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By J. C. Judd

The U.S. Department of the Interior, Bureau of Mines investigated solvent extraction methods for recovering gallium from filtrates produced from sulfuric acid leaching of hydrometallurgical zinc residues. These filtrates typically had pH values of 0 to 0.5, too high an acid content for effective gallium recovery by traditional extractants. Of the nine extractants studied in batch tests, 0.3M octyl phenyl acid phosphate (OPAP) in a kerosene diluent was the most effective extractant for gallium recovery, extracting up to 99.9 pct of the gallium. This high extraction was possible when the iron in solution was maintained in the reduced ferrous state. Batch tests also showed that a 1.5M-H2SO4 strip liquor recovered 98 pct of the gallium from the loaded OPAP. A continuous solvent extraction system was designed and operated based on McCabe-Thiele analyses of equilibrium loading and stripping isotherms. When 12 L of leach filtrate containing 0.32 giL Ga and 20 giL Fe were pumped through the system in 32 hr, results showed that 95 pct of the gallium was extracted in four stages at an A:O ratio of 1:1. Stripping the loaded solvent with a 1.5M-H2S04 liquor recovered 94 pct of the gallium in four stages at an A:O ratio of 1:4. A two-stage, 50-pct-phosphoric acid strip successfully removed ferric iron from the OPAP, allowing high gallium extractions to be maintained.

Jan 1, 1990

"Hecla Limited has been investigating the potential benefits of using a Spendrup/Compact Filter Technic (CFT) diesel filtration system to remove diesel particulate matter (DPM) from the mine atmosphere to optimize use of the available air. The National Institute for Occupational Safety and Health (NIOSH) has evaluated the scrubber efficiency for dust and DPM which is to be presented in another report. This paper reports on a new mining application of a carbon particulate sensor’s ability to continuously monitor long-term mine DPM levels in conjunction with the scrubber evaluation. Magee Scientific is the manufacturer of the Aethalometer®, a continuous black carbon monitor which has been used extensively in worldwide ambient air quality monitoring networks. Two Aethalometers were used at the mine to assess their potential to continuously track underground DPM levels. In this first underground mine application, the Aethalometers were used to continually monitor black carbon (BC) in the intake and exhaust of the CFT scrubber system. Results showed that an off-the-shelf Model AE33 Aethalometer could continually track underground BC levels for extended periods of time over the course of the six-week trial. Data at the scrubber intake over 14 days of testing showed that DPM filtration would be required less than half of the time to maintain BC levels below 50 µg/m3. If the Aethalometer BC data output were “smart coupled” to control a 200-HP in-mine scrubber, significant electrical cost savings would be realized. INTRODUCTION Mine ventilation airflow quantities are calculated to maintain a safe and healthy environment for miners. Dilution of diesel particulate matter (DPM) is often a key factor used to determine the required quantities of ventilation air (Bugarski et al. 2011). DPM emissions and exposure levels are highly variable depending on mine schedules and activities (Janisko & Noll 2010); however, lacking technology to continuously monitor levels, mine ventilation systems are often over-designed in order to handle worst-case scenarios to ensure that DPM levels are maintained at or below regulatory limits. As mines go deeper it becomes more difficult to supply air, and the development of new shafts for air is not economically feasible (Basu et al. 2008). Cleaning and reuse of available air then becomes an attractive option to maintain good air quality."

Jan 1, 2016

By Miles L. Silberman

Three major suites of igneous rocks were erupted in the Great Basin in middle and late Cenozoic time. The first, most likely, resulted from sub- duction of a Pacific plate (Farallon plate) beneath the North American plate, resulting in eruption of andesitic rocks in the northern Great Basin between 42 and 43 m.y. ago. Between 34 and 17 m.y. ago, igneous activity migrated to the southwest, largely as eruption of volumincus rhyolite-quartz latite ash-flow sheets, ending about 17 a.y. ago with the onset of crustal extension and north-south-trending Basin and Range (rift) faulting. The second volcanic suite erupted between 17 and 8 m.y. ago in the northern Great Basin; is mostly basaltic or bimodal basalt-rhyolite. The third volcanic suite consists of andesitic and rhyolitic lavas and some basalt erupted between about 22 to 6 m.y. ago in the western Great Basin along the northwest-trending Walker Lane fault zone. Hydrothermal ore deposits are associated with all suites of igneous rocks, but the largest production has come from mines associated with andesitic rocks of the third (western Great Basin) suite. The oldest (34 to 42 m.y.) andesitic rocks (first suite) of the northern Great Basin have produced some large deposits; slightly younger silicic rocks of this suite are mostly barren. Many small ore deposits are associated with the 17- to 8-m.y.-old basalts (second suite) of northern Nevada.

Jan 1, 1976

By G. Wang, L. R. Moore, Y. Xiong, A. Gorken

"Feldspar may represent 60% of the earth’s crust, but the mineral does not exist in a pure state. It is generally associated with quartz, iron containing minerals such as mica, and titanium containing minerals such as rutile. Feldspar itself can even exist in different lattices from those containing sodium to various divalent ions. As such, its beneficiation has become a process that is quite complicated, when compared to the beneficiation of other minerals. It is this complexity that has made feldspar beneficiation dependent on tailor made reagents to successfully yield the industry specified feldspar products. This paper reviews the mineralogical variations found from two feldspar mines and the development of collectors to specifically target a Fe+Ti concentration of < 0.05%, while maintaining a feldspar recovery greater than 90%.INTRODUCTION Feldspar is a class of aluminum silicates that represents the most abundant mineral on earth (Vidyadhar, 2002). This class of mineral can further be defined by the types or concentrations of cations present in the crystal lattice, MAlSi3O8 (M = K, orthoclase and microcline; M = Na, Albite; M = Ca, Anorthite) (Heyes, 2012). Feldspar is most commonly mined and processed for either the alumina used in glass making or for the alkalis used in ceramics. The alumina is known to improve product hardness, durability, and resistance to chemical corrosion. The alkalis act as flux, which means they not only lower the melting temperature of the mineral mixture, but also assist in connecting the other components present in ceramics. It was stated in 2009 that the world production of feldspar was in excess of 18.3 million tons produced, with the majority coming from China, Italy, and Turkey. Raw feldspar is typically found with quartz, iron containing minerals such as mica, titanium containing minerals such as rutile, and other feldspar minerals (Celik, 2001). Quartz is stated as having a negative effect on the melting point of the feldspar, whilst the iron and titanium based mineral contaminants are known to have negative effects on color and quality of the product produced; these mineral contaminants must therefore be separated prior to application (Barns, 1997 & Celik, 2001). Mineral flotation and magnetic separation are the most common means of feldspar beneficiation. However, conditions vary greatly from mineral contaminant to mineral contaminant (Table 1). Some literature suggests feldspar can be floated with an amine, but this generally yields lower than acceptable recoveries and often still requires the magnetic separation of iron based minerals to achieve the desired grade (Amarante, 1997 & Heyes, 2012). Feldspar flotation can be further improved with the addition of hydrofluoric acid (HF) as a silica depressant and feldspar surface modifier, but this still requires the removal of some of the other minerals first. Rutile can be floated with an anionic fatty acid collector at a pH range of 7-8 or with a petroleum sulfonate at a pH < 4 (Celik, 2001), while mica can be floated with a petroleum sulfonate or an amine collector at a pH of 3.5."

Jan 1, 2016

By R. L. Grayson

Compliance standards for respirable dust which are reduced because of high quartz concentrations have become an industry-wide problem. Research on the physical and chemical character of quartz-classified particles in respirable coal mine dusts indicates that they may exist in intermixed layers or may be coated by clay or coal material. The relative amount of intermixing varies with the coal seam being mined. If most of the quartz particles of a mine dust are "pure," then the dust may have a high fibrogenic risk potential because of truly free silica and the compliance standard should be lowered. If, however, the quartz is locked inside another material matrix, or coated, a reduced compliance standard may not be appropriate.

Jan 1, 1990

By R. R. Klimpel

The grinding of coal-water slurries has received increasing industrial attention during the last decade. In particular, there is long-term interest in the use of pulverized coal-water slurries to replace oil in combustion equipment and the development of coal gasification/liquefaction processes which require coal-water slurries as feed. More specifically, the use of coal-water slurries in gasification requires grinding a high density slurry containing the smallest amount of water consistent with slurry pumping and spraying. As part of a fundamental engineering research support program aimed at the industrial implementation of dense coal-water slurry grinding, this author has published several papers on how specific rates of breakage vary as a function of slurry rheology, Klimpel (1982, 1982/83). These papers demonstrated that there is a consistent pattern of change in specific rates of breakage of coal in dense slurries with controlled variations in slurry rheology. By matching rheological data with laboratory grinding results, it was possible to identify directly slurry conditions that correspond to: (1) slowing down of breakage rates, (2) the occasional acceleration of breakage of some sizes, and (3) conditions where chemical additives will increase rates of breakage.

Jan 1, 1990

By Yi Luo

This paper presents some of the results of a research on long-term subsidence over and near the chain pillar systems -the area available to long-term subsidence in longwall operations. Although the long-term subsidence in most longwall operation is insignificant, it does exist in some cases. For the cases where long-term subsidence is significant, it has been found that the development process can be expressed by an exponential decaying function. The stability factor of the chain pillar system and the potential for long-term subsidence are the two important factors that control the magnitude and development characteristics of the long-term subsidence process.

Jan 1, 1999

It has been reported previously that dolomite samples from different locations showed varying flocculation behavior with Poly(ethyleneoxide) (PEO). In this study it is determined that palygorskite associated with the Florida dolomite samples is responsible for their flocculation with PEO. A detailed characterization of the dolomite samples from different locations and its relevance to flocculation with PEO is discussed.

Jan 1, 1994

By G. L. Mowrey

A new coal interface detection (CID) method is under investigation by the U.S. Bureau of Mines. This novel concept is based on utilizing a passive infrared system (FIR) to measure temperature changes associated with the cutting of coal and adjacent strata. For this method to work successfully, a definite mechanical/hardness contrast must exist between coal and rock. Advantages include that FIR responds instantaneously at the coal-rock interface and it is minimally affected by dust and/or water sprays. Based on preliminary tests, results indicate that the PIR method can be applied successfully in mines having a hard immediate roof.

Jan 1, 1990

By R. A. Thomas, C. D. Litton

"The objective of this study was to develop a laboratory-scale method to rank the ignition and fire hazards of commonly used underground mine materials and to eliminate the need for expensive large-scale tests that are currently being used. A radiant panel apparatus was used to determine the relevant thermal characteristics of the materials: time to ignition, critical heat flux for ignition, heat of gasification, and mass loss rate. Three thermal parameters, TRP, TP1, and TP4, were derived from the data, then were developed and subsequently used to rank the combined ignition and fire hazards of the combustible materials from low hazard to high hazard. The results compared favorably with thermal and ignition hazards of similar materials reported in the literature and represent a simpler approach to quantifying these combustible hazards. INTRODUCTION The potential fire hazard from combustible mine materials is a major concern that can impact the safety of mine workers. Combustible materials tend to be solids and within this category there exist degrees of flammability and fire resistance. These materials may be readily ignitable or they may be able to withstand high temperatures and high heat fluxes. Presently, no simple methodology exists to assess these possible fire behaviors and quantify the resultant fire hazards. Thus, better knowledge of the thermal properties and chemical characteristics of combustible materials is of great importance to the safety of mine workers. Tewarson [1994] developed a methodology which he subsequently used to evaluate the ignition and fire resistant properties of combustible mine materials, but these efforts are more than 20 years old, and do not include many of the new materials used in mines today. The standard practice that has evolved for a combustible material to be approved by the Mine Safety and Health Administration for use in underground coal mines is to conduct extensive and costly large-scale experiments on each new material that is used in a particular mine operation. Results of such tests provide pass/fail answers to the questions of a material’s flammability or fire resistance, yet they provide little, if any, information that will lead to a better understanding of the basic mechanisms involved. Therefore, there is a need to quantify material flammability properties and to develop a methodology to rank the ignition and fire hazards of commonly used underground mine materials. During the heating and ignition stages of combustible materials, properties such as smolder susceptibility, minimum surface temperature for ignition, thermal conductivity, heat capacity, density, and heat of gasification are known to be important [Apte, 2006]. During the flaming stage, additional parameters, such as the heat of combustion, heat release rate, flame spread rate, yields of soot/smoke, and toxic gas production become more important as the flame is sustained and spreads. In very general terms, the heat flux incident on the surface of a solid combustible not only controls the rate of temperature increase of the local fuel surfaces, resulting in expulsion of fuel vapors that sustain the flaming process, but also the heating of other combustible surfaces that subsequently ignite and spread flame."

Jan 1, 2015

By V. P. Mehrotra

The effect of dissolved ions present in plant water on the flotation of phosphate ore from central Florida was investigated. Of the three most abundant ions, namely Ca++, Mg++ and SO4--, calcium ions were the most detrimental in the flotation of phosphate with sodium oleate collector. In one feed, a 10 ppm concentration of Ca++ ions reduced the BPL recovery from 86% to 46% whereas in the other a 40 ppm concentration lowered the recovery from 90% to 52%. The effect of Na2C03, NaHC03, Zeolite and EDTA to remove ca++ ions from water and hence improve apatite flotation was also studied.

Jan 1, 1986

By James A. Jessop

The Bureau of Mines is investigating the use of geophysical methods to determine rock mass characteristics, particularly rock quality, lithology, and fracture parameters that are important for determining the feasibility of in situ leaching operations. As part of this research, two Bureau-developed crosshole geophysical systems are described that can be used for site investigation and rock mass characterization. Results are given for performance tests at field sites in Arizona and Texas and for correlations between the crosshole seismic data with fracture frequency, rock quality, and lithologic variation. Recommendations are made for using elastic waveform propagation characteristics and dynamic elastic properties for improving rock mass characterization in the geomechanical assessment of in situ leaching deposits.

Jan 1, 1990

By S. Tuylu, D. Adiguzel, A. Bascetin, O. Ozdemir, U. Cinar

"In recent years, use of mineral processing tailings as a paste backfill has been significantly increased due to the environmental problems occurred during the disposal of the tailings in mine area. The paste material used to fill the stopes excavated is generally consisted of tailings (solid ratio: 70–85%) and cement (3-7%w/w). Meanwhile, the strength of the paste material is completely related to ratios of these materials. Cemented Paste Backfill (CPB) system is used as support depending on the roof pressure and also as an underground disposal method. Therefore, temporary strength value of the paste material gains a great importance. For example, the paste backfill material must compensate the minimum 4 MPa strength value in order to procure the permanent fortification, and the mixture cost must be minimal as well. For this reason, in this study, the Pb-Zn tailings were used to prepare for the paste backfill material using cement in order to perform the uniaxial compressive strength (UCS) tests. Additionally, a statistical experimental design and linear programming techniques was used to find the optimum mixture in terms of strength and cost. The pulp solid ratios (PSR) of the tailings and cement (by the solid) were varied as 75-85% and 3-8%, respectively. The results of this study; the optimum mixture is ~79.5%PSR and 6.8% cement ratio (by the solid) for 4 MPa UCS (28 day) and 16-17 cm slump value determined by this method with statistically reviewed experimental data. INTRODUCTION Fine-grained tailing waste materials resulting from mining activities are an important environmental problem. In terms of environmental protection, the use of tailings with paste backfill is a very attractive approach and a desirable solution to tailings management. Paste backfill, which is used to fill underground openings, is made by primarly mine tailings mixed with hydraulic binders and water. Its main benefits are not only disposal of mine tailings in underground, but also providing a certain strength to ensure a safe working environment. Therefore, it is necessary to use hydraulic binders due to durability of the system (Benzaazoua et al., 2004). Paste backfill, formed from the mix of dewatered plant tailings (75-85 wt. % solids), water and binder (3-8 wt. %) (Brackebusch, 1994; Landriault et al., 2005).Water is used to determine the chemical reactions between tailing and binder materials for strength as well as increasing pumpability. Portland cement and ash can be used as binder materials (Douglas and Malhotra, 1987). The factors which affect strength and stability of paste backfill separate into two groups called interior and exterior factors. Interior and exterior factors determine the strength, stability and deformation features of backfill in short and long terms and also affect the transportation of backfill to underground (Benzaazoua et al., 2004; Erçikdi et al., 2006; Yilmaz et al., 2010; Nasir and Fall, 2010). Material properties of tailings as well as site operations to fill the underground openings are the most important steps for the design purposes. Grain size distribution, chemical, and physical properties of tailings, and mechanical properties of paste fill material are used to understand the suitability of tailings as paste fill material (Ozturk et al., 2014). In addition, Fig. 1. is also proposed to apply a typical design for the operational stages of paste fill applications on the site."

Jan 1, 2016

By R. A. Haney

The Mine Safety and Health Administration (MSHA) has conducted studies both in coal and noncoal mines to evaluate methods for sampling and separating diesel particulates from mineral dust aerosols. The agency is investigating size separation techniques to distinguish between diesel particulate and respirable coal mine dust. Additionally, ashing techniques have been used to distinguish diesel particulate from mineral dust in metal and nonmetal mines. As part of these studies, personal exposures to diesel particulate for various mining occupations have been measured. This paper describes the measurement techniques and summarizes the exposure levels for various underground mining operations.

Jan 1, 1990

By William T. Ingram

Early in the 60&apos;s the author served as Chairman of the Engineering Foundation&apos;s Coordinating Committee on Air Pollution Research. One of our objectives was the development of a study on the measurement of stack emissions. Three publications of the Foundation resulted from these efforts to focus attention on the stack sampling problem.1,2,3 Two of these were annotated bibliographies covering 197 references. The third reported the results of a workshop meeting at which some 21 individuals from research institutes, industry and government discussed the possible and probable areas of research that would add to or increase the ease, precision and utility of sampling gaseous and particulate matter emitted from stacks. The American Society of Mechanical Engineers has had in its Power Test Codes a Standard Code for Determining the Dust Concentration in a Gas stream4 since 1941, The American Industrial Hygiene Association has published, in its Air Pollution Manual,5 a Section on Stack Sampling. Industry and regulatory agency requirements for stack sampling have utilized some modification of the alternatives in methodology. There has been ho single guide to stack sampling that met the needs of those who have the need to know.

Jan 1, 1970

By D. W. McGlashan

Zeta potential changes were determined as a function of pH, metal ion, and surfactant concentrations. Prom the analysis of data, flotation experiments were carried out to correlate float-ability with the differences in zeta potential. Cassiterite in water bears a negative zeta potential, The zeta potential of cassiterite is sensitive to pH changes; the electric double layer arises from the preferential adsorption of H+ and OH- ions and subsequent ionization of the hydroxylated surface. It is suspected that hydronium-ion concentration may cause reversal in the sign of the zeta potential. The ZPC was found to occur at pH 3.4. Metallic ions and the hydrated ions thereof affect the zeta potential of cassiterite in differing degrees of intensity. Without prior activation, sulfonated surfactants do not change the zeta potential of cassiterite. Sodium alkyl aryl sulfonate is especially active on lead-activated cassiterite and ferric-activated cassiteriten Microcalorimetric fiests confimed the activity of ions on cassiterite and showed the hydrated character of cassiterite smfaces. Close correlation between zeta potential and microflotation results was achieved.

Jan 1, 1969

By J. A. Pigott

Undercut-and-fill mining is a method developed by the International Nickel Company to deal with abnormal pillar mining conditions at one of its mines. Results have been such that its use has been extended to normal, pillar recovery operations in other mines of the Company as a replacement for conventional square-set pillar mining. It is anticipated that the method will also largely replace square-set stoping at INCo mines, especially for mining at deep horizons.

Jan 1, 1962

By Terrence J. Stobbe

On the job injuries to the musculoskeletal system are probably the oldest of the industrial safety problems. Long before we had pinchpoints to guard, chemical vapors to ventilate, and noise to enclose, people were getting strains, sprains, and sore backs from their work activities. In spite of the age of the problem, it remains a long way from being solved. The past ten years have seen a significant commitment by the sci¬entific community to understanding the problem. Engineers, scientists, doctors, and epidemi¬ologists have been involved, and as a result, today we have a much better understanding of the problem. The first step in understanding a problem such as this is to explore its history and epidemiology. The research done by Chaffin (1977), Snook (1970), Ayoub (1978), the ILO (1962), and others, led NIOSH to publish its Work Practices Guide for Manual Lifting in 1981. The guide was based on research carried out in general industry, but subsequent research by Stobbe (1983) and Peters (1983) has suggested that the nature of the problems producing injuries in underground mining is the same as in general industry. The problems are more intense in the mining environment.

Jan 1, 1984

By Edward J. Karkoska

The film you see is a slow motion view of a typical half million ton taconite blast at one of Erie&apos;s mines. This blast is made up of 220 drill holes and a total of 325 thousand pounds of blasting agent. The second view of the same shot is taken by a remote control hi speed camera located 1500&apos; directly in front of the blast. Shortly you will see some violent camera movement as the air blast shock wave strikes the camera location. As the blasting sequence progresses the rock fly becomes evident, and if you will look closely amidst the clouds of smoke, rock fragments weighing up to 1/2 ton begin to cascade into view. These missles are the chief concern of blast security. The next camera shot at 4000 frames/second vividly describes the primacord firing lines detonating the blast. Travelling at 22,000 ft./sec., this white snake of fire travels to each hole, detonating the blasting agent. Before we get into the matter of blast security, let me locate our operation for those of you not acquainted with Minnesota&apos;s iron ranges and tell you something about our company.

Jan 1, 1969

By M. F. Vancas

The San Manuel oxide ore reserve ranges in depth from being exposed on the surface to lying 1000 ft. beneath a conglomerate layer all of which is confined in the subsidence zone of an active sulfide block caving operation. Present day recovery of the oxide copper values is either by open pit mining-dump leaching, or in-situ leaching of the oxide ore using dilute sulfuric acid solutions and solvent extraction/electrowinning processing of the pregnant leach solutions. Future plans are for continuation of leaching with expansion of reserves and increased processing rates. The state of the mining industry and copper in particular has not been good for a number of years. All of the producers that are still operating have made major adjustments in order to remain competitive. Magma Copper Co. is no exception.

Jan 1, 1987