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By C. A. Oestreicher

For sulfide minerals, the presence of dissolved oxygen in the water phase is an influencing factor for separation by flotation. Klassen and Mokrousov(I:217l distinguish three stages of oxygen adsorption; namely, reversible adsorption, activated adsorption and, finally, surface oxidation. While contraversial, limited adsorption of oxygen is believed to improve the interfacial reactivity between collector and mineral. Intensive surface oxidation of sulfide minerals usually results in unsatisfactory flotation with collectors of the xanthate type. The influence of dissolved oxygen in the water phase has received consideration for sphalerite, galena, chalcopyrite, arsenopyrite, chalcocite, and covellite(2,3,4,5,6). However, the nature of surface oxidation products of sulfide minerals has received very limited investigation(6,7). Lightly oxidized galena promotes the adsorptive reaction of surfactant (xanthate or dixanthogen). Tolun and Kitchener(8) state that limited interaction occurs between galena and xanthate under relatively reducing conditions. And, that exposure to oxygen raises the electrochemical potential of the galena to that required for reaction with xanthate or dixanthogen. In the latter condition, the galena has an "effective flotation surface". On the other hand, intensive oxidation of galena reduces the chemical reaction with xanthate collectors (9) . Complete flotation with xanthate-type surfactants with increasing oxygen content in the water phase has been determined to have the following sulfide mineral sequence(3): galena, pyrite, sphalerite, chalcopyrite, pyrrhotite and arseno-pyrite. The work of Cusack(4) suggests that some threshold amount of oxygen is necessary for flotation to occur, but heavy oxidation hinders flotation with xanthates. According to Tolun and Kitchener(8) catalytic oxidation of xanthate to dixanthogen occurs. From infrared studies, Poling and Leja(IO,II) found that in the absence of oxygen in solution, xanthate ions do not adsorb either on

Jan 1, 1972

By I. Mikov

Established in the 60-ies, the Zlatna Panega quarry is a cement raw materials quarry supplying 95% of the needs of the Zlatna Panega Cement Plant with limestone, marl and other silica-aluminates. The two major types of raw materials are formed by accordingly dipping strata of Upper Jurassic limestones and lower Cretaceous marls. Between them there is a ?transition-zone? or ?contact-zone? of a high-silica content layer, about 5-m thick, defined as ?flint? by investigations including spot sampling throughout the quarry. For many years this ?contact-zone? layer was neglected due its negligible thickness. Further, with the increasing demand on the quality of the raw materials feed to the Plant and the focus on variability of the chemical and mineralogical components, it was decided not to mine any of the parts of the quarry contaminated by this material, for about three years. The latter had an impact on the availability of operating benches that could result in a substantial decrease in the total recoverable re-serves of Zlatna Panega, particularly with respect to marls. There were considerations for resorting to a partial, selective excavation of the quarry pit with a reduced lifetime for the lime-stones and re-opening of an other (currently in-active) marl quarry at a distance of 6 Km from the Plant. The Quarry Team of the Zlatna Panega quarry was assigned to study the problem. The ?contact? layer was duly confirmed with systematic geological (including mineralogy) studies, statistical analysis of available data, surface sampling and shallow exploration drills, resulted into distinguishing a third type of reserves and their credible subsurface delineation by means of geological 3-D modeling and quarry design. A new quarry design was finally elaborated with a view of providing flexibility in mining, utilization of the total volume of all available reserves and maximizing the mine life of Zlatna Panega quarry. An industrial test took place in December 2005, by using the contaminated parts of the quarry (co-mining of the ?contact-zone? materials) at a contribution of 15-20% in the raw materials mix, as dictated by their contribution in the total recoverable resources of the quarry. The purpose was to investigate the viability of this solution ? in line with improved quarry planning ? without compromising the quality of cement production and the efficiency of Plant Operations. Results were overall successful.

Jan 1, 2006

By YUSONG MIAO, DI WU, XIANRONG YAN, YIPING ZHANG, JIE LIN, KEBIN LI

Rock fragmentation size distribution is often used as an important index to account for the blasting effect because it directly affects the subsequent loading, transportation, and secondary crushing. Due to the mismatching of explosive and rock wave impedance, high boulder yield often occurs which affects the blasting effect. In this study, methods of measuring rock acoustic impedance, rock strength point loading, and detonation wave velocity have been used to obtain more accurate input parameters. Then, in the watershed image segmentation technique, the Gates-Gaudin-Schuhmann and Rosin-Rammler distribution functions have been used to analyze and quantitatively describe the rock fragmentation size distribution in the existing muck-pile. Finally, taking the rock properties, explosive performance, blasting parameters, and system characteristic variable into consideration, support vector machine (SVM) regression model has been analyzed on the learning and prediction of samples. The results show that SVM has a good prediction accuracy, high precision, and strong generalization ability. The optimized matching coefficient of rock and explosive wave impedance K ranges from 2.50 to 2.58 times. This study has developed a series of simple, accurate methods for rock properties analysis, detonation wave velocity measurement, and muck-pile model image processing, and a basis for predicting and evaluating rock fragmentation size distribution and optimizing the matching coefficient before carrying out a blasting operation.

Jan 22, 2021

By Abdelaziz Baçaoui, Imane Aarab, Mohammed Derqaoui, Abdelrani Yaacoubi, Khalid El Amari, Abdelmoughit Abidi, Abderahman Etahiri

Numerous researchers have addressed the complexity of the direct flotation separation of apatite from low-grade phosphate ore. The majority of these studies were limited to the micro-flotation scale. In fact, there is no warranty that a reagent scheme that showed a direct flotation of apatite at a micro-flotation scale on pure minerals would automatically replicate the same expected results on the natural ore in the mechanical flotation cell and so was the case for the phosphate ore studied. The overall trend was concentrating apatite in the non-float product. This study aspired to thoroughly assess the flotation performance of this never-exploited low-grade sedimentary phosphate using design of experiments (DOE). Unveiling new insights on such a strategy twist, various parameters and operating conditions, namely, desliming, type of collector and depressant, reagents dosages, conditioning times, flotation time, and pH, were addressed by multiple-factor screening using asymmetrical fractional factorial design followed by statistical interpretations. The collectors tested were the sodium oleate and a fatty acid mixture (FAM). The depressants tested were sodium alginate (NaAL), corn starch, carboxymethylcellulose (CMC), and sodium silicate. The results of the multivariable analysis revealed that desliming, pH, collector dosage, and flotation time significantly influenced the flotation process. Desliming was beneficial for the yield, P2O5 recovery, and SiO2 removal; however, there was still a compromise between grade and recovery. A strongly alkaline pH value of 12 stood out from the whole range. P2O5, CO2, and SiO2 grades of the products and flotation recoveries were improved, except for P2O5 recovery, which was negatively affected. A collector dosage between 2000 and 3000 g/t increased the P2O5 grade and reduced CO2 grade and recovery, enhancing the apatite/carbonates separation. A flotation time of 15 min was significant for P2O5 grade and CO2 recovery, nevertheless, at the expense of yield and P2O5 recovery. The products having initial concentrates of 28.08% P2O5 with a recovery of 88.96% and 29.41% P2O5 with a recovery of 62.71% were obtained using 25 g/t of NaAL and 2000 g/t of the collector FAM and sodium oleate, respectively, at pH 12.

Jun 5, 2022

By SANJIT KUMAR PAL, Sarvesh Kumar, Arindam Guha

The integrated geophysical study around pipe 1 and pipe 2 in Wjarakarur Kimberlite Field (WKF) assisted in comprehending the response of pipes. In the present study, we have attempted a combined analysis of very low-frequency electromagnetic (VLF-EM), electrical resistivity tomography (ERT), and magnetic methods to analyze the response from pipe 1 and pipe 2. A comparative analysis along all the profiles revealed the extension, depth, and geometry of the pipes. The results of VLF data and ERT data have been compared for two profiles of pipe 1 and pipe 2. Fraser-filtered grid map generated from all profiles of pipe 1 proficiently demarcated the trend of kimberlite pipe extension with high Fraser filter value associated with kimberlite. The ground-based magnetic data around pipe 1, pipe 2, and pipe 6 have been processed to analyze the magnetic behavior and subsurface structure based on the susceptibility contrasts. Analytical signal (AS) map revealed the extension of pipes, which was validated with the Fraser-filtered grid map. Analytical signal anomaly pattern reveals that pipe 1, pipe 2, and pipe 6 are trending in SW-NE, NE-SW, and NE-SW, respectively. Based on this study, the results indicate the conducting nature of pipe 1, whereas pipe 2 is resistive in nature. Further, the Euler deconvolution technique has been applied to calculate the possible depth of pipes from magnetic data. The comparisons of calculated depth ranges obtained from all geophysical techniques correlate well and validate the result (Table 1). The geological borehole data at pipe 1 and pipe 2 reported by the Geological Survey of India are well corroborated with our combined geophysical study.

Jan 17, 2024

By James C. Hower, B. K. Parekh

Introduction The purpose of coal preparation is to improve the quality of coal to make it suitable for a specific purpose by (1) cleaning to remove inorganic impurities; (2) sizing-crushing or screening, or both; and (3) special treatment, such as dedusting. The properties and quantities of impurities in coal are of major importance in the design and operation of steam generating equipment. Although boilers are often designed and equipped to use a wide range of coals, no boiler will perform equally well with all types of coals. All coals have certain properties that place limitations on their most advantageous use. These limitations are especially stringent for many of the older installations. Because of a general lowering of quality of raw coal in recent years, the need for coal cleaning has significantly increased. Among the factors contributing to the need for increased coal preparation are (1) increased demand for quality brought on by market and environmental requirements; (2) increased extraneous dilution caused by mine, health, and safety laws; and (3) depletion of the higher-quality coal seams. The major characteristics of coal and impurities from the standpoint of mechanical cleaning are particle size composition, generally called size consist; density differences; the relative friability of the coal and its associated impurities; hardness; and strength. Some of the other factors related to preparation are petrographic composition and rank. RANK CLASSIFICATION Coal is a general designation for a great number of solid organic minerals with different compositions and properties; all are rather rich in carbon and have a dark color, generally black. Next to mineral oil and kerogen, coal is the most important reservoir of carbon in the world. Carbon is an indispensable necessity of life as well as the main source of energy. Coals are found in stratified deposits, sometimes at great depths. All coals owe their origins to slow decomposition and chemical conversion of immense masses of organic material. As found in seams, coal is a heterogeneous mixture of organic and inorganic materials. Not only are there large differences in the properties of coal originating from different seams, but also in coal removed from different locations in a single seam. A genetic relationship exists between peats, brown coals, lignites, biturninous coals, and anthracites. As a whole, the process of coal formation, or coalification, proceeds as a continuous transformation of plant material, each phase characterized by a degree of coalification, or rank. As a measure of this rank, the carbon content or some related parameter can be used.

Jan 1, 1991

By John F. Clarkson

White Pine Copper Company built and operated, from August, 1962 through May, 1963, a pilot plant to recover copper using a cyanide percolation leach of the sand fraction (+400 mesh) of the mill tailings (1, 2, 3). The pilot plant runs were metallurgically successful and preliminary engineering studies for a full scale operation were completed under contract by an outside engineering firm. Unfortunately, the net return projected at that time did not justify the proposed capital' investment. In the summer of 1973 the project was revived. The rise in selling price of copper over the previous ten years, compared to the projected operating cost of the leach plant, increased the calculated net return on investment with the most recent figures showing a 30% DCF. Basically, the reason for this favorable shift in profitability was that the cost for many of the reagents did not go up as fast in the previous ten years as did the selling price of copper. Using the Chemical Marketing Reporter costs, NaCN sold for 18.3¢/lb in 1963 and 20.75¢/lb in August, 1973. During this same period H2S04 went from $23.50 per ton to $35.90 per ton. In comparison, producer's prices for electrolytic copper were 31¢/lb in 1963 and 619 in 1973. Thus the price of copper went up 97% while the two reagents only increased 13% and 53%, respectively. The capital investment also escalated but, again, not to the same extent as the selling price of copper. These price differences no longer are valid' but they did provide justification for reexamining the project.

Jan 1, 1975

By Ruichen Ren, Xu Han, Caixia Li, Zengchuan Yue

In this paper, the size-density distribution of coarse middling coking coal (CMCC) with different grinding fineness and flotation concentrates was investigated. Then, the flotation behaviors and liberation characteristics of different size-density fractions in ground CMCC were analyzed, and the different size-density fractions were divided into five categories: mostbeneficial, secondary-beneficial, poorly liberated, insignificant-pollution, and serious-pollution. Based on the content of particles with different liberation characteristics, an evaluation index of the liberation characteristic of ground CMCC, namely, the liberation perfection index (LPI), was proposed. The results indicated that grinding changed the size-density composition of CMCC, which affected the flotation performance. Insufficient grinding and overgrinding were not beneficial to flotation. Insufficient grinding caused a higher content of poorly liberated particles and a lower content of most-beneficial particles. Overgrinding increased the content of serious-pollution particles and reduced the content of most-beneficial particles. The size order of the LPI values of different ground CMCC was consistent with the flotation efficiency, and the LPI values of the other two similar samples were overall positively correlated with their flotation performance, suggesting that the method is universal.

Feb 14, 2023

By T. Bednarski, T. McCallum, L. Moya, B. Gay

"Molybdenum recovery by solvent extraction has gained interest partly due to advances in concentrate leach technology. Historically, both molybdenum and rhenium have been produced using amine-based extractants with the ability to co-extract both metals. Alternatively, a phosphinic acid-based extractant (CYANEX® 600) can be used depending on the flow sheet. CYANEX® 600 selectively extracts Mo leaving Re in the raffinate, which is a valuable feature in plants which require rejection of the rhenium from the molybdenum circuit. CYANEX® 600 piloting results are reviewed and compared to commercial experiences with an amine-based extractant. INTRODUCTION Molybdenum is most commonly found as molybdenite (MoS2) in primary porphyry deposits (about 59% of global reserves) and copper porphyries (about 39% of molybdenum reserves)[1]. Recovery from molybdenite traditionally involves flotation (concentrates typically containing 85 – 95% MoS2) followed by roasting to produce MoO3. Solvent extraction is also used to recover molybdenum from acidic aqueous solutions resulting from scrubbing of roaster and smelter off gasses and leaching of molybdenite (uranium ores). Molybdenum is also frequently found in low concentrations in copper leach inventories[2]. More recently, pressure oxidation has gained interest to treat molybdenite concentrates. Molybdenum Chemistry The chemistry of molybdenum is complex with a series of extractable species existing dependent on the pH and oxidation state as shown in Figure 1.Under strongly basic conditions, molybdenum primarily exists as MoO42- while under strongly acidic conditions MoO22+ becomes the predominant species. In between, a number of isopolymolybdates such as Mo7O246- and Mo8O264- are dominant. Equations (1) – (3) demonstrate some of the proposed molybdenum equilibrium reactions[1]."

Jan 1, 2017

By J. R. Klaehn, M. Shi, T. E. Lister, L. A. Diaz, J. S. McNally

Since 2019, approximately two dozen international battery recycling companies have been operating at the commercial or pilot scale [1]. However, current process capacity does not exceed 100,000 tons of end-of-life (EOL) lithium-ion batteries (LIB) per year, and the current recycling rate is only about 5% in North America [2-3]. Recycling metals from EOL LIBs is not only essential to waste minimization and environmental stewardship, but also a source of critical materials, which can contribute towards supply chain stabilization. LIB recycling can be roughly divided into three categories: direct recycling, pyrometallurgy, and hydrometallurgy. Direct recycling seeks to retain as much of the manufactured effort with minimum chemical processing [4]. The evolution of battery designs and chemical compositions are challenges to this approach [5-6]. Pyrometallurgical processes are robust and flexible allowing a broad spectrum of batteries with minimum pre-processing requirements. Some of the matrix issues are solved through combustion of some constituents. Smelting is strongly tied to fossil energy and air emissions are a concern. In addition, pyrometallurgical processes are not stand alone, as hydrometallurgical refining normally required to separate and recover the metals of interest (Co, Ni, and Cu) from the produced alloy and Li from the slag [7]. Hydrometallurgical processes can be highly selective and efficient in recovering high purity materials [8]. In general, hydrometallurgical processes are considered cleaner and less costly than high temperature processes [4,9-10]. Extensive chemical consumption and waste generation significantly increase the operating costs and overall environmental impact of the hydrometallurgical processes [11]. Leaching of metals from LIBs is dominated by the mixture of sulfuric acid (1 to 4 M H2SO4) and hydrogen peroxide (1 to 15 wt.% H2O2) [10]. Although organic acids have been tested as lower environmental impact alternatives, the reported pulp density operations are low < 60 g/L [10-11]. For the H2SO4/H2O2 leachates, high leaching efficiencies >90% are reported for Co, Ni, Mn, and Li from cathode materials at temperatures over 40 ⁰C and pulp densities between 30 to 200 g/L [9-10]. H2O2 acts as reducing agent to convert Co+3 and Mn+4 into their soluble forms, Co+2 and Mn+2 [12]. Despite the claims of H2O2 being a clean reducing agent [13], production of H2O2 has significant environmental impacts as determined by life-cycle analyses [13-14]. Moreover, the costs and risks associated with H2O2 transportation and storage cannot be ignored [15].

Mar 2, 2022

By Jeff Poe

When a takeup weight is offset to the side of a conveyor, a wire rope is commonly routed from the takeup weight through a series of sheaves to the takeup carriage. The friction developed from a series of wire rope sheaves can effect the actual belt tension at the make up carriage. Through real-mime testing the actual tension on a makeup carriage was recorded with data acquisition using an internally strain-gauged turnbuckle. The wire rope used to transfer the makeup weight to the makeup carriage was routed through a series of lubricated bronzed bushed sheaves. The tension on the makeup carriage was found to be as much as 26% higher than the weight of the makeup due to the friction in the sheaves. A detailed discussion of the testing and results will be made.

Jan 1, 2000

By Herman L. Bauer

The Robinson Mining District near Ely, Nevada, is one of the major porphyry copper districts in the world. Within the district is an east-trending zone of altered porphyry, metamorphosed sedimentary rocks, and sulphide mineralization surrounded by Paleozoic limestone, dolomite, shale and sandstone, Tertiary volcanics, and Quaternary alluvium. Six major disseminated ore deposits have been exploited in this eight-mile zone. About 20% of the district production has come from disseminated ore in metamorphosed sedimentary rocks found adjacent to altered quartz-monzonite porphyry. The following brief statements summarize the inter-relationship of porphyry, alteration, metamorphism, and mineralization: 1. Disseminated copper ore does not occur in unaltered porphyry or in unmetamorphosed sedimentary rocks 2. Sedimentary rocks have been only weakly metamorphosed where intruded, by unaltered porphyry. 3. Intensely metamorphosed sedimentary rocks are found only adjacent to, altered porphyry. 4. Many hydrothermal minerals are common to both altered porphyry and adjacent metamorphosed sediments. 5. Disseminated ore is found in altered porphyry and adjacent meta¬morphosed sedimentary rocks. 6. In general, there is a direct correlation between grade of ore in porphyry and copper content in the adjacent rocks. 7. Chemical changes in the altered porphyry facies and metamorphosed rock facies appear to be related:

Jan 1, 1963

By Hans G. Machel

In low-temperature diagenetic environments (0 c T c 60 - 80°C), elemental sulfur forms by six common processes: via (1) inorganic oxidation of H2S by molecular oxygen, (2) inorganic oxidation of H2S by ferric iron compounds, (3) polysulfide disproportion- ation; and via the microbial metabolic actions of (4) colorless sulfur bacteria, (5) colored sulfur bacteria, and (6) some types of thiobacilli. Economically viable deposits of native sulfur usually are formed by only one process: inorganic oxidation of H2S by molecular oxygen. The amount of elemental sulfur is also limited by three common processes that destroy it in low- temperature diagenetic environments: bacterial reduction, bacterial oxidation, and inorganic oxidation. In high-temperature diagenetic environments (80 - 10o°C c T c 150 - 200°C), particularly in sour gas reservoirs, there are seven common processes forming elemental sulfur: (1) SOdH2S reaction, (2) H2S- hydrocarbon reaction, (3) polysulfide disproportionation, (4) catalytic H2S decomposition, (5) C02-H2S reaction, (6) chlorinification, and (7) exsolution of physically dissolved elemental sulfur. Natural processes that destroy or remove elemental sulfur in such settings are hydrocarbon sulfurization, polysulfide formation, and advection. In most sour gas reservoirs, relatively small quantities of elemental sulfur occur as a free phase (solid or liquid). However, elemental sulfur often precipitates in copious amounts from the reservoir fluids via exsolution and polysulfide disproportionation during production of sour gas reservoirs. By extension, elemental sulfur may precipitate from reservoir fluids in relatively large quantities upon uplift and erosion of deep sour gas reservoirs, or from advection of sulfurous fluids into shallower settings via faults or other high-permeability pathways.

Jan 1, 1992

By J. Patrick Powers, Joel Miller, Arthur B. Corwin

A combination of a slurry trench with dewatering successfully controlled groundwater for a large sloped excavation immediately adjacent to the Arkansas River. The Murray Hydro Electric Project required excavation through sand and gravel to bedrock, 85 feet below high water in the River. The perimeter system of deep wells pumping up to 25,000 GPM (1,600 1/sec) lowered the water level to 15 feet (4.6 m) above bedrock. A short soil/bentonite slurry trench near the toe of the slope cut off the residual head. The method cut six weeks from the critical path. Design of the slopes and the dewatering system demanded high safety factors because of risk to existing Lock and Dam No. 7. A localized pocket of soft organic clay was supported by a rock filled bolster. A major flood event was successfully withstood. The method can be applied to large civil excavations and open pit mines.

Jan 1, 1989

By Z. H. Xie, D. S. He, W. M. Xie

China has rich phosphate rock resources, but with their continuous depletion, the rich ore resources have become increasingly scarce, which makes the exploration of collophane more important. Medium- and low-grade phosphate rock cannot be utilized until their impurities have been removed through mineral processing. However, due to the similar physical and chemical properties of dolomite and phosphorus minerals, the separation of dolomite from phosphate

By Richard E. McGee

The paper presents a description of the procedures that were used to generate a ventilation criterion based on measurements of ambient mine air exhaust pollutant gases and ventilation airflow quantities. Today&apos;s economics dictate that a ventilation requirement be based on actual operating conditions. This will assure optimum economics and a safe working environment at the same time. The paper presents a methodology by which a ventilation requirement is based on the average maximum expected concentrations of nitric oxide (NO), nitrogen dioxide (NO2), carbon monoxide (CO), and carbon dioxide (CO2) that equipment operators in one of our Southeast Missouri mines will be subjected to. A graph has been generated whereby the expected maximum levels of NO, NO2, CO and CO2 are plotted against the ventilation quantity of a mine stope.

Jan 1, 1982

By H. J. Hutson

BRS, Inc. has utilized Carlson&apos;s Natural RegradeTM software to design the reclamation and stabilization of abandoned open pit uranium mine spoils, AML Project 16N, located in the Gas Hills Uranium District in central Wyoming. This work was completed for the State of Wyoming&apos;s Abandoned Mine Land Program (AML) on preSMCRA (Surface Mining Reclamation and Control Act) sites for which no reclamation obligation existed. The project successfully stabilized an eroding spoil dump, enhanced the local view shed, created habitat diversity and provided a significant source of fill material for open pit highwall hazard mitigation efforts. Additionally, the project served to evaluate innovative reclamation design methods, engineering software and grade control technology. This project was nominated for a National Association of Abandoned Mine Land Programs award by the State of Wyoming. The results of this project will be presented as a case history of the Natural RegradeTM approach to surface mine reclamation. Based on the successful application of this new mine reclamation technology on the AML 16N Project, reclamation designs for spoils associated with planned underground uranium mine operations at the Velvet Mine in the Lisbon Valley Uranium District of Utah for Uranium One USA, Inc have been prepared and will be presented. Critical elements will be the encapsulation of unsuitable materials, long term hydrologic stability, and sequencing and staging of ore and mine waste for final reclamation surface construction. The goal of the Uranium One Velvet Project will be to apply best available technology to reclamation planning and fully integrate the reclamation design with the mine planning. Benefits of the approach should include improved public and regulatory perception of the project, facilitate mine permit approvals and the eventual timely release of reclamation bonds, as it is anticipated the stable landform created through this process will result in more rapid stabilization of the site to pre-mine geomorphic conditions. NATURAL REGRADETM CASE STUDY AML Project 16N, D-9 and K Pit Reclamation Project, is located in an intensely disturbed area of the West Gas Hills Uranium Mining District in Fremont County, Wyoming. The overall design calls for the placement of approximately 7 million cubic yards of backfill materials in the D9 and K Pits, imported from adjacent mine spoils including the Central Spoils, and from local highwall excavation. Figure 1 shows the location of the Central Spoils excavation area, as well as the D9 and K Pits fill area. The entire project is located on public lands administered by the Bureau of Land Management, Lander, Wyoming District Office. No access controls are in place at the abandoned mine sites. The public lands are administered under a multiple use land policy including grazing of range land, public recreation, active mining and exploration, and oil and gas. The project site is located along Fremont County Road No. 5, which is a moderately well traveled road. Reclamation of the numerous and large abandoned open pits in the Gas Hills Uranium District has reduced the risk of death or injury to the public. Site History and Background The Gas Hills Uranium District was mined by multiple companies during the period from 1955 to 1981. The uranium was hosted in the Wind River formation, a sequence of soft sedimentary layers of interbedded sandstone, clayey shales, and conglomerates. Uranium was precipitated out of ground water in classic roll front deposits, which are found from very near surface to depths greater than 400 feet. Prior to the enactment of the Surface Mining Control and Reclamation Act of 1977 (SMCRA), very little reclamation work was performed when mines were abandoned or closed. The State of Wyoming Abandoned Mine Lands Program has worked to mitigate hazards associated with large open pits, control migration of unsuitable materials and restore watershed functionality of lands in the district. Additional reclamation is being performed by mining companies on bonded lands. Figure 2, Gas Hills Uranium District aerial photograph, shows the extent of the mining disturbance in the district. [ ] Project Hazards and Concerns The purpose of AML Project 16N is to mitigate hazardous conditions and risk to the public associated with open pit highwalls and toxic, acid forming, and /or radioactive mine waste materials. The D9 and K Pits are contiguous open pits characterized by an 8,800 foot long hazardous highwall that averages over 200 feet tall. In addition to the physical hazard presented by the highwall, the project addresses

Jan 1, 2009

By Richard J. Goff

Technical Support, one of several major organizational units in the Mine Safety and Health Administration, provides technical assistance to the mining community in all areas of health and safety. One of its many areas of expertise is noise control. Depending upon the needs of the operator or inspector, the services provided can range from locating major noise sources to actually building noise controls. Over the years a vast amount of knowledge has been obtained as to what will work and what will not. DESIGN GOALS While many noise controls have been developed for mining equipment and operations, the most acceptable (ergo successful) have achieved the following design goals: - a noise level reduction of at least 5 decibels; - minimum interference with production; - inexpensive to install; and - effective in reducing low frequency noise.

Jan 1, 1984

By Ed Ramsay

The Hill is a cyanidation plant with flotation and roasting. The grinding and treatment have at the present time a capacity of 1400 tons of ore per day and the precipitation has a capacity of 6000 tons per day. The ore is delivered by a 36" conveyor belt and tripper to 8 mill bins having a capacity of 6000 tons. The ore from these bins is discharged onto 20" gathering conveyor belts which are regulated with p.i.v. speed changers to control the belt weight. A weightometer determines the tonnage milled. The feed enters the 8&apos; x 12&apos; rod mill through a rubber lined chute and the product is discharged through a launder to a 7&apos; rake classifier in closed circuit with 10&apos; by 6&apos; hardinge ball mill. Lime slurry is added at the rake classifier to give a strength of .30#/ton. The classifier sand is fed into the ball mill with a scoop feeder and the mill discharge flows in a launder back to the classifier. The classifier overflow is pumped to a 24&apos; by 12&apos; bowl classifier in closed circuit with a 8&apos; by 15&apos; regrind mill. The bowl classifier sands are fed to this mill and the mill discharge is pumped back to the bowl. The bowl overflow at 8% - 200 mesh is the final product from the grinding circuit.

Jan 1, 1975

By S. D. Thompson

Several discrete-event simulation models have been developed for designing the conveying system capacities of underground coal mines. Most of these models require the use of a step-function approach to mimic the loading process at each conveyor loading point. Over the years, practitioners have generally employed one of three versions of the step-function approach. An important issue is how these three versions compare to each other in terms of simulated capacity requirements for belt conveyors. Another is-sue is how they compare to peak capacity design. This paper addresses these issues by statistically and quantitatively comparing their simulation results. The consequences and problems associated with using each version are also discussed and summarized.

Jan 1, 1997