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By B. McFadzean, J. Taguta

The enthalpy of immersion is the heat change arising from the replacement of the solid-gas interface with the solid-liquid interface when a solid surface is immersed in a liquid. Although immersion calorimetry has been established as a valuable and reliable means of determining solid surface properties such as wettability, it has seldom been applied in mineral processing research. In this study, precision solution calorimetry was employed to measure the enthalpies of immersion of different minerals in water. The enthalpies of immersion were then related to the first-order flotation rate constant as determined by the microflotation technique. The same measurements were also made on sulphide minerals whose surfaces were modified to different degrees of hydrophobicity using different percentage coverages of amyl xanthate collector. The enthalpy of immersion technique was capable of distinguishing and ranking minerals according to their wettability. All mineral types and collector-treated minerals conformed to the same relationship that was found between the enthalpy of immersion and the flotation rate constant. This conformed, empirically, to a power law equation. As the enthalpy of immersion decreased, there were large increases in flotation response for very small changes in hydrophobicity. In addition, a critical enthalpy of immersion (CEI) was found, above which no flotation could occur. The scatter in the data points could be corrected for by adjusting the rate constant using the density of the minerals to produce a robust relationship between the rate constant and enthalpy of immersion. These results have shown that the enthalpy of immersion is able to indicate both the natural mineral hydrophobicity and the extent to which collectors can render a mineral hydrophobic. This technique has considerable potential to be used as a sensitive and accurate measure of mineral surface wettability in order to predict flotation performance.

Jan 1, 2018

By Frédéric Bélanger, George P. Demopoulos, Pascal Coursol, Thomas Feldmann

"The use of scandium in industrial applications is expected to rise in the coming years, in large part due to the expected adoption of new, clean energy technologies such as solid oxide fuel cells or Sccontaining high strength aluminium alloys. Given that the supply of Sc is currently dominated by only a few countries (Russia and China) other sources for scandium extraction need to be investigated. This paper looks at different options for the recovery of scandium from waste solutions generated during the production of TiO2 pigments via the chloride process route. In particular, a direct solvent extraction (SX) approach and possible options for concentrating scandium before the SX step were investigated. It was found that the presence of Ti during solvent extraction causes severe formation of stable emulsions, which can complicate the successful application. This was particularly a problem after multiple contacts of the organic phase with fresh aqueous phase. On the other hand it was noticed that the presence of certain elements in the solution, e.g., iron, could possibly be utilized to concentrate scandium in a solid precipitate.INTRODUCTIONMany newly developed high-tech products depend on the use of chemical elements that are only produced in very small quantities. This can result in high and fluctuating prices, as can be seen from Figure 1, which may cause substitution with inferior alternatives, if possible at all (Reck & Rotter, 2012; US Dept. of Defense, 2013; Whittaker, 2011). One example is scandium (Sc) a member of the extended rare earth element family. There are only a few scandium ore deposits, but most Sc is produced from the processing of by-products of other metallurgical processes (Melcher & Wilken, 2013). Increasing demand for scandium metal and oxide (Sc2O3, scandia) is coming from its increased use in applications such as ceramic lasers (Li, Ikegami, & Mori, 2005; Rainer et al., 1982), high strength aluminium alloys (Lathabai & Lloyd, 2002) as well as solid oxide fuel cells (SOFC) (Thijssen, 2011). Especially, the spread of climate friendly SOFC’s, which are a commercial product today (Bloom Energy, 2014), is expected to be a main driver for future Sc2O3 demand. However, the rate of market penetration of this green technology depends on the availability of high purity Sc2O3 (>99.999%), since it specifically contributes to longer life-time and better safety of the SOFC’s (Bloom Energy, 2014; Irvine, Politova, & Kruth, 2005)."

Jan 1, 2016

By Pinnaduwa H. S. W. Kulatilake, Biao Shu

The intact rock properties and discontinuity properties for both Devonian Rodeo Creek (DRC) and Devonian Papovich (DP) rock formations that exist in the selected open-pit mine were determined from tests conducted on rock samples. Special survey equipment, which has a total station, laser scanner and a camera, was used to perform remote fracture mapping in the research area. From remote fracture mapping data, the fracture orientation, spacing and density were calculated using more refined procedures compared to what exist in the literature. Discontinuity orientation distributions obtained through remote fracture mapping agreed very well with the results of manual fracture mapping conducted by the mining company. The GSI rock quality system and Hoek-Brown failure criteria were used to estimate the rock mass properties combining the fracture mapping results with laboratory test results of intact rock samples. Fault properties and the DRC-DP contact properties were estimated based on the laboratory discontinuity test results. A geological model was built in a 3DEC model including all the major faults, DRC-DP contact and two stages of rock excavation. The built major discontinuity system of 44 faults in 3DEC with their real orientations, locations and three-dimensional extensions were validated successfully using the fault geometry data provided by the mining company using seven cross-sections. Numerical modeling was conducted to study the effect of boundary conditions and lateral stress ratio on the stability of the considered rock slope. For the considered section of the rock slope, the displacements obtained through stress boundary conditions seemed more realistic than those obtained through zero velocity boundary conditions (on all four lateral faces). Stable deformation distributions were obtained for k0 in the range of 0.4 to 0.7. Because the studied rock mass is quite stable, it seems that an appropriate range for k0 for this rock mass is between 0.4 and 0.7. The displacements occurred between July 2011 and July 2012 due to the nearby rock mass excavation that took place during the same period; they were compared between the field monitoring results available from the mining company and the predicted numerical modeling results and a good agreement was obtained. Overall, the successful simulation of the rock excavation during a certain time period indicated the possibility of using the procedure developed in this study to investigate rock slope stability with respect to expected future rock excavations in mine planning.

Jan 1, 2015

By John C. Edwards

Standard smoldering and flaming combustion tests using smaIl coal samples have been developed by the U.S. Bureau of Mines as a method to evaluate the response of a smoke detector. The tests are con-ducted using a standard smoke box designed and constructed according to Underwriters Laboratories. The tests provide a standard, easily reproducible smoke characteristic for smoldering and flaming coal combustion, based upon a comparison of the smoke optical density and the response of a standard ionization chamber to the smoke. With these standard tests, the range of threshold limits for the response of a smoke detector and the detector's reliability can be evaluated for nearly identical smoke visibility and smoke physical characteristics. The detector's threshold response limits and reliability need to be well defined prior to the instrument's use as part of a mine fire warning system for improved mine safety.

Jan 1, 2010

By M. A. Schellhardt

In recent years development of instruments and equipment for determining subsurface pressures and temperatures in. gas and oil wells has made it feasible to obtain information that eras not available in the past on many gas-production problems. The material presented in this report illustrates particularly the value of subsurface-temperature data in the study of subsurface leakage from gas wells. An abnormal decline in the shut-in pressure of the Lone Star Gasoline Co. A. Jones No. 1 gas well in the southern part of the Buffalo (Tex.) field was observed during an interval when a negligible volume of gas bras being withdrawn from the reservoir. Wellhead-pressure readings indicated that the shut-in pressure of the well declined from 2,295 pounds per square inch gage on February 3, 1937, to. 1,445 on February 12, 1935. Tests conducted by engineers of the Lone Star Gas Co. suggested that the subnormal pressure gaged on February 12, 1938, probably was not caused by a head of liquid in the well.

Jan 1, 1940

By H. Hamza

Efficient and effective management of extraction process tailings is an essential ingredient for the continued success and development of Alberta's vast oil sands resources. Open-pit oil sands operations north of Fort McMurray recover bitumen from flotation circuits. Process tailings are managed in carefully engineered and monitored external tailings management facilities and in below-grade mine pits. In order to support reduction of production costs and improve upon established tailings management practices, CANMET has been working with the oil sands industry in a collaborative framework to assess the potential to integrate thickened tailings into oil sands extraction process flowsheets. Albian Sands has commercialized the process while others, including Syncrude Canada Ltd., continue to evaluate the suitability of the process for their operations. This paper provides an update on R&D work undertaken in recent years to assess and develop the potential of thickened tailings technology for the open-pit oil sands operators of northern Alberta. The aim of this work was to explore opportunities to optimize process efficiency and environmental sustainability by improving water and energy conservation and develop more alternatives for tailings management and reclamation.

Jan 1, 2004

By C. J. Toberman

Improvements in fine (28x0) and ultrafine (200x0) coal cleaning is an acknowledged goal with potential benefits in both steam and metallurgical coal markets. This work focuses on adaptation to fine coal flotation of FLOTE, a computer simulation model of the froth flotation process originally developed in the hard minerals field. FLOTE describes the flotation of solids based on particle surface activity, composition, and size. The quasi steady state model was applied to predict flotation performance at an Eastern Kentucky, full scale metallurgical coal preparation plant. Parameter estimation and model testing used detailed flotation test data from a five stage rougher flotation bank. Overall grade/recovery predictions compared well with known plant performance under variation of principally feed grade, plant throughput and reagent addition; however, grade/ recovery predictions for each cell weaken under overloading conditions. Analysis of grade/recovery over a range of sizes is presented. Suggested model improvements and extensions are discussed.

Jan 1, 1984

By S. K. Kawatra, H. J. Haselhuhn

"Fine-grained hematite ore can be concentrated by the process of selective flocculation and dispersion, which relies on proper reagent selection and control of water chemistry. While many previous studies have been performed analyzing the effects of different reagents on this process in a laboratory environment, this paper focuses on investigating the water chemistry within the process in a pilot-scale continuous deslime thickener.The pH, sodium concentration, calcium concentration and magnesium concentration were varied to determine their effects on the iron concentrate grade and recovery, and the phosphorus concentrate grade and rejection in the pilotscale selective deslime thickener. The ideal pH for the iron grade and recovery of the process using a starch selective flocculant was found to be 10.5. Phosphorus rejection, however, was increased at lower pH values. Minimization of sodium concentration was shown to improve iron grade, iron recovery and phosphorus rejection. Calcium acted as a nonselective flocculant showing higher iron recovery, lower iron grade and lower phosphorus rejection with increasing concentration. Conclusions could not be drawn from the experiments that varied magnesium concentration.The zeta potential of the solid-liquid interface of particles in each sample taken was also analyzed to show relationships between zeta potential and process performance. In all cases, a maximization of the magnitude of zeta potential correlated with increased iron grade and recovery. This supports the hypothesis that a higher level of dispersion enhances the selective flocculation and separation process.IntroductionThe United States produced 53.2 million metric tons out of the 3,000 million metric tons of concentrated iron ore produced worldwide in 2012 (Tuck, 2013). With this in mind, there has been a large increase in iron ore processing research (Bolen, 2014; Carlson and Kawatra, 2008, 2013; Halt and Kawatra, 2014; Halt et al., 2014; Haselhuhn et al., 2012a,b, 2013a,b; Liu et al., 2014; Manouchehri, 2014; Sandvik and Larsen, 2014; Semberg et al., 2014). Of the 53.2 million metric tons produced, roughly 8 million metric tons of hematite (Fe2O3) concentrate were produced using a process known as selective flocculation and dispersion (Cliffs Natural Resources, 2011). This process is the only economically viable method for concentrating low grade (< 40% Fe) fine grain (< 25µm liberation size) hematite ores. It is scarcely used due to difficulty in controlling the surface chemistry of the ore and high reagent costs (Department of Energy, 2001; Haselhuhn et al., 2013a). The process is highly dependent upon both reagent selection and precise control of the water chemistry during separation. This study shows the dependence of the process performance of a pilot-scale deslime thickener on the water chemistry in the feed slurry."

Jan 1, 2015

By Reinhard Bott

The Companhia Mineira de Metais (Brazil) runs a Zn-recovery plant in Tres Marias, federal state Minas Gerais, Brazil. The Zn-production is about 65,000 t/a, This corresponds to a concentrate throughput of about 255,000 t/a with a Zncontent of about 29.5 % at an output of 85 %, The plant is run continuously 360 d/a and 24 h/d, respectively. From this it can be concluded that: Input: approx. 650 - 700 t/d ore-concentrate and approx. 29 t/h ore-concentrate, respectively.

Jan 1, 1990

By N. J. Sather

A detailed description is given of Bunker Hill&apos;s concentration process employed at the company&apos;s lead-zinc property in the Coeur d&apos;Alene district, Idaho. The plant is equipped to process 3000 tpd of Pb-Zn ore. The history of the Bunker Hill mine dates back to Aug. 26, 1885, when Noah S. Kellogg found the outcrop of the Bunker Hill orebody on the hillside of Milo Gulch above the present town of Wardner, Idaho. A small concentrator was erected in 1886, and before the railroad came, the mine output was transported by mule teams to the head of navigation on Coeur d&apos;Alene Lake. After ten years of production, an interest was purchased in the Tacoma smelter at Tacoma, Wash., to which the entire output was shipped. In 1916 and 1917, a modern lead smelting plant and refinery was erected, and a modern electrolytic zinc reduction plant was built in 1927-1928 by the Sullivan Mining Co. which was 50 pct owned (and now totally owned) by the Bunker Hill Co. The first concentrator, a 100-tpd capacity plant, was built in 1886 near the Reed Tunnel in Wardner. The treatment consisted simply of crushing ore to 2 in. and then to 1/2 in. through rolls in closed circuit with a bucket elevator and trommel screen. The undersize was further sized in trommels to 10 mm and 3 mm. Oversize from the latter was treated in Hartz jigs. The 3 mm undersize was classified in crude hydraulic classifiers furnishing sands to sand jigs, the overflow being dewatered in a long V tank which supplied sands to a Cornish buddle and to Frue vanners. The concentrates were of excellent grade, old records showing assays of 69 pct lead and 29 oz. of silver per ton. Tailings were high inasmuch as the provision for treating the slimes was inadequate. Since the erection of the first mill, the Bunker Hill plant has gone through many changes and improvements. In succeeding years, the flowsheet was enlarged somewhat with Huntington and Chilean mills, vanners, tables, ball mills, and flotation cells being progressively added. In the transition period to the present day practice, the object was not only to improve the equipment for better mechanical performance, but to add new equipment whenever practicable so as to improve both grade and recovery of the product at a lower cost. For the modern plant and flow-sheet, see Figs. 1 and 2. The history of flotation at Bunker Hill dates back to 1913 when experimental work was done in a cell in which the pulp was circulated through a centrifugal pump set outside the cell and discharged against a baffle in the cell. The cell was built with a spitzkasten, and the quiet zone in the latter tended to let the froth break down before it got to the overflow weir. To remedy this situation, a belt drag with light metal flights was arranged for drawing off the froth over a ramp in each cell. Recovery of lead was not as high as present day practice, but was so much better than the vanners that the cells were multiplied gradually to eight cell units. The mill men called them mud hens. Flow between the cells was controlled by a weir, which determined the pulp level in each cell. These early machines were called Bunker Hill machines. The tailings were later treated in pneumatic scavenger flotation machines in a series, producing concentrates for retreatment in the Bunker Hill cleaners as well as the final tailings. This system of flotation served with minor changes until the Fahrenwald machine was developed at Bunker Hill during the period from 1921 to 1925. In 1925, the Fahrenwald flotation machine gradually replaced the Bunker Hill machine, although the Hearing pneumatic scavengers for retreatment of the tailings from the Fahrenwald machines were retained in the lead section until 1928. The use of the Hearing pneumatic machine for cleaning the rougher concentrates was begun in 1924. In 1941, plans were made for remodeling and modernizing the concentrator, the goal being a modern concentrator employing the latest equipment and milling practices. The old wooden building was to be replaced by a steel and concrete structure. Shortly after preliminary work got under way, the United States entered World War II and the demands for metals increased. To meet this requirement, the management decided to increase the capacity of the concentrator to 1800 tpd. This was accomplished in record time by the construction and operation of a sink-float plant. Installation of this plant was the first step in the general plan for modernization. In 1947 the concentrator (including the sink-float plant) was enlarged to handle 3000 tpd. Milling operations today are at the rate of 2300 tpd, five days per week. During 1959, 453,000 tons of ore was beneficiated. The Bunker Hill concentrator is in three sections: the crushing plant, a 3000-ton capacity storage bin, and the concentrator building. The crushing plant is of wood design with a gunnite coating on the exterior.

Jan 1, 1961

"Hematite pellets were electro-deoxidized to iron in molten sodium hydroxide to produce iron containing about 10-wt % oxygen. The anodic reaction was the evolution of oxygen on an inert nickel anode. Before the electrochemical reaction, there is a reaction between sodium hydroxide and ferric oxide to form sodium ferrite and, in order to compensate for the reduction in sodium oxide content, 1 wt % sodium oxide is added to the melt. In a laboratory cell, the cell voltage was 1.7 V with a current density of 5000 A m-2, a current efficiency of 90% and an energy consumption of 2.8 kWh kg-1.Introduction The iron and steel industry contributes about 2.3 bn tonnes/annum of the carbon dioxide to the atmosphere, which is about 10% of the total, and some believe that, in order to prevent a catastrophic rise in the earth’s temperature, carbon dioxide emissions must be reduced by 50% come 2050. However, by 2050, it is likely that the iron and steel industry would have grown significantly so that the production may have risen to 2 or 3 bn tonnes/year so in order to meet the 50 % reduction on today’s figure, the carbon dioxide may have to be reduced to 25% or less, which is a very demanding requirement. There are probably three ways of reducing the carbon dioxide output:1. By capture and sequestration, but this may be technically very demanding.2. Reduction using hydrogen in place of carbon."

Jan 1, 2009

By J. Kretschmann, N. M. Smith, J. F. Brune, P. Goerke-Mallet

Coal production and coal use for electric power generation are diminishing worldwide. Germany will phase out lignite mining as well as lignite and hard coal-powered electricity generation by 2038. In the United States, coal production has dropped by more than 50% since 2006, including in key coal production states like West Virginia and Wyoming (EIA 2021). Closing mines and power plants across entire mining regions has significant socio-economic impacts. Existing industries must be focused on the post-mining economy by attracting new industries, and implementing job retraining programs for displaced workers, as former mining regions transition towards agricultural, industrial and recreational use. This paper examines the German “exit from coal” under the 17 UN Sustainability Goals and illustrates the complex procedures, interrelationships, risks, and rewards of the transition towards renewable energy sources, including wind, solar, hydrogen, and bioresources.

Mar 2, 2022

By C Bellia, A Maldonado, S Strong, H Baxter

This paper discusses the technical and cost benefits of alternative data collection techniques used to collect lithological information, obtain laboratory testing samples and perform in situ testing. A comparison between diamond coring and sonic coring techniques is made using data from a recent geotechnical investigation for a final pit slope cut-back. The impacts of the coring method on laboratory strength testing results, material character testing and, ultimately, the mass strength parameters and subsequent slope design adopted are compared. The sonic coring method was chosen out of necessity and enabled the collection of tube samples of a clay unit of interest as well as bulk samples for material character and geochemical testing. The sonic coring generated unexpected degradation of clastic units, while cohesive clay-dominated units were recovered well. The likelihood that disturbance due to the coring method would result in lower than expected triaxial strength results was recognised ahead of time, with surplus samples collected and single-stage triaxial strength testing undertaken. This testing compared favourably with historic three-stage testing, where sample degradation occurred during the later stages. Material character testing of fine-grained, clay-dominated units was comparable with historic diamond core testing, while grain size estimates for more clastic units tended to be finer than previous results. The use of sonic coring over non-cored methods enabled the collection of material texture data, geotechnical properties and a full suite of geochemical analyses. This provided a comprehensive understanding of the lithological and geotechnical character of the rock and soil mass. A high-confidence geotechnical model was developed. The end result was the development of a slope design that was more aggressive than was previously designed (5&#176; steeper inter-ramp angle), which has proven to be appropriate to the conditions following excavation. Overall, sonic coring was a cost-effective means of obtaining additional geotechnical data for the site.CITATION:Baxter, H, Bellia, C, Maldonado, A and Strong, S, 2015. Case study assessing alternative data collection techniques and analysis methods to assist with geotechnical slope design, in Proceedings Iron Ore 2015, pp 419–428 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By J. N. Brooke

"This paper deals with the training of the mining and mineral engineer and the role that these professional people must play in their company environment, particularly in relation to management.THERE IS SOME CONCERN in the mining industry that the numbers of engineering graduates attracted to it and the academic training that these people have received is inappropriate. Evidence for this is:1. -The meetings and studies initiated by the various mining engineering societies to discuss:(a) attracting students and graduates to the industry; and (b) the appropriateness of the present academic training for the needs of the industry."

Jan 1, 1968

By A. A. Jonke

The combustion of fossil fuels in a fluidized bed of calcined limestone particles is a potentially efficient and economically attractive process for the generation of steam for electric power production and other uses. The process results in greatly reduced emissions of both sulfur oxide and nitrogen oxide pollutants. Development work, evaluations and design studies are being carried out by various organizations under contract with the U. S. Environmental Protection Agency and the Dept. of Interior (Office of Coal Research). The AEC&apos;s Argonne National Laboratory is one of the principal participants in the development program under interagency agreements with EPA and OCR. This paper covers the principal reasons for developing fluidized-bed combustion, its advantages and disadvantages over conventional combustion methods, and some of the potential applications of the process. The participants in the multi-organizational program, the sponsoring agencies, the general status of development of the overall project, and the chief program areas that still require resolution are discussed. The main features of the program of work at Argonne National Laboratory and the results achieved so far are also included.

Jan 1, 1974

By C. J. Wilhelm

The Carthage field, in Pamela County, Tex., is one of the largest gas-condensate in respect to both areal extent and volume of recoverable gas in the United States. The field (shown in fig. 1) is approximately 25 miles long by 20 miles wide on a southwest to northeast axis and includes 306 producing wells as of July 1, 1948. 3 distinct reservoirs in the Rodessa and Sligo formations have been found in the Carthage area the two main producing zones being the Upper and Lower Pettit, less important production being found in the Hill zone. Although a small amount of commercial production has been obtained in the Travis Peak formation, it has not been included in this report. An engineering study has been made by Bureau of Mines engineers, with the ultimate objective of arriving at an estimate of the gas reserves, by application of the data made available by the operators in the Carthage field. During World War II the Bureau of Mines submitted a restricted petroleum-engineering report on the Carthage field to the Petroleum Administration for War, at which time (March 1944) the development consisted of only 19 producing wells, 7 of which were dual completions. Because of the restricted nature of the first report and the limited development at the time of writing, a more complete investigation and subsequent publication was considered advisable. Bureau of Mines engineers have made use of all available records such as electric logs, drilling-time logs, core descriptions and results of core analyses, well-completion records, and production data in developing this report.

Jan 1, 1950

By J. L. Hoffman, F. W. Wood, R. A. Beall, W. E. Anable

For melting and purifying reactive and refractory metals , two electronbeam furnaces have been constructed and operated on the principle of highpowered vacuum diodes . It was necessary to design the vacuum and electrical systems , mechanical features , and operating procedures to cope with dynamic conditions . Vacuums of 10-5 torr , tens of kilowatts of power at 5 to 10 kilovolts , and the evolution of impurities at thousands of liters per minute constitute a challenging combination . The furnaces have been used to melt either button specimens weighing several hundred grams or , by feeding material as a consumable electrode , small ingots weighing several kilograms . Hafnium is among the metals that can be purified . For example , hafnium starting material with Rhn 94B was softened to Rhn 70B after 2 hours of melting under a 25 - kw electron beam , and impurity levels were reduced as follows : Oxygen , from 700 to 350 ppm ; nitrogen , from 71 to 20 ppm ; hydrogen , from 6 to 1 ppm ; aluminum from 75 to < 20 ppm ; chromium , from 80 to < 10 ppm ; iron , from 100 to 40 ppm ; and silicon , from 40 to < 20 ppm .

Jan 1, 1963

By Robert Martin, Brian W. Stump, David P. Anderson, W Scott Philips

Since 1994, various mines in the US have cooperated with research scientists at the Los Alamos and Lawrence Livermore National Laboratories to address issues related to verification of the Comprehensive Test Ban Treaty (CTBT). The .CTBT requires that no country may conduct any nuclear explosion in the future. While the CTBT is a significant step toward reducing the global nuclear danger, verifying compliance with the treaty requires that the monitoring system be able to detect, locate and identify much larger numbers of smaller amplitude seismic events than had been required previously. Large mining blasts conducted world-wide will be of sufficient amplitude to trigger the monitoring system at the lower threshold. It is therefore imperative that research into the range various blasting practices employed, the relationship of yield to seismic magnitude, and identification of anomalous blasting results be performed. This paper will describe a suite of experiments funded by the Department of Energy and conducted by the Los Alamos and Lawrence Livermore National Laboratories in cooperation with the US mining industry. Observations of cast blasting, underground long wall generated coal bumps, stoping, and explosively induced collapse of room and pillar panels will be presented. Results of these dual use experiments which are of interest to the mining community will be discussed. These include 1) variation of amplitude of seismic energy at various azimuths from cast blasts, 2) identification of the extent of back failure following explosive removal of pillars, and 3) the use of single fired shots for calibration of the monitoring system. The wealth of information and discovery described in this paper is a direct result of mutual cooperation between the US Government and the US Mining Industry.

Jan 1, 1997

By Norman Paley

A blasting project was undertaken at the Teck Alaska Red Dog Mine to determine the extent that Semi-Autogenous Grinding (SAG) mill throughput could be increased by using electronic detonators. The Julius Kruttschnitt Mineral Research Centre (JKMRC) blast fragmentation, jaw crusher, and SAG mill breakage models were used to relate changes in the blast fragmentation measured with the SPLIT image analysis system to the SAG mill performance. For optimum performance a SAG mill requires a particular size distribution; a number of large rocks to facilitate breakage by impact, as few as possible middle size rocks that become critical size, and the maximum fines which will pass quickly through an open-circuit mill. Physical constraints of drill and charging equipment size limit the extent that pattern geometry can be changed, and the powder factor increased, to maximize fines. However, a growing number of studies have shown that the use of electronic detonators can lead to increased fragmentation and greater muckpile uniformity, providing an alternate means of increasing primary breakage by blasting. Unfortunately results in the literature rarely quantified the changes to the fragmentation distribution. The measured gyratory crusher fragmentation of 2,750,000t of ore was used to compare stockpiles constructed from shots using pyrotechnic detonators to those using Orica i-kon electronic detonators. In the first stage of the study, stockpiles shot using 25 ms inter-hole timing with pyrotechnic detonators were compared to stockpiles shot using the same inter-hole timing with electronic detonators. In the second stage of the study, stockpiles shot using 17 ms inter-hole timing with pyrotechnic detonators were compared to stockpiles shot using 8 ms inter-hole timing with electronic detonators. When electronic detonators were used with the inter-hole timing kept constant, fragmentation uniformity was slightly increased, but mean fragment size increased by 20%. When electronic detonators were used with faster inter-hole timing, fragmentation uniformity was slightly reduced, but mean fragment size decreased by 30%. Modeling was used to determine the SAG mill throughput which could be expected from this change in fragmentation to determine the pairs of pyrotechnic and electronic blast pattern geometries that would provide similar throughput. The economic benefit from the electronic detonator shots was calculated to assist in determining the optimum blast designs for their use.

Jan 1, 2010

By E. I. Gedgagov, A. V. Tarasov, V. M. Paretsky, A. D. Besser

"Hydrometallurgical technology and equipment have been developed in the Gintsvetmet Institute (Moscow, Russia) for processing of rhenium-containing alloys waste to produce commercial- grade metal products with recovery of all valuable components. The technology is based on waste treatment with hydrofluoric acid with subsequent decomposition with nitric acid at pH = 1 to 1.5 to ensure maximum degree of rhenium conversion to the liquid phase of pulp. Some experimental lots of ammonium rhenate and ammonium molybdate were produced at a pilot plant. The following main techno-economic performance values have been achieved when processing grinding residues (balance-based tests were conducted with a lot of SP-2 alloy waste) to produce ammonium rhenate of high purity, as well as tantalum-niobium, nickel-cobalt and tungsten-molybdenum concentrates: rhenium recovery of 92.3% into a commercial product from grinding waste; production cost of 170-210 US Dollars per 1 kg of rhenium in commercial-grade ammonium perrhenate taking into account its recovery; tantalum and niobium recovery of 75- 77% into bulk concentrate; nickel and cobalt recovery of 88-89% into a mixed concentrate; tungsten and molybdenum recovery of 88% to 91% into concentrate. Introduction During the recent years the use of rhenium has been increasing at a fast pace worldwide [1]. The growing prices for this refractory metal are an essential factor encouraging collection and processing of rhenium-containing wastes. Double-component waste, such as molybdenum- rhenium or tungsten-rhenium alloys can be readily decomposed and the process of rhenium recovery does not cause, therefore, any difficulties. The main separation of metals takes place at the stage of waste digestion with acid, because molybdenum and especially tungsten remain in precipitate in the form of hydrated oxides and rhenium reports in the liquid phase, from which it can be easily extracted by solid ion-exchange or solvent extraction techniques. However, most of rhenium is contained in super-alloy waste [2]. Such waste has different contents of valuable components, such as tantalum, niobium, molybdenum, tungsten, cobalt, nickel, chromium and other metals. The basis of all such super-alloy waste is nickel. Refractory metals in the super- alloy composition are in the form of metallized target-oriented eutectics, and due to this reason hydrometallurgical decomposition of such alloys is a difficult technological task."

Jan 1, 2008