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By MAURICE SUNKPAL, 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 at around 1.5 per year between 2009 and 2021. To date, a standard methodology for the design of coal rib supports that can accommodate the wide range of unique conditions found in coal mines in the USA 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 is proposed, beyond which the addition of primary support cannot stabilize the failed rib structure.

Aug 5, 2023

By Isabel F. Barton, MarkD. Barton, Christian A. Rathkopf

Molybdenite is the world’s principal source of rhenium (Re), a critical element in multiple high-tech applications. However, the Re contents in molybdenite vary by orders of magnitude on scales ranging from single grains to whole deposits. In order to better understand the systematics of this variation and what geochemical factors control molybdenite Re concentration, and hence overall Re resources, we examine global patterns in molybdenite Re contents through a compilation of > 3000 measurements of Re in molybdenite from > 700 mainly ore-bearing moderate- to high-temperature hydrothermal systems of different types. Our results are similar to but expand on those of earlier studies. Rhenium concentration in molybdenite has a lognormal distribution and varies systematically with type of geologic system, intrusive lithology, and Mo grade. The lowest-Re molybdenite occurs in greisens (geometric mean 1 ppm ± a multiplicative standard deviation of 9), quartz vein-hosted W-Sn deposits (2 ± 5 ppm), unmineralized granites and granodiorites (12 ± 8 ppm), intrusion-related deposits (24 ± 8 ppm), and porphyry W-Sn deposits (16 ± 11 ppm). Rhenium is most enriched in molybdenites from volcanic sublimates (23,800 ± 5 ppm), with skarn Fe and Au (560 ± 5 ppm and 540 ± 3 ppm respectively) and porphyry Cu and Cu-Au deposits next (470 ± 4 and 430 ± 7 ppm respectively). Among porphyries, skarns, and quartz vein-hosted deposits, Re is most highly concentrated in molybdenites from Cu and Au systems and its concentration decreases systematically through Cu-Mo, Mo, Sn, and W deposits. In nearly all cases, molybdenites from systems associated with intermediate igneous rocks contain more Re than molybdenites from systems of the same type with more felsic rock associations. The disparity between Re contents of molybdenite in felsic and intermediate systems is largest for porphyries, quartz vein-hosted, and skarn deposits and is near zero for subeconomic or barren granite and granodiorite Mo systems; felsic intrusion-related deposits have slightly higher molybdenite Re than their equivalents associated with intermediate intrusions. In most systems, molybdenite Re content does not correlate with metal grade, but may have an inverse correlation with Au grade in intrusion-related deposits (based on a small number of data points) and does exhibit a strong inverse correlation with deposit Mo grade. Dilution of Re through larger amounts (higher deposit grades) of molybdenite explains about 40% of this correlation, but the relative enrichment of Re in molybdenite from low-Mo deposits must also reflect some selective enrichment of Re/Mo in porphyry Cu systems compared to porphyry Mo systems. We found no evidence for secular increase or other systematic temporal variation in molybdenite Re content. The data regarding the use of molybdenite Re content as a proxy for mantle influence are ambiguous. Nearly all observed empirical correlations can be traced back to differences in redox state and sulfide concentration, the two geochemical factors identified here and by previous experimental work as the controlling influences on Re mobility under hydrothermal conditions. Hydrothermal systems with reducing conditions (W- and Sn-rich) tend to have low molybdenite Re even though compiled whole-rock data indicate that their source rocks have as much or more Re as those of more oxidized systems (e.g., Cu-rich). Vapor-phase exsolution, crustal assimilation, and mixing with external fluids may all enrich molybdenite Re concentrations in individual deposits and deposit types, but their extent and importance in overall hydrothermal concentration of Re is uncertain. Thus, it appears that the available molybdenite Re resource in an ore deposit largely depends on how the deposit’s redox and sulfidation conditions have varied over time and space during the timespan of hydrothermal activity. Oxidized, high-sulfide conditions tend to concentrate Re in molybdenite, whereas reducing conditions tend to leave Re dispersed at low concentrations in the bulk rock.

By Lei Zhang, Zhibo Ma, Pingye Guo

The mechanical behavior of hard roofs often leads to significant deformation and strong mine seismicity in gob-side entry retaining (GER) of the longwall mining, affecting the mining safety and efficiency of coal mines. Field investigations indicate that the failure process of the retained entry by hard roofs can be divided into three stages. In this paper, theoretical analysis and numerical simulations are first used to evaluate the damage process of the retained entry under the hard roof. The results show that the instability of the retained entry in the conventional GER method results from the combined static-dynamic loads (in stage II and stage III) from the hard roof. Then, an innovative method, through fracturing roofs to maintain entry (FRME), based on GER with the hard roof is introduced to solve this problem. In this method, the roof of the mining area is pre-fractured, and when the coal seam is mined, the direct hard roof will collapse along the fractured surface due to its deadweight and mining pressure. The hard roof control effect on the static and dynamic loads, taking into account the innovative method, is verified with simulations and theoretical analyses. The results are compared with those of the conventional GER method. Furthermore, based on the numerical simulation study of dynamic responses, it is found that compared to the GER method, the dynamic response of the retained entry using the innovative method is more moderate and more conducive to the stability of the entry. Field application also validated the effectiveness of the innovative method to control the mechanical behavior of the “massive sudden collapse” of the hard roof.

Aug 14, 2023

By Mehmet Guney

Acceptable working environments in production areas and service openings of Mindola Mine on the Zambian Copperbelt are currently maintained with an air volume of 1000 cu m/s and 5500 kW of installed refrigeration plant. It is planned to open up additional deeper areas of the mine for future production. Temperature problems, due to the increased heat generation from various underground sources are recognized as potential hinderances to efficient production. Findings of research work indicate that the present total heat output of 58,000 kW will inevitably increase by the second five-year period of the deepening program to 52,900 kW and by the fifth five-year period to 63,200 k\V. Estimated WB and DB temperatures, representing the total heat generation in stope faces and development headings during these same periods are predicted to increase to 53.8/40.2 °C and 35.5/42.1 °C respectively. It is assessed that the nominal refrigeration capacity needs to be progressively increased to 9,000 k\V This paper deals with the thermal problems in Mindola Mine, current heat output calculations, future prediction of temperature and heat load in various working areas of the mine with the application of computer programs and interpretation of data in order to find the effective solution for long-term planning of ventilation and air-conditioning.

Jan 1, 1982

By Yuchen Wu, Jing Mei, GUOZHONG LIU, KE GAO, Songjie Tian, Jianwei Cheng, Yu Wang, Yuhang Wu, Hui Wang

The stability and integrity of protective coal pillars between the longwall working face and development roadways are greatly affected by the mining activities such as the development face advancing or retreat mining. Due to underground strata pressure, the coal pillar cracks could be further developed and then the air leakage passages (mainly cracks) could be produced greatly. The air circulation caused by the atmospheric pressure difference between the two sides of the coal pillar may also provide suitable conditions for the possible spontaneous combustion of the residual coal in the mine gob. In this paper, a typical the protective coal pillar in a Chinese coal mine is studied. The deformation and stress of coal pillars during the mining process of longwall working face are simulated by using FLAC3D. The displacement and strain of coal pillars are analyzed. Two indexes, the linear fracture rate (LFR) and area fracture rate (AFR), are introduced to characterize the development of internal cracks’ development within coal pillars. By using field measurement, the air leakage intensity distribution due to coal pillar cracks is obtained; the equation of air leakage resistance of coal pillar under the influence of mining is derived. Then, the possible location of large air leakage of coal pillar is determined, which can be used as guidelines to study the possibilities of coal spontaneous combustions happening in pillar or mine gob. The research results help to control such potential hazards in advance.

Jun 16, 2022

By J. Kalliokoski

As part of a funded project we have completed a comparative literature survey of the US regional Proterozoic geology. For our com¬parative study model we used the geological settings of the very large pitchblende deposits in Saskatchewan, Canada and the Northern Terri¬tory of Australia. These occur in a variety of Archean and Lower Proterozoic host rocks, at and below a regional unconformity that is capped by subhorizontal Middle to Upper Proterozoic conti¬nental sedimentary rocks. We find that there have been recorded in some parts of the US similar unconformable sequences of Middle to Upper Proterozoic continental sedimentary rocks that overlie older basement rocks.

Jan 1, 1979

By John C. Gould

GENERAL DESCRIPTION The Climax mine is located 161 km (100 miles) southwest of Denver on Fremont Pass astride the Con¬tinental Divide. Climax employs 2800 people to pro¬duce 42 670 t/d (47,000 stpd) of molybdenum ore. The Storke level, the main underground level, produces 19 970 t/d (22,000 stpd); the 600 level, the lowest ac¬tive underground area, produces 7260 t/d (8000 stpd); and the open pit presently produces 15 430 t/d (17,000 stpd), Fig. 1. A panel caving method is possible because of the favorable geology and geometry of the Climax ore body. Geology The Climax deposit is composed of three separate ore bodies formed in a massive Tertiary stockwork which penetrates Precambrian granites and schists. The ore bodies are elliptical in plan with overall dimensions of 914 x 1219 m (3000 x 4000 ft). In section the ore bodies are arcuate with an overall height of 457 m (1500 ft), although the depth is not firmly ascertained. Ore grade decreases rapidly into a highly siliceous cen¬tral core and more gradually toward the outer bounda¬ries of the deposit. The mineralization is relatively continuous in the ore zone and contains few dilution-causing waste inclusions. Molybdenite is contained in pre-ore fractures which were largely recemented during the three mineralization pulses. Postore fractures, with a spacing of a few inches to several meters (feet), render the deposit amenable to caving when it is undercut. The peripheral rock caves similarly to the ore, but the central core is competent and is difficult to cave. Caving Feasibility Successful caving action requires an undercut area of approximately 122 by 122 m (400 by 400 ft). When the undercut area exceeds this minimum distance, the arching effect is lost, and the ore column collapses. Ore extensions from the main cave area less than 122 m (400 ft) wide usually cannot be caved. Waste or low¬grade areas within the deposit can be left if they are large enough to carry weight around openings passing through them. Experience indicates that these areas should be over 30 m (100 ft) in length and width. Smaller unprofitable units of waste, such as postore dikes, are mined to avoid weight problems and result in ore dilution. Successful caving and ore recovery are dependent upon timely ore removal from the undercut zone because columns of immobile ore inhibit caving and concentrate weight onto the underlying workings. Climax experience indicates that a finger (drawpoint) spacing* averaging 10.4 m (34 ft) yields an excellent caving action and a 92.5% ore recovery at the estimated grade. An additional 15% of the ore is recovered-at a diluted grade (Fig. 2). Mine Layout The slusher drift is the primary production facility (Fig. 3). The ore is undercut and caved into the fingers,

Jan 1, 1982

By Hao Zhang, Mi Zhou, Lichun Jiang, Xihong Zhang

This study investigates the roof stability of underground limestone mining in the presence of stratified joints during the excavation process. An analytical solution is derived to predict the stresses and deformation of underground tunnel structures, with the structure simplified as a double-hinge arch. The accuracy of the analytical model is validated through comparison with finite element modeling results. A parametric study is then conducted to assess the influence of different design parameters on the optimization of tunnel stability during excavation. The results indicate that the best stability of the tunnel can be achieved in engineering practices when the applied load (q) is 5 MPa and the joint inclination angle (θ) is 45°, with the roof span and thickness set at either 12 m and 8 m or 15 m and 10 m. These findings provide valuable design guidance for underground limestone mines with stratified joints.

Oct 18, 2023

By Glenn E. Hoffman

Relax: the world today is not running short of many mineral commodities, even though prices are spiraling upward. Einstein?s observation still holds true that everything is still relative, as long as one can pass on the incremental costs of production to the consumer. However, a strong consensus exists among many of the companies comprising the minerals processing industry that it is becoming increasingly difficult to find the easy-to-mine, high-quality resources that make juicy profit margins, which shareholders like. For the most part, the resource cherries have all been picked, and established mineral processing companies that are exhausting their reserves, or in the process of looking to add reserves so as to create longevity for their company, have no recourse but to look long and hard at what were once considered the dregs of the earth. Even end users, such as steel companies, are now looking to become ?miners,? to establish captive iron ore resources. Owning a captive resource facilitates a dependable supply of feed materials, which translates into effective cost control and a guarantee for a thriving sustainable business in the future.

Jan 1, 2012

By C. L. Kosmoski

In recent years, coal mining safety has attained national attention due to several highly publicized disasters. Despite these threats to worker safety and health, the U.S. relies on the mining of coal to meet its need for electrical power, with 42% of electricity in the U.S. generated through the burning of coal (U.S. Energy Information Administration, 2012b; United Mine Workers of America, n.d.). Furthermore, the production of coal continues to increase and reach record levels every year (National Mining Association, 2012; U.S. Energy Information Administration, 2012a). For this reason, the coal mining industry must continue to find ways to protect its workers while maintaining productivity. One potential approach to protecting workers while maintaining productivity is through improving the safety culture at coal mines. In order to achieve this culture, operators, employees, the inspectorate, etc. must share a fundamental commitment to safety as a value. This type of culture is known in other industries as a ?safety culture,? and can be defined as the characteristics of the work environment (such as the norms, rules, and common understandings) that influence facility personnel?s perceptions of the importance that the organization places on safety. Fundamental to improving safety culture within an organization is understanding the existing safety culture; improvements cannot be made without first elucidating areas that require improvement. This manuscript details the process undertaken by researchers at the National Institute for Occupational Safety and Health (NIOSH) Office of Mine Safety and Health Research (OMSHR) to assess the safety culture at five underground coal mines across the United States. The purpose of this manuscript is to present the aggregated results from those five safety culture assessments conducted at underground coal mines. Additionally, some general recommendations for areas where coal mines might focus their attention as part of their attempts to improve safety culture are offered.

Feb 23, 2014

By J. A. Organiscak

The National Institute for Occupational Safety and Health (NIOSH) Office of Mine Safety and Health Research (OMSHR) has recently studied several redirected scrubber discharge configurations in its full-scale continuous miner gallery for both dust and gas control when using an exhaust face ventilation system. Dust and gas measurements around the continuous mining machine in the laboratory showed that the conventional scrubber discharge directed outby the face with a 12.2-m (40-ft) exhaust curtain setback appeared to be one of the better configurations for controlling both dust and gas. Redirecting all the air towards the face equally up both sides of the machine increased both the dust and gas concentrations around the machine. When all the air was redirected towards the face on the off-curtain side of the machine, gas accumulations tended to be reduced at the face at the expense of increased dust levels in the return and on the curtain side of the mining machine. A 6.1-m (20-ft) exhaust curtain setback without the scrubber operating resulted in the lowest dust levels around the continuous mining machine, but this configuration resulted in some of the highest levels of dust in the return and gas on the off-curtain side of the mining face. Two field studies showed some similarities to the laboratory findings, with elevated dust levels at the rear corners of the continuous miner when all the scrubber exhaust was redirected towards the face either up the off-tubing side or equally up both sides of the mining machine.

Feb 23, 2014

By Mark C. Russell, Charles T. Russell, Brooke Colburn, Parker Bailey, Desiree Bernhard, Kaito Durkee, Tristen Cutshall, Rob P. H. Urselmann

Hypersonic tunnel boring (HTB) uses directed, high velocity projectile impacts to remove rock. HTB is highly independent of rock strength and targets hard, abrasive rock and variable rock environments. This paper discusses history, theory and presents test (2×2 m and .75×.5 m tunnels) field trial results and technical and economic evaluation. The HTB tests are compared against the historic ‘project REAM’ and the concept against tunnel boring machines and drill & blast in design studies for full-scale 12 m and 4.5×5.5 m tunnels. The positive results indicate that under these conditions HTB can replace Drill & Blast and Tunnel Boring Machines.

Jun 13, 2021

By A. Bakalarz, M. Duchnowska

"The copper sulfide ore from the Legnica-Glogow Copper Basin in Poland, which is processed by KGHM Polska Miedz S.A., consists of three lithological fractions: dolomitic, sandstone and shale. The copper ore, depending on the mining area, contains all fractions in different ratios. The lithological fractions differ in the particle-size distributions of their useful minerals. In the dolomitic part of the ore, most of the metal sulfide particles are 50-200 microns in size, while those in the sandstone and carbonaceous shale are 30-200 and 5-40 microns in size, respectively. This complex composition results in different mineralization, physicochemical properties and, in particular, upgrading selectivity. Due to the specific nature of the examined copper ore, two of the crucial parameters determining the beneficiation efficiency are size and liberation of sulfide particles. We analyzed the influence of grain size of feed on scavenging selectivity in one of the KGHM flotation plants and performed evaluation and analysis of the upgrading results of industrial beneficiation of the ore, containing mostly dolomite and shale. The upgrading efficiency was estimated using a selectivity indicator obtained by means of the Fuerstenau upgrading curve, which was then correlated with the particle-size distribution of the flotation feed.IntroductionFroth flotation is used extensively in commercial ore treatment flotation plants to separate minerals from gangue and to separate minerals from each other. The floatability of ores and their benefication selectivity are determined by many factors. One of them is particle size. The effect of particle size on flotation performance has been widely studied by many authors (Gaudin et al., 1931; Trahar, 1976; Trahar and Warren, 1976; Lynch et al., 1981; Pease et al., 2004). Particle size also plays a critical role in the probability of particle collision and attachment to bubbles (Solari and Gochin, 1992; Schulze, 1993), kinetic flotation (Taggart, 1956; Ahmed and Jameson, 1985; Debacher, 2002; Brozek and Mlynarczykowska, 2007; Potulska 2008), collector consumption (Sutherland and Wark, 1955; Vianna et al., 2003) and mechanical entrainment (Lynch et al., 1981; George et al., 2004; Konopacka, 2005). Most of the results have shown that there is an optimum particle size range for a given flotation system and that the simultaneous optimization of the recovery and concentrate grade of these particles in conventional flotation equipment is extremely difficult."

Jan 1, 2015

By Wern-ping N. Chen

Pressure tunnel includes water-supply tunnels and tunnels for hydroelectrical power plants. Their design principles are different from those of water/wastewater/conveyance and transportation tunnels. Though theoretical pressure tunnel design guidelines exist, a comprehensive guide to summarize these guidelines to facilitate the planning and design of a pressure tunnel program is lacking. The purpose of this paper is to fulfill this need. It first provides key pressure tunnel design concepts in hard rock environment; then, it illustrates associated final lining design principles, including the design of un-reinforced concrete lining, steel lining, and reinforced concrete lining. Design diagrams and a flowchart for the reinforced concrete lining design are provided.

Dec 1, 2022

By Jianbo Hu, Tao Zheng, Zhao Qihua, Rui Su

Rock mass discontinuities play a significant role in evaluating the stability of rock slopes. It is necessary to collect and analyze these discontinuities to fully understand the mechanical and deformational behaviors of rock masses. Based on the digital images of some slope faces, this paper proposes a framework based on machine learning for detecting and extracting discontinuities. First, digital images and photo poses are combined, the original scene is reconstructed, and 3D point cloud data are obtained. All captured point clouds can be mapped with each image by the direct linear transformation algorithm. Next, the detection machine learning algorithm is used to segment rock mass discontinuities on 2D digital images. Finally, based on the mapping relationship, the 2D digital images that were segmented are transformed into 3D point cloud data. The geometric parameters of the corresponding discontinuities can be extracted by the DBSCAN and the Skmeans algorithm and the characterization of specific rules. The proposed framework was evaluated utilizing a quantity of slope digital images in the Yebatan hydropower under construction, located in Sichuan Province, China, as a case study. The results illustrate that most deviations are less than 10° for the dip direction and dip angle, which verifies the reliability of the proposed framework and maintains acceptable measurement accuracy.

Nov 6, 2022

By A. Bieg, M. Samson, Z. Bell, A. Zinsser, J. Schuyler

As the world progresses toward the renewable energy transition, the demand for minerals increases exponentially. According to the International Energy Agency (IEA) 2021 Net Zero by 2050 Report, critical mineral production needs to grow six times to combat climate change. It’s estimated that four to seven percent of global greenhouse gas (GHG) emissions come from the mining sector’s Scope 1 and 2 (direct and indirect) emissions. Including Scope 3 (upstream, downstream, and embedded) emissions, the sector accounts for 28% of global GHG emissions. In a typical mining operation, diesel fuel is the biggest source of direct GHG emissions, and haulage is often the main source of diesel emissions. Large haul trucks have been the industry standard for over 50 years and offer flexibility, relatively low capital cost (especially with equipment lease options), and guaranteed technology with known maintenance. The challenge will be to justify their continued use when there are more economical options that lead to significant reductions in GHG emissions. The purpose of this study is to compare five haulage options in terms of GHG emissions, using the methodology defined by the Greenhouse Gas Protocol Scopes 1, 2, and 3 (Category 2). Additionally, this analysis compares operating and capital costs associated with the five haulage options (emerging and proven technologies): 1) Diesel-Electric Haul Trucks, 2) Battery-Electric Haul Trucks, 3) Fuel Cell Haul Trucks, 4) Railveyor, and 5) Conveyor. This analysis looks at the emissions and costs associated with operating the equipment (fuel and electricity), the embedded emissions and costs associated with the construction and installation of the equipment itself, and the associated infrastructure (roads, powerlines, and rail).

Feb 1, 2023

By V. Miller

C. Hertzler The difference of means of milling characteristics by rock and ore type is tested in a statistical manner in the paper. The null hypothesis being tested is that the sample means of two different populations are equal. It is assumed that the null hypothesis is true when performing the test. Because of this, acceptance of this hypothesis leads to the conclusion that the sample means of the two different populations are equal (or there is no significant difference between the sample means), not that the samples come from the same population, as the author asserts. Rejection of the null hypothesis leads to the conclusion that the sample means of the two different populations are not equal. A P value for this test is reported in Table 2 and is incorrectly identified as the probability that the two samples came from different populations. Actually, what was recorded was 100(1 - p) where p is the usual p-value. The usual p-value measures, on a scale of zero to one, the extent to which the data agree with the null hypothesis. It is the probability of observing a value of the test statistic (Z score) that is at least as contradictory to the null hypothesis as the one computed from the sample data (McClave and Dietrich II, 1982). The author makes no mention of the fact that were these tests to be conducted together, the chances of determining that the null hypothesis is not true when it is really true are increased considerably (Snedecor and Cochran, 1980). This is a multiple comparisons problem and should be dealt with accordingly. The usual methods involve adjusting the level of the individual tests. Table 2 reports an incorrect 100(1 - p) value in the comparison of concentrate grades for other rock types with .30 to .69 % Cu vs other rock types with .70 + % Cu; 86 % should be 98.6 %. Although the article can be criticized for poor data presentation and discussion of that data, the point the author makes regarding the use of these statistics in the ore reserve estimation process is well taken. It does not appear that inaccuracies in the statistical interpretations have hindered that usefulness. References McClave, J.T., and Dietrich, II, F.H., 1982, "Statistics," Dellen Publishing Co., San Francisco and Santa Clara, CA, 766 p. Snedecor, G.W., and Cochran, W.G., 1980, "Statistical Methods," The Iowa State University Press, Ames, IA, 561 p. Reply by V. Miller After reading Ms. Hertzler's comments, it appears that the differences noted were basically semantics. If the paper is ever presented at a math symposium, some changes will be needed. When deciding the best way to present the data to mining engineers and geologists, I decided that presenting the probability that the two sample groupings came from different populations (reverse to the Null hypothesis) was easier and would require less verbage to explain. Hence, the 100(1 - p) difference which was noted. Ideally, I would have preferred to give the recovery and concentrate grades, then comment which ones had significant differences. Unfortunately, Kennecott Corp. did not grant permission to publish the ore body's actual recoveries and concentrate grades. The conclusion is that there was, as expected, a relationship between the ore's geologic properties and the metallurgical properties, and by determining this relationship, the overall mineral model was greatly improved.

Jan 8, 1984

By D. Blackwell, W. R. Reed, G. Ross, S. Peters, M. Shahan

Dust collectors for roof bolting machines generally use a dry box to collect the roof bolting material. Recently, an underground mining operation converted a dry box dust collector to a wet box dust collector with a unique exception from MSHA for testing purposes. Water is routed to the roof bolter from the main water line of the continuous miner. The wet box utilizes a water spray to wet the incoming material. Testing was conducted comparing the two different collector types. Respirable dust concentrations surrounding the roof bolter with the different collection boxes were similar. The main difference in respirable dust concentrations occurred when cleaning the dust boxes. The average respirable dust concentration during cleaning of the wet box was 0.475 mg/m3, and during the cleaning of the dry box the average respirable dust concentration was 1.188 mg/m3. The silica content of the roof material was high, ranging from 28.9% to 52.7% during this study. The results from this study indicate that using the wet box as a collector reduced exposure to respirable dust up to 60%when cleaning the collector boxes. INTRODUCTION Currently, the dry vacuum dust collection system is the standard dust collector on roof bolters in underground coal mining. It has been proven to be an effective dust control to prevent roof bolter operators’ overexposure to coal mine respirable silica dust during roof bolting operations [Potts eta l, 2011 and Colinet, 2013]. The dry collection system works as follows: drill cuttings are vacuumed into the precleaner. The precleaner dumps oversized material onto the mine floor while undersized material enters the dust box via vacuum airflow produced by a blower. The cuttings entering the dust box are either deposited on the floor of the dust box or into a dust collector bag. The vacuum airflow leaves the dust box through a final filter which removes any small particles from the airflow. Figure 1 shows the internal view of two dry collection system boxes; one that does not use the dust collector bags and another that uses the dust collector bags. Dust collectors that use dust collector bags to contain material inside the dust collector box have been shown to provide better protection to roof bolter operators than dust collectors that do not use the bags [Listak and Beck 2008]. The dry collectors are very efficient at reducing respirable coal mine dust exposure from roof bolting activities. Recently, Blue Mountain Energy’s Deserado underground coal mining operation, outside Rangely, CO, made a modification to the dry collector system to convert it to a wet collector system with a unique exception from the Mine Safety and Health Administration (MSHA) for testing purposes. The wet collector system still uses a vacuum system to pull material from the vertical roof bolt hole, but incorporates a water spray nozzle located inside the dust box to wet the incoming material. The flow through the nozzle ranges from 1.9–7.6 lpm (0.5–2.0 gpm) at 0.69 MPa (100 psi). The conversion to the wet collector system removes the precleaner and the cyclones inside the dust box. The wet material (sludge) then empties through a specialized valve in the bottom of the box. The final filter is still used to prevent respirable dust from leaving the dust box through the exhaust. During operation, the wet collector dust box is opened and hosed clean after roof bolting every place. After every two entries the final filter is changed out. Figure 2 shows the inside of the wet collector box.

Jan 1, 2019

By L. L. Damon

A geochemical study of cyanide neutralization processes was conducted at Brohm Mining Corporation to gain a better understanding of the leach pad behavior and to develop a simplified procedure to quantify the attainment of compliance with permitted cyanide neutralization criteria. Such attainment is a prerequisite for off-loading spent ore from the leach pad at the "load/unload" gold heap leach operation. The geochemical study program consisted of size fraction analyses and weak acid dissociable (wad) cyanide analyses for various sections of the spent ore from the actual leach pad. Further testwork included size fraction analyses and wad cyanide analyses on the ore from four 29 foot laboratory test columns which were leached and neutralized with different neutralization treatments in parallel with an actual active leach pad. Wad cyanide was analyzed daily on the neutralizing columns and various parameters were monitored weekly on the influents and effluents during the leach cycle and neutralization period. A comparison was made of neutralized effluent chemistry and porewater extraction fluids, as well as an analysis of the selective adsorption of cyanide on various particle size fractions.

Jan 1, 1992

By M. A. Islam, H. Mischo, G. Barakos

Among the thirty-five declared critical raw materials (CRM) in the United States, there is no sustainable domestic production for the fourteen of them, while for some others domestic supply does not meet the national demand. Numerous potential critical raw material projects are in advanced exploration stages. However, the price uncertainties for many of these commodities put feasible future production into question and come into contrast with overly optimistic price forecasts published in all feasibility studies related to such exploration projects. Thus, in this work, a newly developed smart computational tool is used to evaluate the real perspectives of several CRM exploration projects implementing actual commodity prices, and to conduct reverse calculations in order to find the prices at which each project can be viable.

Mar 2, 2022