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By R. Krog, S. Schatzel, H. Dougherty

Longwall coal mine operators in the U.S. are required to ventilate multipanel longwall districts, but may have little or no knowledge about what happens to the ventilation air between the inlet evaluation points (IEPs), bleeder evaluation points (BEPs) and bleeder fans. The effectiveness of bleeder performance can directly influence the ability of a ventilation system to remove and dilute coal bed methane emissions. U.S. coal mining stakeholders have acknowledged their belief that the T-junction split at the longwall face tailgate corner is very important in controlling the distribution of ventilation air. To obtain direct measurements of bleeder performance, a tracer gas, sulfur hexafluoride (SF6), was released into the ventilation air stream on active and inactive longwall panels. Testing was performed on multiple longwall panels that included various phases of longwall development and variable path lengths of ventilation air transport to bleeder fan installations. Changes in the T-junction ventilation air distribution over the life of a longwall panel, and its variable effect on panel ventilation airflow are discussed. These findings will assess the effectiveness of commonly applied ventilation strategies for improving air distribution and ventilation controls to meet statutory requirements.

Jan 1, 2011

By Robert E. Kirby

The potash mine and refinery of AMAX Chemical Corporation is located 26 miles east of Carlsbad on the north end of the potash producing areas. This property was developed between 1949-1952 and production of muriate began in August 1952. Two shafts, located 400 feet apart, were sunk to the ore zone 900 feet below the surface. A 15-foot diameter ore hoisting shaft equipped with a 1000-ton ore pocket, crusher and automatic skip loading equipment plus a 20-foot diameter service shaft with a curtain wall for exhausting air were provided. Sylvinite production to date has been from the first ore zone, stratigraphically, the lowest of ten identified mineralized zones in the salt section. This first ore zone has been the top producer of Muriate of Potash over the years. In our area we do have some very marginal mineralization in the third ore zone some 35 feet above the first. We have not mined from this seam as yet but we are investigating the economics very closely with access slopes and an experimental mining panel. Our ore seam is generally flat lying but does have local gentle rolling with grades from 2% to 5%. The mine was designed and developed from the beginning with conventional mining practices and rail haulage from the mining section to the ore pocket. As mining progressed in the early years, production increased from less than one million tons annually to a maximum of 3,935,000 tons in 1964. Efficiency also kept pace with the increasing production and continually improved until the efficiency of the production crews reached an average of 100.3 tons per manshift in 1964.

Jan 1, 1974

By D. F. Lillie

The milling plant of Bicroft Uranium Mines Limited has been in operation for a period of 2 years. During this time many changes, both physical and chemical, have been made in an effort to improve plant efficiency. Investigational work has been facilitated greatly by the fact that the Bicroft mill consists to a large extent of two parallel circuits; however, the number of inter-related variables is great, and it is difficult to measure, let alone control, all of them simultaneously. The Bicroft mill staff feel that much progress has been made, but there is much yet to be done and many questions still to be answered. The property of Bicroft Uranium Mines is located in Cardiff Township in the Highlands of Haliburton County, southeastern Ontario, some 140 miles northeast of the city of Toronto and 70 miles due north of Lake Ontario. The ore-bearing dikes are classed as mafic, non-segregated, syenite and granite pegmatites. Microline, microline perthite, quartz, uraninite, uranotharite, fluorite, titanite, cyrolite, calcite, pyrite, pyrrhotite and molybdenite as important accessories. The mill was designed for a nominal capacity of 1,000 tpd with all sections except grinding, leaching and clarification capable of 1,500 tons. Feed was entered to one mill unit on October 29, 1956 and to the second parallel line on November 13. The total feed rate was at 1,000 tpd on November 15. The first precipitate was produced on November 12. By August 1958 the average entry had been increased to 1,410 tpd without the installation of any additional major equipment. The overall recovery has been greater than 93,0 pet do several months, but a recovery of 92.5 pct appears to be most economical. Operating time for 8 months in 1958 has been 97.1 pet although all grinding mills were relined during this period.

Jan 1, 1959

By Paul H. Johnson

Chemical mining may be defined as the in-place extraction of metal values from ores located within the confines of a mine (that is, broken or fractured ore, stope fill, caved material, or ores in permeable zones) or ores located in dumps, prepared ore heaps, or tailing ponds. The term solution mining has, in general, been used to mean the in-place dissolution of water-soluble mineral salts with leach solutions. Chemical mining is presently in its embryonic stage. Basic research needs to be done to understand the influence of physical and chemical factors in chemical mining, and applied research needs to be done on both a small and a large scale to develop practical techniques of field practice. The areas of research include the following: (1) Leaching chemistry, the influence of physical, chemical, and biological factors in in-place leaching (2) Hydrodynamics of chemical mining systems (3) The development of new, as well as old, leaching agents for use under chemical mining leach conditions (4) The development of new means of contacting fine and coarse materials (with solutions) under static leach conditions (5) The development of new processes for generating and regenerating chemical mining leach solutions (6) The development of metal recovery processes

Jan 1, 1969

By J. Jorge, J. Mariño-Salguero, B. Álvarez-Rodriguez, J. M. Menéndez-Aguado, J. J. de Felipe

"This work focuses on the characterization of heat generation during dry fracture by direct impact of a tantalum ore. Moreover, an attempt was made to quantify the heat energy loss to the environment during the grinding process of a ball mill. The study was carried out using mineral samples from a Spanish mine that were characterized by measuring their density and specific heat. During the experiments, increases of temperature ranging from 5.9 to 7.3 °C after the fracturing process were observed. A numerical model was then applied to calculate temperature distributions through the mill wall. Three possible combinations of heat transfer coefficients were analyzed, from (1) Kapakyulu and Moys, (2) Duda and (3) the heat transfer laboratory of Universitat Politècnica de Catalunya. The models with the coefficients from Kapakyulu and Moys and the heat transfer laboratory of Universitat Politècnica de Catalunya were found to provide results that were in best agreement with the experimental data. IntroductionEvery year, several billion tons of metallic ores, minerals, cement and other solids used in the ceramic and chemical industries are subjected to size reduction in ball mills around the world. The U.S. Department of Energy (2005) reported that in the United States about 29 percent of the total energy used by the mining industry, or 0.4 percent of total U.S. energy consumption, is used in comminution. In Canada and Australia, comminution is responsible for about 2 percent and 1.5 percent of total energy use, respectively. In nonindustrialized countries where mining constitutes higher fractions of their industrial sectors, comminution costs represent a significant contribution to mine operating costs. The specific energy consumption values for the grinding of these materials typically range from 5 to 50 kWh/t. Thus, a significant amount of electrical energy is consumed in the ball mill grinding operation (Aguado et al., 2006; Gupta and Sharma, 2014). Grinding can represent up to 50 percent of the direct operating cost in a mineral processing plant, especially as comminution devices operate in highly inefficient ranges of 4 to 8 percent (Wang, Yang and Yu, 2012)."

Jan 1, 2017

By K. Avarmaa, P. Taskinen

"The solubilities of rhodium (Rh) in a molten silver alloy in doré-smelting conditions for treating copper refinery slime at temperatures of 1,000 to 1,300 °C were measured using a high-temperature equilibration-quenching technique and direct phase analysis by electron probe X-ray microanalysis. The results obtained indicate that the solubility of rhodium in molten silver is very limited: only 0.04 to 0.26 weight percent at 1,000 and 1,300 °C, respectively. The thermodynamic properties of rhodium in molten silver were estimated from the solubility data. The compositions of the solid rhodium alloys in equilibrium with molten silver-rich doré metal were also measured in this study. The excess rhodium in the doré alloy will precipitate during the refining as solid, greater than 99 weight percent Rh.IntroductionSilver refining is a core step of the processes of anode slime treatment and precious metals recovery in copper refineries (Hait, Jana and Sanyal, 2009; Ludvigsson and Larsson, 2003). It separates the oxidic materials, such as nickel oxide (NiO) and barium sulfate (BaSO4), from metallics after the de-coppering and de-selenation steps of the anode slime, and refines the silver bullion from the undesired elements present (Cooper, 1990; Hyvärinen, Lindroos and Yllö, 1989). Doré metal is a silver-rich alloy which after the smelting step will be refined electrolytically, and the remaining constituents of gold as well as the platinum group metals (PGM) will be subsequently recovered in their own processing stages.Because of low concentrations of PGMs in copper, they are typically soluble in the doré metal in anode slime treatment. This can be concluded from their binary phase diagram data with silver (Massalski et al., 2007). One exception is sometimes encountered in the anode slime smelting operations if the smelter feedstock is rich in rhodium, as may be the case if the raw materials contain secondary raw materials such as e-scrap. The solubility of rhodium in solid silver is less than 0.04 weight percent (Królas and Sternik, 1995), and its solubility in molten silver is highly uncertain. The only experimental information in the literature (Królas et al., 1993; Elliott, 1985) suggests that the silver-rhodium alloy system at 960 °C forms a eutectic, which locates less than 2 °C below the melting point of pure silver (Arblaster, 2015). Early observations (Rössler, 1900) propose that rhodium is insoluble in molten silver even if solid silver-rhodium solutions have been recently produced on a nanoscale (Kusada et al., 2010).The aim of this study is to measure experimentally the solubility of rhodium in molten silver at doré metal smelting and refining temperatures. The silver alloy used to mimic the doré alloy contained silver and about 1 weight percent of the alloying elements copper, gold, palladium, rhodium and tellurium. The experiments were carried out in metal-slag equilibria with nitrogen-oxygen mixtures as the atmosphere, employing an equilibration-quenching technique and using direct phase composition analysis by electron probe X-ray microanalysis (EPMA) (Jak, Hayes and Lee, 1995)."

Jan 1, 2017

By F. Arslan

Gold extraction from refractory ores is still a problem. Gold losses in the cyanidation process are mainly due to the presence of sulfides and carbonaceous materials. Oxidation of carbon and sulfide minerals by sodium hypochlorite is a possible method to render the ore amenable to cyanidation, and earlier work by the USBM indicates that electrochemical generation of hypochlorite is promising. Laboratory work was carried out on the oxidation of pyrite and carbon in an electrochemical cell with a sodium chloride electrolyte. The results of preliminary tests are discussed in relation to the mechanism of electrooxidation of refractory gold ores and the estimation of current efficiency.

Jan 1, 1991

By David T. Kao

The energy situation the world faced during the decade of the Seventies resulted in actions throughout the free world to reduce its dependence on OPEC oil. Many nations heeded the indicators early in that decade and commenced efforts to reduce their oil requirements. Those nations which did not heed these warnings had the message brought to them more emphatically in the last portion of the decade with the further rapid and continuing rise in prices of OPEC oil. Furthermore, the demonstrated unstable nature of large oil-producing countries has made it clear to all free world nations that their energy demands of the future must be provided for by some more stable and assured supply. Now, early in the decade of the Eighties, we see the need for expanded use of coal as an energy supply to be essential. The result in 1980 and 1981 has been a great increase in demand for steam coal worldwide. In essence, a new business for the United States has developed in steam coal export. In the United States, the large growth in foreign demand for our steam coal is now well recognized and documented. A major impediment in meeting this growing demand became abundantly clear in 1980. This impediment is the congestion existing at U.S. coal export ports. With the recognition of this constraint, a plethora of activity has been observed in announcements for alleviating the problem by improving and expanding current export port facilities. In addition to improvement and expansion, an abundance of totally new facilities has been announced in an effort to find ways to meet this large demand in the new business of steam coal export. This is needed; and hopefully, many of these plans will be translated into action. While total implementation of all stated plans is not expected, much of the implementation depends upon large- scale dredging, large waterside ground storage sites, and improved handling equipments in the conventional type of export ports. The time and capital required for implementation is great, and environmental problems associated with dredging are substantial. Host significantly, and unfortunately, implementation that does include dredging our ports will still not permit use of the truly large colliers of 140,000 DWT and greater, which are evolving as the "state of the art" in long-distance sea transport.

Jan 1, 1981

By W. J. Douglas

The mine manager in today's environment is faced with a wide variety of problems and challenges which must be addressed if efficient and profitable operations are to be achieved. Expansion of operations has, in many instances, resulted in inexperienced work crews, face supervisors, and superintendents. The complexities of a mining system preclude rapid movement through the learning curve for moat personnel. The management of production suffers both from this inexperience and the lack of standard procedures and methods for planning, schedul ing and analyzing production systems. The lack of visibility of information, as well as of the operations themselves, due to their remoteness, adds to the problem of production management.

Jan 1, 1980

By Paul G. Mahoney

The original Radum aggregate plant, built by Henry Kaiser in 1931, is the oldest of the nearly 200 individual operations established by the Kaiser companies. Since that time it has been producing aggregate products primarily for the San Francisco Bay Area. Currently, and for many years, this plant has been listed as the Nation's No. One aggregate producer. Early this year a new $10,000,000 facility went on line replacing the old plant. Construction of the facilities was performed under the direction of Kaiser Engineers construction management teams. The existing plant continued to operate without interruption during the entire construction period.

Jan 1, 1970

By R. Wild

[When a blast furnace has a certain defined burden and is operated under fixed conditions of blast temperature, etc., the fuel efficiency is determined by the extent to which the reducing gases can remove oxygen from the burden in the furnace stack. This is determined by two distinct factors: 1) The uniformity of gas-solid contact, and 2) The ease with which oxygen can be removed from individual pieces of burden. This latter is often called burden reducibility. When burdens were poorly prepared the first factor was by far the most important and a study of the reducibility of individual lumps was of rather academic interest recent years good burden preparation with emphasis on sized material has led to greatly improved gas distribution in the stack, and thus the second factor has become much more important and there has been a marked increase in interest in methods of measuring reducibility.]

Jan 1, 1964

By David Oliveira

"INTRODUCTIONOne of the key drivers for the selection of a mechanized tunneling method in good quality rock masses is a good excavation advance rate. In order to achieve satisfactory advance rates, a good geotechnical performance of the rock grippers in main beam open tunnel boring machines (TBM) is required, besides other factors related to ground support and the performance of the cutter head.One well known potential issue of main beam gripper TBM is associated with the bearing capacity of the tunnel side walls while crossing fault zones or poor rock masses in general. Such a mechanism is similar to that of footings on rock and perhaps the most common problem associated with gripper performance. Depending on the quality of the rock, the rock under the gripper contact may experience significant damage and crushing (Figure 1), eventually affecting the TBM advance rate due to a lack of tunnel wall reaction/support. Such a mechanism is primarily dependent on rock quality, applied loads and gripper size.In the case of large diameter TBMs (e.g., D > 12 m), a second potential failure mechanism may arise associated with the loss of confinement within a rock pillar between twin tunnels, even in good quality rock but particularly in bedded sedimentary strata. Punching of the rock pillar may develop due to excessive shear along bedding partings and sub-vertical joints (Figure 2), particularly under a low stress environment. As a result, this mechanism is dependent on both the rock pillar width between twin tunnels and stress conditions around the tunnels, in addition to the previous factors described for the first failure mechanism."

Jan 1, 2016

By J. F. Burr, D. P. Snyder, G. W. Luxbacher

"The Mine Improvement and New Emergency Response Act (MINER Act) of 2006 established the Office of Mine Safety and Health (OMSH) within NIOSH and provided authority for awarding contracts related to the development and implementation of new mine technology and equipment. Subsequently, Congress provided additional funding to NIOSH to initiate and continue the program. OMSHR has fulfilled this Congressional mandate through annual Broad Agency Announcement (BAA) solicitations, directed Requests for Proposals (RFP) and Inter-Agency Agreements (IAA), and has further expanded the program through capacity building BAA contracts with U.S. university mining departments in the areas of mine ventilation and ground control. This paper describes the extramural funding program and broadly examines the results of the 150+ contracts issued to date, grouped into key focus areas. INTRODUCTION The Mine Improvement and New Emergency Response Act (MINER Act) of 20061 established the Office of Mine Safety and Health Research (functioning as the Office of Mine Safety and Health Research or OMSHR) within the National Institute for Occupational Safety and Health (NIOSH), and provided authority for awarding contracts related to the development and implementation of new mine technology and equipment. The OMSHR external contracts program was formed in accordance with the MINER Act and has since issued over 150 contracts focused on worker health and safety. Initial funding for this program was provided by Congress through two Emergency Supplemental Appropriations (ESA) and was focused on emergency oxygen supply, communications and tracking, and refuge chambers (also referred to as refuge alternatives). Since the MINER Act, Congress has continued to fund the program through the annual NIOSH mining appropriation. The program was supplemented starting in 2009 with an academic capacity building program directed toward developing university tenure-track faculty and facilities and supporting graduate students in the area of mine ventilation; ground control was added as an additional focus area in 2011. A short synopsis of each contract is published on the NIOSH Mining website2 and final reports are available in digital form on request. A summary of the progress OMSHR has made over the 10-year period since the Act was published previously.3"

Jan 1, 2018

By William X. Chavez

The Mantos Blancos copper-silver district is located in northern Chile about 45 kilometers northeast of Antofagasta, as shown in Figure 1. Disseminated and veinlet copper and copper-iron sulfides containing economically significant silver values are hosted by andesitic to rhyolitic volcanic rocks of probable Jurassic age (Chavez, 1983). Field, mineralogical, textural and lithochemical evidence show that three major alteration types may be recognized within the district: (1) district-wide sodium metasomatism, manifested as albitization of feldspars and development of albite veinlets and pore-fillings, (2) local to widespread, incipient to intense addition (chloritization) or removal (bleaching) of magnesium and iron and (3) local to widespread, moderate to intense hematitization, usually spatially associated with bornite-digenite-covellite (as specularite) or digenite-covellite-chalcocite (as red hematite staining with or without specularite) mineralization. Carbonatization, occurring as veins, veinlets, vesicle fillings and irregular patches of calcite, varies in magnitude of development but is widespread. Sericitization of feldspars and groundmass is only locally intense yet is generally present as incipient alteration throughout the district. The district-wide albitization of the Mantos Blancos volcanic sequence has altered what were originally more potassic and calcic, less sodic, rocks to what are now more properly referred to as keratophyres (Battey, 1955; Gilluly, 1935, Hughes, 1975). The petrographic and lithochemical characteristics of the three major alteration types noted above are summarized below.

Jan 1, 1984

By I. E. Odom

The physiochemical properties and industrial uses of the clay mineral montmorillonite are reviewed. Montmorillonite is the most abundant species of the smectite clay mineral group and is the predominant mineral composing bentonite deposits. The crystal structure, chemical composition, exchangeable ion type, and small crystal size are responsible for several unique properties, including a large chemically active surface area, a high cation exchange capacity, and interlamellar surfaces having unusual hydration characteristics. The type of exchangeable ion, sodium (Na), calcium (Ca), and/or magnesium (Mg), has a major influence on industrial and chemical uses of montmorillonite because of their hydration characteristics. Natural occurrences of Na mont¬morillonite in commercial quantities are geographically limited, but Ca-Mg montmorillonite deposits are known worldwide. Foundry, well drilling, pelletizing, and civil engineering constitute the major uses of montmorillonite, but the total array of uses exceeds that of any other mineral.

Jan 1, 1988

By P. S. Sundar

A laboratory method for the simulation of in-situ leaching of uranium was developed under contract to the Interior Department's Bureau of Mines. Simulation tests using sulfuric acid, ammonium bicarbonate and sodium carbonate as reagents and hydrogen peroxide as oxidant were carried out with representative ores from Texas and Wyoming. The tests not only yielded recovery rates of uranium and permeability changes, but also qualitative inferences on the chemical and physical phenomena which occur underground. Leaching with basic solutions was found to be kinetic constrained and with acid solutions, equilibrium constrained. The rapid permeability changes observed when leach solutions were introduced, were probably caused by gas blockages in the packing. Precipitation of calcite or gypsum and the reaction of clay to leach solutions had slow and gradual effect on the packing permeabilities.

Jan 1, 1978

By J. F. Allan

The main conditions which affect investment in mineral exploration and development are compared in five countries currently considered attractive for mineral investment - Canada, the U.S., Australia, Brazil and the Philippines. The primary parameters - the probability of occurrence and discovery of mineral deposits and the short and long term viability of mineral development - are evaluated and compared on a relative basis by a form of factor analysis. This analysis indicates that of the five countries studied, Canada is second only to the United States in degree of favourability for mineral exploration and development.

Jan 1, 1976

By K. Marutiram, T. Umadevi, R. Sah, K. Abhishek

Low-grade iron ore received from the West Singhbhum area of Jharkhand state in India was used to develop a beneficiation flow sheet to remove alumina (Al2O3) and silica (SiO2 ) from the ore to get iron (Fe)-rich product for steel plants. Mineral characterization showed that the main mineral constituents present in the head sample are goethite, limonite, hematite, quartz, clay and gibbsite. The head sample contained about 49.2 percent Fe, 8.79 percent alumina and 12.01 percent silica. The received iron ore sample is amenable to upgradation through beneficiation to get pellet grade concentrate. The beneficiation process was established, involving crushing, screening, jigging, spiral separation, size reduction, desliming and magnetic separation. The process showed significant enrichment in Fe values to 63.5 percent Fe in concentrate, from 49.2 percent Fe. From the developed flow sheet, it is possible to obtain pellet grade fines with 63.5 percent Fe, 2.92 percent SiO2 and 2.61 percent Al2O3 with 50.6 percent weight recovery. The overall tailing loss is 49.4 percent with 34.3 percent Fe, 21.28 percent SiO2 and 14.64 percent Al2O3

Compagnie Minière Espérance mines a primary gold orebody in an openpit. The ore is processed through crushing, primary gravity concentration, grinding and secondary gravity separation, generating tailings with remaining fine gold. Analysis of the material in the pond has indicated a residual gold grade sufficient to justify a chemical treatment process to recover the gold. However, direct cyanidation would require a prohibitively large installation for the site, which is located in a remote rainforest environment. It was decided to test the feasibility of preconcentration using flotation. As lab tests proved up the technical feasibility of the flotation process, an on-site pilot operation was carried out to assess the continuous circuit performances, define the best configuration and obtain enough data for industrial circuit sizing. Data collected during the pilot campaign was used for calibrating a process simulator built with the USIM? PAC software. The simulator validated the feasibility of upgrading the existing plant, including the reprocessing of tailings to obtain an acceptable fine gold recovery from a small amount of concentrate.

Jan 1, 2010

By R. A. Pyper

Laboratory leaching experiments were conducted to ascertain the effectiveness of acidic solutions of thiourea, CS(NH2)2, for dissolving gold from finely disseminated gold ores. The ore tested was from the Chipmunk orebody of the Northumberland mining district in Nye County, Nevada, USA. Extractions of up to 90 percent were attained from the ore. Extraction was shown to be dependent on thiourea concentration, ferric sulfate oxidant concentration, acid concentration, and leaching time. Optimum conditions were an ambient temperature leach near pH 1.0 utilizing 100-200 kg/t thiourea and 50 kg/t ferric sulfate for 8 to 24 hours.

Jan 1, 1981