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By J. Keith Seidler, David R. Lentz, Douglas C. Hall, Nicholas Susak

Abstract - The Ta-Nb-Zr-REE-rich McKeel Lake alkali granite pegmatite-aplite dike swarm is hosted within the late Silurian multiphase alkali granite intrusion (Welsford) near Browns Flats in southwestern New Brunswick. A steep easterly dipping series of pegmatite-aplite dikes trending 170º can be traced 150 m along strike; they are up to 40 cm wide, closely spaced (10-20 cm apart), and interconnected by narrow branching dikelets that generally trend perpendicular to the overall strike direction. The Welsford intrusion has mantle A-type affinities characterized by enrichment of Zr, Ta, Nb, Th, and rare earth elements (REE). Two separate sets of dikes have been identified: (1) an unmineralized fine-grained porphyry, which occurs at depth and seems to be the youngest intrusive phase; and (2) a mineralized pegmatite-aplite dike swarm (419 ± 20 Ma) that crops out at surface. The mineralized pegmatite-aplite dikes likely represent a late magmatic, highly fractionated phase of the Welsford intrusion. XRF and INAA data indicate that the Welsford intrusions and related pegmatite-aplite dikes show elevated levels of Ta (74-220 ppm), Nb (750-3,040 ppm), Zr (1,450-36,000 ppm), Y (540-3,070 ppm), Th (170-870 ppm), U (60-250 ppm), Ce (430-1,950 ppm) and Yb (110-290 ppm). Electron microprobe analysis of two samples identified Ta-Nb- and REE-rich phases: euxenite, fergusonite, aeschynite, (Y)-aeschynite, and a REE carbonate.

Jan 1, 2005

By Keith A. Heasley, Frank E. Chase, Christopher Mark

Deep cover retreat mining (overburden in excess of 750 ft) is an important emerging issue which will intensify in the future as the more easily mined shallow seam reserves are depleted. Analysis of Mine Safety and Health Administration (MSHA) statistics indicates that deep cover pillar recovery accounts for a disproportionate share of the underground coal mine roof/rib fall fatalities and injuries. Past research has shown that previously recommended Analysis of Retreat Mining Pillar Stability (ARMPS) stability factors (SF’s) may be excessive for deep cover pillar design. The objectives of this study were to evaluate the various methods and strategies by which panels, production pillars and barrier pillars are developed and extracted under deep cover, and to develop appropriate design guidelines. In the course of the research, 29 mines in 7 states were investigated to collect panel design case histories. At each mine site, underground geotechnical data were collected on the pillar line in order to document roof rock, coalbed and floor conditions. The analyses indicated that squeezes were the most likely failure mode where the depth was less than 1,250 ft, but bumps predominated in the deeper cover cases. Immediate roof rock quality, the ARMPS SF’s, and barrier pillar stability factors were all found to be important parameters in determining the outcomes of the case histories. Design guidelines, including suggestions for barrier pillars to isolate active panels from nearby gobs in bump prone ground, are also proposed.

By P. L. Sibrell

The U.S. Bureau of Mines has tested cyanide leaching for recovery of palladium (Pd) from spent petroleum processing catalysts. Three different catalyst samples were supplied by a spent-catalyst processor. These catalysts consisted of a zeolite base and contained 0.4 to 0.7 pct Pd. During alkaline cyanide leaching, the catalysts exhibited ion-exchange properties due to their zeolite matrices. Hydrogen ions were released from the zeolite in exchange for sodium ions in solution, resulting in a significant decrease in solution pH. This could present a safety hazard because of the potential for release of toxic hydrogen cyanide gas. A pretreatment .step where the catalysts were contacted with a 1.0M sodium hydroxide solution was found to prevent the pH shift from occurring. Following the sodium hydroxide pretreatment, two stages of leaching at 160 °C with solution containing 1 pct sodium cyanide and 0.1M sodium hydroxide gave at least 75 and up to 95 pct Pd recovery. The Pd was quantitatively recovered from the leach solution by thermal decomposition in an autoclave at 250 °C for 1 h. The Pd content of the precipitate was over 50 pct. Thermal decomposition also decreased the total cyanide content of the barren solution to less than 0.2 mg/L. The catalyst leach residues passed the Federal Toxicity Characteristic Leaching Procedure and the California Waste Extraction Test, indicating that landfill disposal of the leach residues would be acceptable.

Jan 1, 1995

By P. L. Sibrell

The U.S. Bureau of Mines has tested cyanide leaching for recovery of palladium (Pd) from spent petroleum processing catalysts. Three different catalyst samples were supplied by a spent-catalyst processor. These catalysts consisted of a zeolite base and contained 0.4 to 0.7 pct Pd. During alkaline cyanide leaching, the catalysts exhibited ion-exchange properties due to their zeolite matrices. Hydrogen ions were released from the zeolite in exchange for sodium ions in solution, resulting in a significant decrease in solution pH. This could present a safety hazard because of the potential for release of toxic hydrogen cyanide gas. A pretreatment step where the catalysts were contacted with a 1.0M sodium hydroxide solution was found to prevent the pH shift from occurring. Following the sodium hydroxide pretreatment, two stages of leaching at 160 °C with solution containing 1 pct sodium cyanide and 0.1[M] sodium hydroxide gave at least 75 and up to 95 pct Pd recovery. The Pd was quantitatively recovered from the leach solution by thermal decomposition in an autoclave at 250°C for 1 h. The Pd content of the precipitate was over 50 pct. Thermal decomposition also decreased the total cyanide content of the barren solution to less than 0.2 mg/L. The catalyst leach residues passed the Federal Toxicity Characteristic Leaching Procedure and the California Waste Extraction Test, indicating that landfill disposal of the leach residues would be acceptable.

Jan 1, 2010

By G. L. Oldright, Virgil Miller, F. K. Shelton, Keck. W. E., Horst Sieg

Potash is one of the principal foods and is necessary for the growth of most vegetables, grains, and fruits. For many years, this country depended entirely upon foreign sources for its supply of potassium compounds. However, during the last 15 years the domestic production of potash has gradually increased, as shown in table 1. 7/ Imports now supply only half of the American market. In 1936, the equivalent of 207,000 tons of K20 was imported, and 210,000 tons was produced domestically. In 1935, domestic consumption was approximately 420,000 tons of potash, a 57 percent increase over that for 1934 and surpassing previous peaks of 1928 and 1930.

Jun 1, 1937

By Yuanbo Zhang, Zhixiong You, Daoxian Duan, Guanghui Li, Tao Jiang, Xiaohui Fan

"Presently, high-grade manganese oxide ores are usually utilized by a coal-based reduction roasting process. However, there were some obvious drawbacks, including serious environmental pollution, high energy consumption, etc. The reduction temperature is relatively high (850-900 °C). In this study, a kind of biomass, pine black charcoal, was used as a reductant to reduce low-grade manganese oxide ores. Under the optimal conditions of reduction temperature of 600~650 °C, roasting time of 30~45 min and ratio of ore to reductant 5:1, the leaching efficiency of manganese reached more than 98%. Compared with the traditional coal-based reduction process, the reduction temperature is reduced by 250-300 °C.IntroductionManganese plays a very important strategic role in the national economy, and it is widely used not only in the metallurgical industry, but also in the light industry, chemical industry, medicine, agriculture, etc [1]. The average grade of manganese ores in China is evenly about 21%, which belongs to typical low-grade manganese ores [2]. Therefore, large quantity of high-grade manganese ores is required to import in order to meet the rapid development of Chinese manganese industry. As reported, the conventional treatment of reduction roasting has been widely applied for processing manganese oxide ores, in which coals or reducing gases (CO, H2 or CHi) were usually used as reducers [3]. However, there were some insuperable problems, such as high energy consumption, poor operating conditions, serious environmental pollution, etc [ 4-9]."

Jan 1, 2013

By J. Langwaldt

Ammonium-nitrate is a main component in blasting agents and dissolved ammonium residues may be discharged with mine water. The European Union aims to reach a good state of the water quality and recommends ammonium content below 0.5 mg/L for drinking water. Natural zeolites adsorb ammonium from water. In this work, the removal of low-concentration ammonium by commercial natural zeolites was studied in continuous tests. Natural zeolites made up between 83% and 95% of the mineral composition. The major zeolite minerals were chabazite and clinoptiolites and minor amounts of erionite and mordenite were as well present. The five natural zeolite samples, 500~710 µm, were conditioned to obtain zeolites in the Na-form. Ammonium adsorption onto natural zeolites depends mainly on the zeolite mineral type. The chabazite containing zeolite was superior over the four clinoptiolite samples for the removal of low concentrations of ammonium from water (pH of 7.5, flow 5 mL/min, room temperature). The adsorption capacity of chabazite was 48.3 mg NH4+/g zeolites, i.e. 73% of the theoretical exchange capacity 69 mg NH4+/g, during duration over 107 days and almost 6700 bed volumes exchanged. The four tested clinoptiolites had an ad-sorption capacity of 15.7mg to 25.0 mg NH4+/g zeolites. The natural zeolites were rapidly regenerated by an alkaline sodium chloride solution fed to the columns.

Jan 1, 2014

By G. Biles

The Eskay Creek Mine, a high-grade gold and silver underground mine in northwestern British Columbia started operations in late 1994. For the first three years of operation, ore was crushed to minus 2 inch; blended to satisfy smelter terms and direct shipped to smelters in Japan and Canada. Before commencing the above operations, preliminary evaluations had indicated that an extremely complex metallurgical facility, requiring a large capital investment, would be required to recover the precious and other economic metals. Subsequently markets for raw ore were identified and underground mining and ore transportation operations, at a nominal production rate of 300 tonnes per day, were commenced. Exploration diamond drilling in 1995 and 1996 discovered additional lower grade mineralized zones, which are less economic or uneconomic as direct shipping ore. Testwork demonstrated that these lower grade zones could be upgraded to marketable concentrates in a simple flotation mill. The mill was built in 1997 and has been in operation since late 1997. The mining rate now has increased to 400 tonnes per day including 250 tonnes per day of the direct shipping ore and ISO tonnes per day of the mill ore. Prime Resources Group Inc. wholly owns the Eskay Creek Mine. Homestake Canada Inc. owns 50.6% of Prime's shares and manages and operates the mine on Prime's behalf. This paper discusses the mining and milling operations as well as ore shipment and smelter contracts' administration.

Jan 1, 1999

By Luis W. Pommier

The wastes of the plating industry are generated in the process of rinsing the plated objects and during the cleaning of the plating shops. High volumes of hazardous waters are neutralized on site producing recyclable water, generating a metal rich, hazardous, filter cake. Occasionally, a complete plating bath has to be discarded which increases the metal content of the filter cakes. These filter cakes are discarded as hazardous at a cost of over $ 250 per ton. The metals found in the cakes are generally copper, nickel and chromium or zinc, lead and cadmium, assaying from 2% to 15% or even higher content of each metal The metals and their contents are highly variable. Some plating shops discard their products directly to hazardous waste sites, others, have them collected by companies which consolidate wastes from several producers prior to discarding. These wastes, richer in metal content than mine ores, need neither mining nor crushing and grinding for concentration. The commonly found impurities in the mine ores, i.e.: pyrites, arsenic, and valuable metals which are present in minute quantities, are absent in the filter cakes. The processes available to selectively recover the metals from the filter cakes, are discussed based on their technical and economic merits. The flow-sheets include leaching, solvent extraction when required, and electrolytic refining. Examples, showing the separation of copper, nickel and chromium and also of zinc and lead in pilot plants, show the technical and economic viability of the processes.

Jan 1, 1996

By M. Clair Smith

The Bureau of Mines investigated the Rip Van Winkle mine, which is about 25 miles northwest of Elko, Nev., from October 1942 to February 1943 to determine possible resources of lead-zinc-silver ore. The property is owned by the Rip Van Winkle Consolidated Mining Co., which maintains its offices in Salt Lake City, Utah. Ore was discovered in the 1860's, and active mining was carried out intermittently until 1939, when the Rip Van Winkle Consolidate Mining Co. acquired the property. This company equipped and developed the mine, erected a flotation mill, and operated on a 65-ton daily basis. The ore is mined from veins, mineralized faults, and replacement deposits of wall rock. The country rock comprises cherty shale, granite, porphyry, intrusive breccia, and breccia, and brecciated shale. Mining operations were carried down to the 700-foot level. The Bureau of Mines drilled 5 diamond-drill holes, totaling 1,084 feet, from the 600- and 700-foot levels, excavated a 250-foot surface trench, and took 26 underground samples. The ensuing sections note the location and physical features of the district, mention the equipment, and specify the property and ownership. The history of the mine is briefed, and the geology, mine workings, and plant are described. The work performed by the Bureau of Mines is summarized and supplemented in detail by assay maps, logs of diamond-drill holes in appendix A, and by tables of sample analysis in appendix B.

Jan 1, 1949

By Mary F. Koehler

Interoxidic compounds, such as vanadates, have not been studied extensively in the field of thermodynamics. This is a report of three vanadates of sodium--metavanadate, pyrovanadate, and orthovanadate. Vanadates of other metals will be considered in future reports. The heats of formation at 298.15° K. of the three sodium vanadates (both from the oxides and from the elements) were derived from the results of sodium hydroxide solution calorimetry. Free energies of formation were obtained by combining heat and entropy values. The results support the conclusion that the three sodium vanadates were stable with respect to all possible decompositions in which vanadium retains its pentavalency.

Jan 1, 1960

By S. C. Schaefer

Thermodynamic properties of indium oxide (In203) have been studied with a high-temperature galvanic cell employing stabilized zirconia electrolytes. In the temperature range 790° to 1,230° K the free-energy change for the cell reaction 21n(1) + 3Cu20(c) = In203(c) + 6Cu(c) together with the standard free energy of formation of cuprous oxide gives the standard free energy of formation of indium oxide from the elements as ?Gf° = -218.78 + 75.34 X 10-3T ±0.7 kcal/mole. A third-law analysis yields the standard enthalpy of formation: ?Hf°298.15 = -221.5 kcal/mole.

Jan 1, 1971

By W. Mirihange

Considerable mechanical property decline is always observed at the Heat Affected Zone (HAZ) of AA 5083, This occurred as a result of the thermal effect on the hardened material, It is highly beneficial to have a prospective metallurgical procedure to overcome this decline of properties. Therefore, experiments were set to evaluate the effectiveness of a quick post weld heat treatment process with the AA 5083, Optimum results were obtained at 473K and with approximately 600 seconds time, The interplay between the time, temperature and mechanical properties are possible to explain by the Lifshitz-Slyozov-Wagner (LSW) theory and by considering the composition of the material.

Jan 1, 2005

By B. J. Dave, Ashok J. Desai, Allen J. Hulshizer

This paper presents various design and construction considerations involved in developing an underwater installation of lined, large diameter shafts not known to have been previously accomplished for the diameter size, depth or ground conditions identified with the Seabrook Power Station. Outlined approach provides for driving tunnels independent of shaft sinking operation, permits the use of prefabricated shaft lining and incorporates a safe dry tap method for connecting the shafts to the tunnels. INTRODUCTION The Seabrook Power Station, located two miles inland from the Atlantic Ocean, at Seabrook, New Hampshire, is a two unit, nuclear powered, electric generating plant, each unit having a capacity of 1150 megawatts (electric). Typical to all steam driven generators is the requirement for large quantities of water. Requirements for the Seabrook Plant will be met by intaking and discharging approximately 850,000 gpm of ocean water at a considerable distance from the shore line in water depths that render the environmental impact inconsequential. Obtaining, conveying and discharging such large quantities of water represents a major undertaking, particularly under today's environmental controls. Tunnels are well suited to reducing the environmental impact at the Seabrook site by conveying the water about 250 feet deep under the salt marsh (a valuable ecological feature) and under the beach well out into the ocean without any noticeable ground surface disturbance. An alternate approach utilizing trenched piping would have created some disturbances and was environmentally less desirable. The tunnels connect to intake and discharge

Jan 1, 1976

A novel, low-pressure adaptation of Praxair?s coherent jet technology utilizing CoJet lances was developed for application at Inco&apos;s Copper Cliff Smelter where oxygen is available in large tonnage quantities but only at relatively low supply pressures (< 193 kPa header pressure). The project objectives were to evaluate the effect of coherent jet technology on controlling oxygen lance accretion (?horn?) formation and to increase oxygen efficiency in the Inco Oxygen Top-Blowing Nitrogen Bottom Stirring (TBBS) MK Reactor used for converting nickel- contaminated chalcocite to semi-blister copper. This paper describes the development and subsequent implementation and testing of two CoJet lances in this application.

Jan 1, 2005

By D J. Williams, S Islam

Worldwide production of black coal amounts to about 8 billion tonnes. Perhaps 20% of this annual production is washed to improve its quality, generating perhaps 250 million tonnes of coal tailings annually, generally in the form of a slurry. The coal tailings slurry is typically thickened to about 25% solids by mass prior to disposal to a tailings storage facility (TSF), and hence comprise mostly water (about 85% by volume), plus sand, silt and clay-sized mineral particles and carbonaceous material. While the tailings settle and consolidate on deposition in the storage facility, their high moisture content makes them prone to run-out should the tailings dam fail, and makes rehabilitation of the tailings difficult. Hence, it is important to understand the consolidation and shear strength of coal tailings. Laboratory consolidation testing of coal tailings from a slurry to a soil-like state has been carried out in a purpose-built, large-scale, slurry consolidometer, instrumented with load cells and pore water pressure (PWP) transducers. Laboratory vane shear testing of coal tailings produced at different gravimetric moisture contents has also been carried out. The test results have allowed a regression model to be developed relating the peak vane shear strength of the coal tailings to their gravimetric moisture content.CITATION:Islam, S and Williams, D J, 2018. Consolidation and shear strength of coal tailings slurries, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 62–71 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By John P. Albers

During the 1950&apos;s and 1960&apos;s a program of geologic mapping of counties has been underway in Nevada by geologists of the U. S. Geological Survey and the Nevada Bureau of Mines. The aping has been done principally at reconnaissance scales, with more detailed work in local areas. Geologic maps are either published or open-filed for about a dozen counties at scales of 1:200,000 or 1:250,000. One by-product of the geologic mapping has been the recognition of about 80 Tertiary volcanic centers from which large or very large volumes of volcanic material ranging from rhyolitic to andesitic in composition have been erupted. For purposes of this discussion these are regarded as the major centers; the tally does not include hundreds and perhaps thousands of small rhyolitic domes and plugs and basaltic vents. About 15 to 20 of the 80 major centers are believed to be calderas ranging in diameter from 3 to 25 miles. Some have been described in the literature by geologists of the Nevada Test Site and others. However, the vast majority of the 80 centers have not been described in print, and we are indebted to our colleagues for locating them for us in many of the counties. Briefly, we would like to express our appreciation to the following people all members of the USGS unless identified otherwise: Robert Coast, who has identified the centers in Elko County; Ron Willden, Robert Yates, and

Jan 1, 1968

By A. W. Fletcher

Report of discussion at February, 1971, IMM general meeting (Chairman: M. J, Cahalan President) of a paper published in Transactions/Section C (Mineral Processing & Extractive Metallurgy), vol. 79, December, 1970, pp. C247-52, and contributed remarks. Contributions are included from: Cahalan M.J., giving figures related to leaching Cu from pyrite at Rio Tinto 1880-1910; Noakes F.D.L.; Embling P.M.K.; Bradshaw A.V.; Le Roux N.W.; Kelly D.P.; and Tuovinen O.H. Questions were raised regarding impurity removal, effects of light and ultraviolet light on the bacteria, the economics of solvent extraction and the possibilities for other organisms or other metals

Dec 1, 1971

By S. K. Pan, R. Venugopal, L. A. Kumaraswamidhas, Rakesh Prasad, Chandan Pandey

Cold compressive strength (CCS) and apparent porosity (AP) of pellets are essential properties regarding the burden of blast furnaces. The optimal required values for industries are 2.5 KN CCS and 21% AP. Blaine fineness and firing temperature are the decisive parameters analyzed here. During firing, microstructure phases such as silicate, hematite, and pore are formed. The phase’s grain density during induration influences the pellet’s CCS and AP. The influence of the factors over the response is evaluated by I-optimal (response surface methodology) design. Blaine fineness, firing temperature, pore phase grain density (PPGD), silicate phase grain density (SPGD), and hematite phase grain density (HPGD) affect the CCS and AP of pellets. The RSM-I-optimal predicted responses with a coefficient of determination (R2) of 0.89 for CCS and 0.91 for AP of iron ore pellets. Optimal conditions were Blaine no. = 1668 cm2/g, 1250 °C firing temperature, PPGD value 90 no/mm2, SPGD value 250 no/ mm2, and HPGD value 490 no/mm2, with 4.5 KN CCS 27%. The induration processes were investigated through statistical design of experiments for process optimization. Statistical analysis indicated the direct and interactive influence over CCS and AP. In this study, RSM-I-optimal modeling is used for iron ore pellets processing. The I-optimal prediction could be applied to make the process economical because information is obtained through fewer experiments.

Aug 18, 2020

By W. J. Schlitt, K. McNaughton

A heap-leach field test was conducted during the 1993-1994 winter season at Carmacks, Yukon Territory, Canada. The purpose of the test was to demonstrate the viability of heap leaching Carmacks oxide copper ore in an arctic setting on a year-round basis. The test itself utilized a 5-m-diam crib loaded with 250 t of composite ore from the Carmacks copper deposit. The ore was stacked to a total height of 7 m in the crib, including 1 m of ore as an over- layer above the emitter system. This matched the plan for the commercial heap at the time when the test was started. The design also replicated an interior segment of the commercial heap. Leaching was done at a flow rate of 0.2 L/min/m2 using leachate at 21 °C. This temperature was maintained with no external heat input other than normal process heat transferred from electrowinning to the leachate via solvent extraction. Winter conditions were quite typical of those expected at the mine site. The temperature reached -54°C, while averaging -13°C over a three-month period. In spite of some flow system problems, leaching continued unabated, and freezing was limited to the insulating ore over-layer and isolated points near the outer crib walls. Thus, the test demonstrated that year-round heap leaching of Carmacks copper ore is practical in an arctic setting.

Jan 1, 2000