve the recovery ratio. While when the mining-caving ratio is larger than 1:1, the recovery ratio of two caving methods is basically equal; however, the advancing time is significantly shortened under DGCM. This study are of great significance to realize high efficiency and precise caving in LTCC with an inclined seam.

Jul 7, 2020

By Juan J. Monsalve, AMAN SONI, ALEX PFREUNDSCHUH, Nino Ripepi

Ground control failures are one of the main causes of accidents in the underground stone mining industry. Some of the fundamental tools for rockfall hazard identification are related to rock mass characterization and geotechnical discontinuity mapping. Recent technological advances in these methods are related to remote sensing techniques and point cloud processing software for automated discontinuity mapping. Remote sensing techniques, such as LiDAR and photogrammetry, generate multimillion point clouds with millimetric precision, capturing the structure of the rock mass. The automated point cloud processing tools offer alternative algorithm-based methods to characterize and map these discontinuities. However, their applicability is constrained by multiple factors such as site specific conditions of the rock mass and the parameters used within the mapping algorithms. This paper evaluates the performance of automated discontinuity extraction software compared with manual virtual discontinuity mapping. Sampling windows from laser-scanned sections in an underground limestone mine are defined and mapped using discontinuity set extractor (DSE). Results from the virtual discontinuity software are compared with manually extracted fractures from I-Site based on reviewing orientation, trace length, spacing, number of extracted discontinuities, and processing time. The analysis determined that the automated mapping algorithm was able to identify the same discontinuity sets that had been manually mapped. The automated mapping software mapped an excessive amount of smaller fractures, which caused the comparison of both mapping techniques to be unsuccessful in terms of trace length and spacing.

Mar 19, 2021

By J. G. Groppo

In 1992, 48 million tons of fly ash (not including bottom ash, boiler slag and FGD material) was produced in the United States, according to the American Coal Ash Association. Of this amount, 27.2% was utilized; although the percentage of fly ash utilized increased from 20% in 1980, there is significant potential to increase utilization. One of the primary driving forces behind increased utilization is increased disposal costs due to more stringent environmental controls and regulation. Fly ash is classified as a natural pozzolan and finds significant use as a cement additive. It has been found to increase the plasticity of the cement mixture while lowering the amount of mix water needed. In addition to improving leveling and finishing properties, fly ash improves pumpability as well as decreases the heat of hydration of the concrete mix which is particularly important in massive concrete structures such as dams where the concrete needs to cool as quickly as possible. The Loss on Ignition (LOI) of fly ash is an important quality when considering utilization. The LOI is primarily due to the presence of unburned carbon and the standard specifications for the use of fly ash as a mineral admixture in Portland cement concrete are given in ASTM C618-89a. which sets an LOI limit of 6%. However, many states have set a more stringent limit of 3%. The primary reason for the lower LOI specification is that carbon adsorbs air entraining agents which limits freeze-thaw resistance. This is particularly important in climates such as the midwestern US where multiple cycles of freezing and thawing are typical in winter.

Jan 1, 1995

By Michael Moats, Weston Hartzell

The stable domestic supply of gallium, indium, and germanium has been identified as a critical need for the USA. Electric arc furnace dusts processed to recover zinc produce a process residue that could be a potential source of these critical materials. An industrial zinc leach residue was characterized to determine the predominant mineral phases and leached with common mineral acids. Atmospheric agitated leaching using sulfuric acid and hydrochloric acid were examined to determine the extractions of gallium, indium, and germanium as a function of concentration, temperature, and time. Extractions of gallium were found to be similar to iron in all experiments. The maximum percentage of Ga, Ge, and In leached into solution were 82%, ~100%, and 89%, respectively.

Aug 8, 2023

By K. Fuenkajorn, S. Serata

Discussion by S. Serata and K. Fuenkajorn Background Results of the above study in the August 1991 issue of Mining Engineering offer valuable lessons in the solution of cutter-roof problems. The original study plan was initiated by the discussion authors to solve the problems using the "stress control method" of mining (Serata 1976, 1982; Serata, Carr and Martin, 1984; Serata and Gardner, 1986; Serata, Gardner and Preston, 1986; Serata, Gardnerand Shrinivasan, 1986; Serata and Kikuchi, 1986; Serata, Preston and Galagoda, 1987) However, the plan and the planner were changed to the arrangement reported in the paper. The change was considered reasonable at the time due to the mine engineers' uncertainties about the stress control method. Consequently, the basic principle of the study was shifted from the original stress control method to the method described in the paper, which will be called the "yield pillar method" for the purposes of this discussion. The paper convinces the reader that the yield pillar method fails to solve the cutter-roof problems. This doesn't mean that the stress control method also fails. Actually the contrary is true, as discussed below. Limitation of the yield pillar method The paper illustrates clearly how poorly the yield pillar method performs in solving the problem. The reason for this failure is the lack of the protective stress envelope needed to stabilize the cutter roof. Unfortunately, the protective envelope cannot be formed properly without utilizing the stress control method of mining. Changing the pillar size does not make much difference in the roof stability. Stress measurement The key issue is how to form the global stress envelope to make the gate entries safe for production. Therefore, measuring the stress condition of the ground around the mine opening is critically important to solving the cutter-roof problem, regardless of the method applied. With regard to the stress measurement, there is a serious question as to the reported stress state of [6 i = -51.7 MPa (-7499 psi), G2 = -44.5 MPa (-6458 psi) and 63 = -30.8 MPa (-4465 psi)]. It is mechanically impossible to have such a stress state at any location in the mine ground since the known initial vertical stress [o,,] is less than or equal to 800 psi. There may be a large stress state in the [61] direction, but that is possible only at the expense of the [63] value. Having the above stress tensors in the mine is simply impossible. The questionable, reported stress values could be attributed to the application of the overcoring method, which tends to produce erroneously large stress values in the extremely nonelastic mine ground. Stress control method The paper should be considered as a major contribution demonstrating the limitation of the yield pillar method. At the same time, the paper does not disprove the stress control method. However, in comparing the paper with stress control studies conducted in other similar failing grounds, the stress control method appears to be able to solve the problem more effectively. Therefore it is advisable that the mine not give up its efforts to solve the problem. [•] References Serata, S., 1976, "Stress control technique - An alternative to roof bolting?," Mining Engineering, May. Serata, S., 1982, "Stress control methods: Quantitative approach to stabilizing mine openings in weak ground," Proceedings, 1st International Conference on Stability in Underground Mining, Vancouver, BC, Aug. 16-18. Serata, S., Carr, F., and Martin, E., 1984, "Stress control method applied to stabilization of underground coal mine openings," Proceedings, 25th US Symposium on Rock Mechanics, Northwestern University, June, pp. 583-590. Serata, S., and Gardner, B.H., 1986, "Benefits of the stress control method," invited paper, American Mining Congress Coal Convention, Pittsburgh, PA, May 7. Serata, S., Gardner, B.H., and Shrinivasan, K., 1986, "Integrated instrumentation method of stress state, material property and deformation measurement for stress control method of mining," invited paper, 5th Conference on Ground Control in Mining, West Virginia University, Morgantown, WV, June 11-13. Serata, S., and Kikuchi, S., 1986, "A diametral deformation method for in situ stress and rock property measurement," International Journal of Mining and Geological Engineering, Vol. 4, pp. 15-38. Serata, S., Preston, M., and Galagoda, H.M., 1987, "Integration of finite element analysis and field instrumentation for application of the stress control method in underground coal mining," Proceedings, 28th US Symposium on Rock Mechanics, Tucson, AZ, pp. 265-272.

Jan 1, 1993

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 John S. Livermore

I feel very privileged to be added to the distinguished group that has received this award since the first recipient, Reno Sales, received it in 1954. If my research is correct, the last geologist to be honored was Bill Callahan in 1982. Perhaps the Awards Committee decided it was about time to select a geologist, and I became the beneficiary. Regarding the discovery of the Carlin ore body, it was the culmination of several studies - one going back as far as 1939 when a US Bureau of Mines engineer named W.O. Vanderberg recognized an unusual type of gold mineralization in sediments in northern Nevada. What was unusual was the fact that the gold was so fine that it would not concentrate in a pan. This, plus the study of the Roberts Mountain thrust by Ralph Roberts (1960) and conversations with Harry Bishop, manager of the Gold Acres Mine, pointed Alan Coope and me in the direction of the Lynn Window and, eventually, the Carlin discovery. There has been a great deal written about this discovery. I thought it would be interesting to discuss the evolution of Carlin gold exploration in Nevada since this time and to discuss the changes in perception of this unique type of mineralization. The method used by Coope and me in following up on Roberts' work was to map geologically the Roberts Mountain thrust and conduct geochemical sampling of favorable outcrops and float. Coope was highly experienced in geochemical exploration, as he had graduated from the Royal School of Mines in London, studying under John S. Webb, one of the pioneers of geochemistry as applied to mineral exploration. We first considered using pathfinder elements, but finally concentrated on gold analyses alone, even though the sensitivity of fire assays used at that time was only about one-half part per million. According to the spatial relationship of the mineralization to the Roberts Mountain thrust, as pointed out by Roberts (1960), we were looking for mineralization within the thrust itself. Our models were the known Gold Acres and Getchell mines, which contained reserves on the order of 0.9 to 1.8 mt (1 to 2 million st). These were not of substantial size, but I had hoped we might find more than one of these tributary to a mill that would justify an operation for our employer, Newmont Mining Corp. What we ended up with was the type of deposit we were looking for - an open pit deposit of micron gold, which the old timers had missed because the gold was not capable of being panned. But never in our wildest dreams did we think there was a chance of finding a body containing 10 mt (11 million st) of ore at 0.30 oz/st (10.3 g/t).The deposit was not in the thrust proper but was found in a favorable horizon of Roberts Mountain siltstone that was below the thrust. This is the first example of how a model should not be followed blindly. Newmont was successful in keeping this discovery quiet for some time. After a considerable amount of drilling, a rumor circulated that they had discovered a large low-grade ore body of around 1.7 to 2.1 g/t (0.05 to 0.06 oz/st), which of course would not have been economic with gold at $1.13/g ($35/oz), which was the fixed price at that time. The company, for obvious reasons, did not discourage the rumor. When, however, the true results came out there was a lot of feverish activity. Because of this new type of so-called "invisible gold," which could have been missed by the old timers, the thought was that other deposits would be discovered very soon. The Carlin and other smaller deposits, including Bootstrap, Blue Star and Gold Quarry, were in windows where the lower plate of the Roberts Mountain formation (below the thrust) had been exposed by erosion. This then became the new

Jan 1, 1997

By B. A. Hancock

It is not at all surprising that causticized potato and potato derived amylopectin starch solutions performed much better than their parent starches. Some preparation is required to rupture the starch granules to effect the polymeric adsorption and interparticle bridging necessary for selective flocculation. In laboratory work comparing the deslime performance of causticized and autoclave cooked laboratory corn starch solution preparations, it was found that higher deslime weight rejections, with attendant proportionally greater iron unit losses, occurred with the causticized starch. These results may be specific to the ore involved but they do suggest that cooking and causticizing cause different starch granule rup- ture and/or starch breakdown, which have an effect on desliming response. I calculated from the data in the article that the slimes product grades were high - 24.3% and 20.3% Fe when 53.7% and 54.5% Fe concentrate products were obtained, respectively, in Table 4, and 21% Fe with a 62.6% Fe concentrate in Table 5 - using the natural tailings sample, which had a head of 34.3% Fe. It may be advisable for the authors to consider different starch preparations in future investigations. The combination of upgrading and selectivity results presented in Table 4 are not as good as the authors suggest. The authors' claim that a system has been developed to produce saleable concentrates from the Kudremukh tailings is quite disconcerting. There are many hurdles yet to be crossed before commercial application of selective flocculation becomes possible because differ- ences between the very small-scale laboratory tests conducted and commercial application are rather large. Among the many differences are varying circuit feed grades that will occur from use of tailings, the apparent face that much lower tailings grades will be encountered in practice (it is much easier to achieve a high concentrate grade with reasonable recoveries using 34.3% Fe tailings as in the study rather than 25.3% Fe tailings grades that the plant apparently averages), the hydraulic nature of the thickeners used in operations compared to the static system used in laboratory tests, the different size distributions that will be obtained from a plant closed grinding- classification circuit, and differences in water used in a plant operation and the laboratory. The authors wrote that it was necessary to overgrind to be sure that the coarse gangue would not settle with the iron oxide floccules. This situation is likely to be exaggerated in commercial operations where it is assumed cyclones would be used for classification. Because cyclone classification is greatly influenced by particle densities, there will probably be an even greater difference in size between the iron and gangue particles in the plant, which would make the gangue slightly coarser still in relation to the iron. This would make the selective flocculation-desliming separations using the procedure employed by the authors even more difficult and, using the dispersant system the authors employed, greater overgrinding would be required. To grind finer to minimize the coarse gangue in the flocculated iron oxides is quite inefficient and appears not to broach the problem. The actual problem appears to be insufficient dispersion of the ground pulp. In this situation, addition of a dispersant would likely be required to attain a sufficiently high pulp dispersion level to efficiently effect a selective flocculation-desliming separation. Although the very coarse particles would still have a tendency to settle with the floccules, it probably would be found unnecessary to overgrind as much as indicated. Use of an optimum combination of dispersant and pH modifying reagents may also significantly improve the selectivity of desliming. Additionally, although it is possible that sufficient dispersion may be obtained by pH control alone in some situations, it is quite probable that added dispersity was obtained in the reported work from using distilled water. It is research experience that distilled water enhances dispersion. In commercial operations it may not be expected that sufficient dispersion will be obtained by pH control alone, unless the water used in the process is by nature quite dispersive. Overall, a change in the Kudremukh tailings dispersant scheme appears necessary where a dispersant is used in conjunction with a pH modifying reagent. With this change, different dispersion-flocculation responses will result that would have to be further evaluated. Therefore, it is still an open question whether an efficient and effective selective floccula- tion separation using Kudremukh tailings may be obtained that will produce saleable concentrates.

Jan 1, 1987

By Kenneth N. Han

The effect of anions such as sulfate, chloride, and nitrate on the solubility of rare earth element-bearing phosphates and fluorocarbonates has been examined using relevant thermodynamic data. The study has found that these anions have significant influences on the overall dissolution of rare earth elements (REEs) from various sources in the aqueous media. The thermodynamic calculations in this study have indicated that the speciation of REEs with the anions tends to increase the overall solubility of the REE-bearing minerals. However, in some cases, strong precipitation of REEs with some of these anions causes an adverse effect on the solubility of these elements. It has been found that sulfate ion has the most pronounced effect on the solubility compared to nitrate and chloride, in which the latter two acids exhibited almost identical results. The calculated results have indicated that REE-oxide (hydrated) is the easiest compound to dissolve, followed by carbonate, phosphate, fluoride, sulfate, fluoro-carbonate, and oxalate in that order.

By Nilufer Kursunoglu

Despite being crucial for meeting the world’s energy demands, coal mining operations present considerable health risks, especially when it comes to exposure to coal dust. Longwall coal mining, an advanced technique, has considerably enhanced production but made dust control more difficult. Due to its explosive nature, coal dust continues to be a serious and dangerous concern in the industry despite developments in safety technology. This study aims to comprehensively assess the risk of coal dust explosions by integrating fuzzy risk matrix and Pareto analysis methodologies, offering precise risk ratings for various root causes. A fuzzy risk matrix was developed to assign exact risk ratings to identified root causes. Additionally, a Pareto chart was used to identify the main reasons and suggest particular preventive measures. 17 significant causes were determined for underground coal mines based on the study’s methodology, and some preventive measures were suggested. In accordance with fuzzy risk rating values, root causes were divided into six categories. The root hazard with the highest risk rating was identified as ‘accumulation of inflammable gas/gases’. The proposed approach was used to examine incidents caused by coal dust explosions in Turkish coal mines in order to demonstrate the effectiveness of combining the fuzzy risk matrix and Pareto analysis for assessing coal dust explosions. This study gives not only a solid framework for risk assessment but also critical information for classifying the risks associated with coal mine dust explosions. The results highlight the critical necessity for proactive and preventive actions to reduce the hazards of coal dust explosions and protect the safety of mining workers.

Nov 16, 2023

By J. L. Carr

On average, one to two fatalities occur every year when a miner is struck or pinned by a continuous mining machine (CMM). Proximity detection systems offer an effective means to prevent these types of injuries and fatalities by sensing the presence of a miner and automatically disabling machine motion. The Mine Safety and Health Administration (MSHA) is expected to finalize a regulation that would require the use of proximity detection systems on all CMMs. Approximately one-fourth of all CMMs in use in the United States are already equipped with such a system. The use of these systems will save lives, not only by automatically preventing accidents but also by acting as a tool for educating miners and reinforcing safe work practices. Researchers at the Office of Mine Safety and Health Research (OMSHR) have conducted a series of field evaluations to quantify the performance of these systems in an underground coal environment. These field evaluations measured the distance at which the warning and stop zones were activated. The technology provided consistent activation of the warning and stop zones under many of the conditions tested. However, some performance inconsistencies were observed which present opportunities to improve the reliability of proximity detection system technology. Some of the notable variables that caused significant changes in performance and should be considered in the design, implementation, and use of proximity detection systems include interaction with the trailing cable, performance changes during tramming, and orientation of the miner-worn component.

Feb 23, 2014

By John G. Kuhm

INTRODUCTION The goal of any development drilling program is the drilling of the exact number of holes necessary to adequately determine grade and dimensions of an orebody, no more and no less. This cost of development drilling represents a large part of the total development budget. Any technique which eliminates unnecessary drilling costs produces beneficial results, both to the company and to the person who can claim the savings. This high development cost factor has produced a quantity of research to develop techniques, for efficient control of drilling programs. Recent mining literature is full of articles describing methods for measuring and increasing the effectiveness of these expensive projects. The method described in this paper was developed during the latter part of a major development drilling project at the Christmas Mine of Inspiration Consolidated Copper Company in Gila County, Arizona. It was used to test the need for drilling individual holes but could be expanded to control an entire project. This technique was used at Christmas as a practical means for evaluating adequacy of drilling on an individual hole basis, and it was possible to eliminate several previously approved holes. Application of this method involves an evaluation of geological conditions present and a good bit of common sense. Projects involving large disseminated, rather homogeneous orebodies of the porphyry-copper type would be the most likely candidate for application of the technique. Discontinuous orebodies would not be amenable to this technique. The method is based on the measurement of the linear correlation between the spatial distribution of two sets of data. The Pearson product-moment correlation coefficient is used. This is a standard statistic commonly used to measure linear correlation between two sets of data. The statistical significance of the correlation coefficient at a 95% confidence level is the main criterion for controlling the project. For purposes of reference we will call our technique the "correlation method" in this paper.

Jan 1, 1971

By R. Bissonette, C. Sunderman, R. Jacksha

The ability of wireless systems—e.g., voice communication, proximity detection, tele-remote operation, telemetry, etc.—to function satisfactorily (coexist) in the presence of other wireless systems is critical to the safety and health of mine workers. The failure of wireless systems to coexist could result in the delay, corruption, or outright loss, of critical data, voice, or video information. However, no mining-sector-specific regulations, standards, or guidelines exist to ensure the safe coexistence of wireless systems. This paper introduces the concept of wireless coexistence, implications in the mining industry, and the National Institute for Occupational Safety and Health’s (NIOSH’s) research approach to develop guidelines and recommendations to ensure safe wireless coexistence.

Feb 1, 2023

By J. O. Marsden

This paper presents an overview of the innovative methods for gold and silver extraction that have been put into actual practice in North America since 1987. It is an update of a paper of the same title which was presented by Sass and Mansanti at the AMC conference in San Francisco in 1987. New practices, new equipment and variations of older practices that have been introduced in the last five years are discussed in each of the following categories: process selection trends, comminution, gravity concentration, flotation, oxidation, agitated and vat leaching, heap leaching, solution concentration and purification, zinc precipitation, electrowinning, tailings disposal detoxification, and reclamation.

Jan 1, 1993

By P. Somasundaran, S. Venkatachalam, N. K. Khosla, K. Khosla

Froth flotation of alumina particles was studied by employing and hydrogen gas bubbles. The effect of current density was studied over a range of 0.1 to1.6 kA/m2. Compared to hydrogen, the use of oxygen bubbles at any given gas evolution rate resulted in a higher flotation recovery. The existence of an optimum current density, for each given particle size was indicated. Flotation at higher current densities resulted in better flotation kinetics, while flotation at lower current densities resulted in better Spas utilization.

Jan 1, 1996