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By B. Jones, W. Zhang, Q. Li

Rare earth elements (REEs) are critical materials to the United States due to the imbalance between their supply and demand. Considerable research and development efforts have been devoted on REE recovery from coal refuse using inorganic acids as lixiviants, however, the cost of inorganic acids is normally high, and the environmental friendliness is low. As alternatives to inorganic acids, organic acids of small molecules such as lactic acid and succinic acid have been used as efficient and green lixiviants for metal extraction. In addition to the supply risk of REEs, food waste generation and disposal is another emerging and critical issue since nearly one-third of the edible food is lost or wasted each year. In this research, organic acids were microbially converted from food waste and used as lixiviants to leach REEs from density fractions of an Illinois Basin coal. Experimental results indicate that REE recovery increases of up to 15 absolute percentage points were obtained using organic acids relative to inorganic acids. The conversion of food waste into value-added, renewable, and environment-friendly chemicals promotes the economic viability of REE recovery from coal refuse.

Mar 2, 2022

By Roger L. Baer

Introduction Brief Review -SEC mission & history -Development of SEC mining disclosure guidelines -Principles of good public company ‘disclosure’ SEC staff views of alternative approaches to mining company disclosure -SME Resource & Reserve Committee – April 2003 Food for thought

Jan 1, 2003

By James K. Gowans

Polaris Mine, a Cominco Ltd. property located on Little Cornwallis Island, Northwest Territories, Canada, (Lat. 75°30'), is the world's most northerly base metal mine. Polaris Mine, which commenced operations in late 1981, is characterized by its unique process barge. The lead/zinc concentrator, powerhouse, shops, offices, dry and warehouse have all been incorporated on a 400 ft. long by 100 ft wide by 65 ft high barge. This process barge was constructed in Southern Canada, at Trois Rivieres, Quebec, then towed 3500 miles into its final resting place on the high arctic. This paper describes the unique design of the barge concentrator and deals with the start-up and continuing operation of the Polaris Concentrator.

Jan 1, 1984

By A. Juganda, Jr. Bogin, J. Brune, J. Grubb, S. Lolon

"Face ignitions at the longwall face are a serious hazard in underground coal operation that can lead to a major mine explosion. Despite having methane monitoring system mounted on the shearer and at various locations on the longwall face, undetected accumulations of methane can still occur and result in face ignitions. With the use of Computational Fluid Dynamics (CFD), the interaction between the air flow at the longwall face and factors that contribute to the accumulation of methane around the longwall face can be modeled and visualized in great detail. The results confirm that the tailgate corner of longwall face is a critical area that is prone to face ignitions and needs to be properly monitored. The occurrence of roof falls on the tailgate entry inby the face and/or poor caving conditions behind the shields can pose a safety risk in the longwall operation. Poor gob caving can lead to insufficient face air quantity to dilute methane at the tailgate corner, while blocking of tailgate by a roof fall can carry methane contaminated air from behind the shields back into the face near the tailgate corner and pull the explosive gas zones (EGZs) inside the gob closer to the face. INTRODUCTION Ignitions of accumulated methane gas at longwall faces are well known to be among the major causes of methane explosion in underground coal longwall mining operations. Some of these ignitions can lead to major mine explosions, such as the Upper Big Branch mine disaster in 2010, in which methane migrated from the gob to the shearer cutting drum where it was ignited, resulting in a major mine disaster (Phillips, 2012). History has shown that, despite having methane monitoring systems mounted on the shearer and various locations at the longwall face and alarms set to 1% CH4 (Mine Safety & Health Administration, 2012), undetected accumulations of explosive methane-air-mixtures can still occur and result in face ignitions. Recent study by Verma and Brune (2016) state that frictional ignitions on longwall faces contributed to 379 incidents or 23% of the total 1,637 cases of face ignitions recorded from 1983 to 2014. The common cause of mine explosions usually involves poorly designed ventilation systems, insufficient ventilation, and inadequate monitoring in critical areas that are prone to methane accumulations. In longwall operation, the tailgate corner of the longwall face can be considered as one such critical area. On-site studies by Krog et al (2006) and Thakur (2008) show that increasing the longwall panel width may also increase cumulative methane emissions at the tailgate. There are also indications that increasing longwall face lengths can result in an increase of the void space behind the face under certain roof conditions (Schatzel et al, 2006). This may cause higher leakage rates from the face to the gob, thus reducing the amount of fresh air available at the tailgate."

Jan 1, 2017

By M. Kalin, H. D. Sharma

INTRODUCTION The inactive uranium mill tailings investigated in this study are located in two mining districts, Elliot Lake and Bancroft, Ontario, Canada. The sites exhibit a mixture of surface features consisting of dry areas, with or without vegetation and areas covered with water. On the edges of the water bodies, indigenous vegetation has invaded the tailings beaches [ Typha] spp. (Tourn.) L. a dominant plant on these tailings beaches, has been studied for the uptake of radium-226 and lead-210 (Kalin and Sharma, 1981b) from the tailings. It was found that most of the radium-226 remains in the roots of the plants, and that the solubility of radium-226 in control soil differs from that in tailings. The uptake of radium226 by vegetation and other biota is related to the solubility of the element in water. Factors controlling the solubility of radium-226 in uranium mill tailings are of interest in assessing the environments effects of these wastes. Rusanova (1962) found that the soluble or extractable amount of radium from the soil is inversely related to the total concentration of calcium and magnesium in the soil. Experimental work has clarified some aspects of the leachability of radium from uranium mill tailings (Levins, et al., 1978; Wiles, 1978, and others). Halvik, et al., (1967) studied the effects of pH and chemical composition of surface water on the liberation of radium from uranium mill tailings and uranium ore. He found that an increase in pH, up to a value of pH 9, decreases the amount of radium released from the tailings and the ore. A positive effect of calcium chloride was noted on the leachability of radium. Benes (1981) reviewed the physicochemical forms of radium and its migration in water. Based on experimental work, he identifies primary factors which determine leaching of radium from uranium mill tailings. The ratio of the volume of the leaching solution to the weight of the leached sample; the composition of the leached solids and the leaching solution, and finally the pH of the leaching mixture are of importance. He emphasized that a scarcity of field data exists, which would relate experimental work to the actual situation in the tailings ponds. Inactive tailings ponds in Ontario are 16 to 23 years old, and processes which are of importance in evaluating the long term effects of uranium mill tailings in the environment can be studied. The objective of this work was to investigate the leachability of radium-226 from the tailings under field conditions. MATERIALS AND METHODS Sample Collection A description of inactive tailings sites in Ontario, where the tailings and the water for this study were collected, can be found in a report by Kalin (1981). [Typha] spp. specimens were excavated from 16 different locations along with the tailings attached to the roots of the plants. A tailings sample from the surface area, around the plant, was secured before the excavation. Tailings samples were also collected from dry areas free of vegetation in the vicinity of the wetland stand. Surface samples (depth 0-5 cm) and samples at a depth of 20-25 cm were secured. Water was collected from the shallow tailings beaches around the vegetation stands. Sample Preparation The roots of the plants were washed free of tailings with distilled water. The resulting thick slurry of tailings was allowed to settle in the wash basin for five minutes and the supernatant water was decanted. The remaining saturated tailings were brought to 400 ml volume with distilled water in a beaker. The slurry was mixed with a magnetic stirrer for two 24-hour periods. After the first 24-hours, the slurry was allowed to settle for 20 minutes and the supernatant water was removed. The remaining sediments were suspended again in distilled water (400 ml total mixture) and leached for a second 24hour period. The tailings slurries from the different locations had a solid to liquid ratio, which ranged from 0.5 to 1.4 grams of dry tailings material per millilitre. The variations in the ratios are the result of differing fractions of coarse and fine tailings on the sites. In the final leachates, the solid to liquid ratios of the samples were 0.17 (±0.1) g/ml. The pH of the surface water and the leachate was determined with an I.L. Portomatic pH meter. All the tailings samples were homogenized in a mortar and brought to dryness at 75 to 85°C. Approximately 0.5 g of tailings were

Jan 1, 1981

By F. C. Eberhardt

Systematic planning of underground work is increasingly important in this era of advancement of rapid underground excavation techniques. The day when planning consisted of deciding which men would shovel and which would push wheelbarrows - and the answer to any problem was more men - has faded into the past. The modern underground constructor faces economic and competitive conditions which demand exacting planning, logistics and scheduling; the most efficient operating methods and equipment known; and realistic, comprehensive, analytical risk evaluation with adequate contingencies or other protection. Proper planning to cover all these elements - plus one other element: old fashioned "good luck" - will insure a successful underground project. Underground work, because of its inherent characteristics, is among the most risky types of construction. The forecasting of underground conditions is at this time only educated predictions based on sometimes faulty data, insufficient data, or no data at all. The predicted conditions of numerous operations have deviated far from the actual conditions encountered. Opportunities for error exist in profusion seldom present for surface operations. Other than this unusual risk factor, the planning for underground work does not differ greatly from planning for surface work. In planning and starting an operation there is seldom available the information or time to thoroughly analyze all unknowns. It then becomes necessary to develop a plan predicated on assumptions and trade-offs that appear to be safe and economical and to add time and cost contingencies that will provide some degree of protection. On competitively bid fixed price work, choice of markup on estimated

Jan 1, 1974

By Chester A. Rowland, David M. Kjos

INTRODUCTION Mineral ore comminution is generally a feed preparation step for subsequent processing stages. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding - primary and secondary), natural or manufactured nonmetallic media (pebble milling) or manufactured metallic media - steel rods, steel or iron balls, or a combination of autogenous media and steel balls (SAG milling). This chapter covers rod and ball mills utilizing manufactured metallic grinding media. MILL DESIGN The interior surface of rod and ball mills exposed to pulp and/or grinding media are protected from wear and corrosion by rubber, metallic or a combination of rubber and metallic wear resistant materials. Rod and ball mills essentially draw constant power, thus are well suited for use of synchronous motors with power factor correction capabilities as drive motors. A net of approximately 120 to 130 percent of running torque is required to cascade the charge in these mills. The pull-in torque is about 130 to 140 percent with the pullout torque to keep the motor in-step (in-phase) generally in excess of 150 percent. When rod and ball mill are started across-the-line the starting and pull-in torques result in inrush currents exceeding 600 percent which possibly result in high voltage drops. To deliver 130 percent starting torque to the mill the motor design must take into account the maximum anticipated voltage drop. Motor torque decreases as the decimal fraction of the voltage available squared. E.g., a motor rated 160% starting torque with a 10% system voltage drop will deliver 160% x or 129.6% torque to its output shaft When it is not possible or practical to start a fully loaded synchronous motor across-the-line it is possible to utilize the motor's pull- out torque to start the mill. By using a clutch, normally an air clutch. between the motor and the mill, the motor is brought up to synchronous speed before the clutch is energized. If the motor has an adequate amount (175 or greater) of pull-out torque the pull-out torque starts the mill without major disruptions on the electrical system. Since the energy release at initial cascade of the mill charge is an inverse function of acceleration time, a minimum acceleration time of 6 to 10 seconds or more is recommended to prevent damage to the mill or the mill foundation. Economics at the time of plant design and mill purchase determine the drive to be used. The simpliest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinionshaft by either an air clutch or flexible coupling. Using a speed reducer between the motor and pinionshaft permits using motors having speeds in the range of 600 to 1000 RPM. In this speed range, if power factor correction is not required induction motors can be used; squirrel cage where there is no restriction on inrush current; slip ring where a slow start and low inrush current is required. Air clutches can also be used to ease starting problems with squirrel cage motors. In some areas of the world induction motors and starters are less expensive than synchronous motors at a sacrifice of motor efficiency and power factor correction. Dual drives, that is two pinions driving one gear mounted on the mill, become economical for ball mills drawing more than 3500 to 4000 horsepower (2600 to 3000 kilowatts). Further developments of the low frequency, low speed synchronous motors with the rotor mounted on the mill shell or an extension of one of the mill trunnions could improve the cost picture for these "gearless drives", making them practical for large ball mills. The percent of critical speed, which is the speed at which the centrifugal force is sufficiently large to cause a small particle to ad- here to the shell liners for the full revolution of the mill is given in mill specifications. Critical speed is determined from the following: Where D is mill diameter inside liners (specified in meters). Cs is critical speed in RPM. When D is specified in feet

Jan 1, 1998

Relatively lower recoveries (60%) of coarse phosphate (-16 +35 mesh) particles are attributed to the difficulty in transportation of coarse particles from the pulp to the froth column and rapid drainage of these particles from the froth column to the pulp. Increasing the hydrophobicity of coarse particles by higher, more efficient coating of the collector and using larger froth volumes have been suggested to overcome these problems. In this study, the role of emulsification of the fatty acid-fuel oil collector on the flotation recovery of coarse particles was examined. It was determined that emulsification can not only increase the coarse particle recovery, but do so at higher fractions of fuel oil in the collector, thereby reducing the reagent cost as well.

Jan 1, 1989

By N. A. Abdel-Khalek

Laboratory flotation tests were carried out on six Polish samples of different rank. Each coal sample was screened into 7 size fractions. The floatability of size -by- size for various size ranges for different samples was investigated as: well as the original samples. The results showed that in flotation of size -by -size, the intermediate size had the maximum floatability. The range of this intermediate sizes varied with the rank of the coal. Entrainment was shown to be an important mechanism to the recovery of fine particles in flotation of the original samples".

Jan 1, 1989

By G. D. Elliot, R. L. Grayson, G. L. Winn

For the years 1990 and 1991, West Virginia's underground coal small-mine fatality rate was approximately 10 times higher than the national coal-industry fatality rate. Although the number of fatalities went down from 22 in 1991 to 17 in 1992, West Virginia still led the nation in coal-mining fatalities. A team of mining and safety professionals from West Virginia, interfacing with the Mine Safety and Health Administration and the West Virginia Office of Miners' Health, Safety and Training, studied this problem. A clear path for formal intervention was determined and is presented here.

Jan 1, 1996

By Ashma Singh, Nichole Maistry

Due to declining iron ore grades, low-grade ores with complex mineralogy — namely, banded iron formation (BIF) — are being explored to maximize iron ore resources and increase the life of mine. Flotation is the standard global process used for the beneficiation of BIF ores, as it is a complex ore reserve, and liberation is typically achieved below 1 mm. This study focuses on the flotation of a South African BIF, evaluating its response to different collectors, as well as effect of grind. As flotation is an expensive process and South Africa ranks high on the production cost curve, preconcentration by gravity or magnetic separation is performed prior to flotation. Thus, flotation was also performed on a gravity concentrate to determine if a grade suitable for steelmaking (approximately 63 percent) is achievable. The results indicate that the use of an alkyl ether amine-based collector partially neutralized with acetic acid is suitable for separation of hematite from illite and quartz at high pH values. Further optimization tests are being pursued to improve overall recovery and grade.

May 1, 2025

By Jerry L. Purswell

INTRODUCTION The principal OSHA standards dealing with tunneling activities are found in 29 CFR Part 1926. A reference for the applicable subparts which apply to tunneling is included as Appendix A. In addition, some standards from 29 CFR Part 1910 have been identified as applicable to construction and are included in the recent printing of an interim vertical standard for construction (1910/1926 Project). The remainder of this presentation will be focused on past OSHA activities to revise Subpart S of 1926, the principal subpart dealing with tunnel construction.

Jan 1, 1979

By N. T. Rempe

The underground voids that result from mining provide opportunities for deep geologic waste disposal (DGWD). Such disposal could solve many environmental problems. DGWD takes careful planning, technology transfer and true partnership between waste generators and mine/repository operators. The US government is now planning on using specially mined excavations to dispose of radiotoxic and mixed waste. Using the experience gained from the domestic projects that are now in preparation and from foreign repositories that are now in operation, US mines can claim their rightful share of the growing market for DGWD.

Jan 1, 1998

By Henry P. Cole, Andrzej M. Wala

Engineers and psychologists at the University of Kentucky have developed simulation exercises that teach and test proficiency of mining technical personnel for coping with planned and accidental connections between separate air courses in underground mines. The exercises are based upon investigations of actual mine disasters. The exercises are judged as valuable and interesting training materials by experienced mining personnel. Although these specialists generally perform well on the exercises, potentially fatal errors are not uncommon. The use of such realistic exercises may better prepare miners for preventing and coping with underground mine ventilation emergencies.

Jan 1, 1987

By S. K. Kawatra, S. K. Valluri, S. J. Root

Carbon dioxide, a by-product of combustion reactions, is a greenhouse gas linked to climate change. At Michigan Technological University, we have studied the capture of CO2 using alkali scrubbing solutions in a pilot scale packed bed counter-current scrubbing column. To do this, we have simulated flue gas by combining streams of CO2 and compressed air. Real flue gas has impurities such as CO, SO2, and NOx, and has lower levels of O2 than our simulated flue gas. In order to study CO2 capture from a real flue gas, the Department of Chemical Engineering worked with the Facilities department to install a pilot scale scrubbing column in the Michigan Tech steam plant. Experiments were conducted using a sample stream of the flue gas from the boiler in the steam plant. Data collected from these experiments were compared to the data collected from identical experiments conducted on simulated flue gas in the lab. The capture efficiency of the real flue gas scrubbing column is discussed in this paper. BACKGROUND The most common equipment used to separate CO2 from a flue gas stream is a counter-current packed-bed scrubbing column. The department of Chemical Engineering at Michigan Technological University has two such pilot scale columns. In order to study the effect of impurities found in real flue gas, one is installed in a controlled laboratory using a simulated flue gas created by mixing CO2 and compressed air. This column will be used to establish a baseline for percent capture analysis and is displayed by figure 1. The second scrubbing column is identical to the first, and is installed in the Michigan Tech central energy plant. The plant produces up to 130,000 lb/hr of 80 psig steam to provide hot water and heat for campus buildings. A typical wintertime load burns around 25,000 SCFH of natural gas; this load releases 1.4 metric tons of CO2 per hour. The flue gas the steam plant produces is roughly 7.5% CO2 by volume. In order to study CO2 capture from this flue gas, a tap has been installed on the main exhaust duct which draws a sample of the flue gas into the pilot scale scrubbing column. Figure 2, a simplified diagram of the steam plant, shows that the tap was installed between the economizer and the stack condenser.

Jan 1, 2019

By Caelen D. Anderson

An investigation of a hydrometallurgical method for the recovery of tellurium dioxide from cadmium telluride (CdTe) thin-film manufacturing scrap was conducted. Sulfuric acid/ hydrogen peroxide lixiviation was employed to selectively leach cadmium and oxidize tellurium to tellurium dioxide (TeO2). Experimental variables included solution pH, pulp density, oxidant (H2O2) concentration, temperature, agitation rate and residence time. After one hour of lixiviation, no noticeable increase in grade or recovery occurred; thus, all experiments were performed for one hour. Leaching results showed it was possible to recover at least 74% of the tellurium, while producing a 98% pure TeO2 product. Additionally, an apparent activation energy value of 12.1 kJ/mole (2.9 kcal/mole) was determined. This value, along with results from agitation rate experiments, indicated the reaction was primarily controlled by diffusion through the product layer. During lixiviation, partial dissolution of tellurium proved inevitable. Thus, a second step: ammoniacal precipitation was introduced to recover the aqueous tellurium as a precipitate. A statistical experimental design (DOE) was used to identify the significant effects from the residual effects. Experimental variables included precipitation pH, temperature, initial tellurium tenor and agitation rate. Precipitation results showed that it was possible to produce a low-grade (46% Te) precipitate, while recovering 98% of the aqueous tellurium species. The precipitation pH and temperature (°C) were determined as the significant effects for both the grade and recovery of tellurium. Agitation rate and initial tellurium tenor were determined as the residual effects and did not influence the grade or recovery of tellurium. An illustration of tellurium recovery, as a function of the significant effects, is illustrated in Fig. 1.

May 1, 2012

By H. Cetin

The availability of coal resources of selected 7.5-minute quadrangles of southwest Indiana was examined in this study. Resources were defined as original coal in place, remaining after mining, and currently available considering land-use and technological restrictions. The amount of coal mined from each study area varies depending on the mining practices of the companies operating within the area. Of the three quadrangle investigations, the available resources range from 50 percent to 79 percent of the original resources. Most of the coal resources under deep-minable category were not available because of technologic restrictions, such as insufficient coal thickness.

Jan 1, 1996

By Ralph P. Hackl

Biooxidation processes for treating arsenical and other types of refractory gold ores are touted by many as being simple, inexpensive and environmentally friendly. One of the advantages claimed by proponents of biooxidation is that arsenic is disposed of as insoluble ferric arsenate which is suitable for long term disposal in tailings ponds. However, there is considerable controversy as to the stability of ferric arsenate with respect to re-dissolution of arsenic. In particular, there is concern that the amorphous ferric arsenate produced at low temperature is not stable, especially at the alkaline pH's required for cyanidation. There is a general lack of published information on the stability of arsenical bioleach tailings. Past work on ferric arsenate stability has usually involved purely chemical systems and the use of synthetically prepared precipitates. Upon completing the Salmita bioleach pilot plant campaign (Hackl and Wright, 1989) the author had the opportunity to test the stability of the arsenical bioleach/cyanide tailings which were produced. This paper summarizes the results of that work.

Jan 1, 1990

By Mark William Johns

Carbon-in-pulp (CIP) is the preferred method of gold recovery om cyanided plp. However, other gold adsorbents (namely, ton-exchange resms) are being introduced to the industry as they are considered to have several advantages over carbon for the recoery of gold. A set of test methods have been developed by the industry to evaluate the quality of different carbons. owever, many ?f these tests are not applicable to evaluating ton-exchange resms. Hence, a need has arisen for a method to evaluate the different adsorbents. This paper introduces a philosophy to evaluate an adsorbent that results in an estimation of the capital cost and the operating cost associated with the adsorbent. Tests that have been devel­ oped to evaluate the ability of an adsorbent to adsorb gold (in­solutton . counter-current) and to evaluate the optimum adsorbent for a specific pulp (in-pulp batch) are introduced. A method of converting the test data into an economic dimension is present­ed. Case studies are presented to highlight the discussion.

Jan 1, 1996

By Peter G. Mason

Pressure oxidation is now a well estab­lished process for the treatment of refracto­ry gold ores and concentrates. The process utilizes pure oxygen to oxidize mineral sulphides in an aqueous slurry at elevated temperature and pressure. Gold released from the sulphides becomes amenable to subsequent cyanidation. A major component of the operating cost of this process the supply of oxygen. The quantity required is dependent on the sul­phide grade, the extent of oxidation and the percentage utilization of the oxygen. As the sulphide grade is fixed and complete oxida­tion is generally required to maximize gold liberation, the only opportunity for saving in oxygen consumption is to improve its utilization. Under circumstances where the feed material contains substantial carbonate, oxygen utilization can be improved by strip­ping carbon dioxide from the autoclave vent and recycling the associated oxygen. Fluor Daniel Wright included an oxygen recycle circuit in the Getchell pressure oxidation plant, which is operated by First­Miss Gold in Nevada. FirstMiss Gold has applied for patent rights to the process as implemented at Getchell.

Jan 1, 1990