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By MAURICE SUNKPAL, Taghi Sherizadeh, Kutay E. Karadeniz, Dogukan Guner, Samuel Nowak

Despite numerous empirical, numerical, and hybrid studies conducted on the subject of coal rib failure, mine fatalities related to rib failures continue to occur at around 1.5 per year between 2009 and 2021. To date, a standard methodology for the design of coal rib supports that can accommodate the wide range of unique conditions found in coal mines in the USA has not been developed, forcing mine operators to rely on the trial-and-error process or industrial legacy practices for the design of coal rib support patterns. In this study, to better understand the support systems’ effectiveness, various rib bolting scenarios were investigated. A coal mass constitutive model is implemented to a distinct element method (DEM) code, and rib factor-of-safety (RibFOS) analyses were performed by applying strength-reduction for the coal matrix and explicitly modeled face cleats. Ribs with various overburden depths and mining heights composed of solid Banded-Bright coal (BBC) were considered in this study. Different scenarios such as rib bolt length and the number of bolts were studied to quantify the effect of primary rib support density on rib stability. As a result of the unsupported coal rib analyses, a decaying exponential relationship is proposed between the mining height-overburden depth and the RibFOS. The developed coal rib models clearly show the coal-rib bolt interaction and illustrate how the addition of rib bolts restricts the movement of the coal rib. The results reveal that the strength reduction method in DEM code with the coal-mass model can aid in the development of practical primary support charts, which can help in the rib support design process. In addition, an unsupported RibFOS cutoff is proposed, beyond which the addition of primary support cannot stabilize the failed rib structure.

Aug 5, 2023

By James Westfield

I welcome this opportunity to appear before you and to review briefly the various health and safety services of the Bureau of Mines available to the mining industry, other than coal, and also to tell of our plans for extending these services to improve safety and health conditions. The Bureau's health and safety activities are basically services by which we attempt to develop and disseminate information on improved techniques in mine accident prevention and promote mine safety education in the mineral industries. Unfortunately, some mining companies, particularly some of the smaller operations, seem to be unfamiliar with these services, The Bureau of Mines is charged by law with the responsibility for inspecting coal mines; other services to the coal mining industry and are provided only on request, Safety inspections and other services that we offer are provided for noncoal mining industries solely upon request.

Jan 1, 1958

By D. L. Lampshire

Biological leaching using native heterotrophic bacteria was studied by the U.S. Bureau of Mines as a means of extracting manganese from a domestic low-grade wad ore. Column heap leaching simulation techniques were employed using molasses as the nutrient source for the bacteria. These column studies were designed to investigate the effects of molasses medium flow rate, concentration, and replacement interval on the extent and rate of manganese extraction. .Results showed that flow rates between 0.41 and 8.1 L/min/m2 had little impact on manganese extraction. The total amount of molasses used did have a direct effect on the extent and rate of manganese extraction. Manganese extractions over 90-pct were obtained in 12 weeks of bioleaching with 2-week medium replacement cycles.

Jan 1, 1993

By N. A. Karl, D. R. Wilburn

"This summary of international mineral exploration activities for the year 2015 draws upon information from industry sources, published literature, SNL Metals & Mining (SNL) (Charlottesville, VA) database and specialists in the U.S. Geological Survey (USGS) National Minerals Information Center. The summary provides data on exploration budgets by region and mineral commodity, identifies significant mineral discoveries and areas of mineral exploration, discusses government programs affecting the mineral exploration industry, and presents analyses of exploration activities performed by the mineral industry.Three types of information are reported and analyzed in this annual review of international exploration: budgetary statistics expressed in U.S. dollars provided by SNL; regional and site-specific exploration activities that took place in 2015 as compiled by the USGS; and regional events that affected exploration activities including economic, social and political conditions, which were derived from published sources and discussions with USGS and industry specialists. Commodity and regional compilations are presented in this summary. Because multiple sources were used to develop commodity and regional compilations, statistics may vary depending on the source and type of data that are being reported.The SNL data summarize planned company budgets for worldwide exploration activities in 2015 for 24 nonfuel mineral commodities and metal alloys and are based on company surveys. Surveys focused on precious metals (gold, platinum-group metals and silver), base metals, (cobalt, copper, lead, molybdenum, nickel, tin and zinc), bulk commodities (bauxite, chromite, iron ore, manganese, phosphate and potash), and other commodities (diamond, graphite, heavy-mineral sands, lithium, niobium, rareearth elements (REE), tantalum, tungsten and uranium).USGS data compilations and analyses are based on information provided by USGS mineral commodity and country specialists, as well as industry contacts and published trade journals. The USGS data summarize exploration site data collected for more than 80 minerals and materials, with a focus on base metals, diamond and precious metals. The USGS analyzed available information to assess the level of exploration activity in 2015 and to report trends in mineral exploration activity for the period 2005 through 2015. These analyses identify where mineral exploration is occurring by commodity and region and the intensity of activity that is taking place in each region for selected mineral commodities, and determines those factors that most affect changes in exploration activity."

May 1, 2016

By J. G. Groppo

A series of performance studies have been conducted using counter-current column flotation to recover a marketable clean coal product from existing preparation plant fine waste streams. The test work was conducted at several preparation plant sites in Kentucky and Virginia and utilized a 0.15m (6 in.) I.D. column that was 5.79m (19 ft.) high. The refuse streams tested included both thickener feed and classifying cyclone overflow with a typical size distribution of 100 mesh X O. The percent solids of the column flotation feed varied from 1.5 to 15% while the ash content ranged from 30-60%. The concentrates produced contained 2-10% ash with BTU recoveries of nearly 90% using 0.12-1.25 kg/ton (0.25-2.5 lbs/ton) fuel oil and 0.12 kg/ton (0.25 lb/ton) MIBC. These results were consistent with or better than results obtained in preliminary laboratory scale column flotation evaluation. Test results from 7 different seams are discussed as well as scale-up relationships utilized to design a full scale installation presently in operation.

Jan 1, 1990

By D. R. Hanson, H. Lawson

In this study, NIOSH researchers conducted a machine learning analysis to examine whether a model could be constructed to assess the probability of dynamic failure occurrence based on geochemical and petrographic data. Random forest, cluster analysis and dimension reduction were all applied. The objective of dimensionality reduction was to explore patterns and groupings in the data and search for relations between compositional parameters. Cluster analyses were performed to determine if an algorithm could find clusters with given class memberships and to what extent misclassifications of dynamic failure status occurred. A random forest analysis performed on data from the Pennsylvania Coal Sample Databank cross-referenced with accident data from the Mine Safety and Health Administration (MSHA) determined that 7 parameters of the 18 examined exerted the most influence on results. Cluster analysis on data after dimensionality reduction resulted in a hierarchal clustering algorithm finding four clusters, with one relatively distinct dynamic failure cluster, and three clusters consisting mostly of control group members but with a small number of dynamic failure members.

Feb 1, 2023

By H. Mischo, R. Ellmies

Uranium has been mined in Namibia since the 1970s. The worldwide increasing need for energy in the early 2000s has led to the development and opening of a second uranium mine in 2007 and has triggered an increased interest in uranium exploration, often referred to as “the Namibian Uranium Rush.” In all, four mining licenses have been granted by the Namibian Government, which currently has two mines in operation, a third undertaking trial mining and the fourth in project finance negotiations. In addition, highly intensive prospecting activities at additional deposits are at an advanced stage. The finalization of the strategic environmental impact assessment (SEA) on uranium mining in 2010, the first of its kind and scale in the world, has enabled the Namibian government to assess this uranium rush and its tremendous legal, socioeconomic and environmental impacts and to prepare for the different future scenarios conceivable when this article was written in 2011, including both a skyrocketing and a complete breakdown of demand scenario. The Fukushima disaster and plans of the Namibian government to significantly increase royalties and company taxes in 2011 have threatened the market situation, forcing investors to reevaluate Namibian uranium mining projects. However, with the tax and royalty increase initiative since withdrawn, most projects are back on track at this point in time.

Jan 1, 2012

By John C. Ferm

The role of geology in coal exploration, mine planning and mine development has, in the past, been relatively small. The primary reason for this lies in. the simplicity of geologic concepts that were necessary to bring these operations about. These concepts can be briefly stated as follows: 1) coal occurs in the beds or layers and is overlain and underlain by shales, sandstone or limestone in varying porportions, 2) each coal bed has been given a name and certain quality characteristics can often be associated with a named bed and 3) coal beds (and their adjoining rocks) are often folded into broad anticlines or synclines and, in some places, displaced or faulted. In actual practice, however, these simple principles have been found wanting. Thickening, thinning, pinch outs or changes in coal quality occurred but the occurrence was apparently unpredictable and the only real solution to such problems seemed to be practice and experience. This combined with ingeneous engineering solutions have provided answers to most questions. Perhaps a more fundamental reason that geology could offer so little input to coal mining is that the thickest, most persistant and best quality coal seams followed the above stated rules fairly well. In many areas, however, such seams are now nearing exhaustion and the increased demand for clean, non-polluting, safe energy is quickly bringing a need for techniques of exploration and mining procedures that will make development of formerly unminable seams a profitable venture. For the past 20 years some considerable body of knowledge has been added to the framework of the geology of sedimentary rocks and it is the purpose of this paper to present some of these newer ideas that will be useful in solving some of the present problems of coal exploration and development.

Jan 1, 1976

By P. Karimi, A. Akcil, A. Amini, H. Abdollahi

"In this study, cyanidation experiments were carried out on a massive sulfide ore sample, whose grades of the precious metals gold and silver were found to be 4.2 g/t and 311.9 g/t, respectively. Other elements in the ore were arsenic with average grade of 1,751.6 g/t; antimony, 619.7 g/t; zinc, 781.4 g/t; copper, 454.6 g/t; lead, 101.8 g/t; mercury, 324.3 g/t; and bismuth, 0.6 g/t. The first objective of test work focused on determining the best conditions for extracting gold from the ore, ground to less than 75 microns. Different influential parameters such as particle size fraction, cyanide concentration, pH and leaching time were investigated. The optimum parameters were found to be particle size fraction of less than 25 microns, cyanide concentration of 2,000 mg/L, pH of 11 and cyanidation time of 24 hours. In these optimum conditions, the highest gold and silver leaching efficiencies were 76.88 and 59.63 percent, respectively. In addition, roasting as pretreatment at a temperature of 700°C over a period of two hours was evaluated. Cyanidation testing of the roasted ore in the optimum conditions found increases in the gold and silver leaching efficiencies to 85.97 and 67.57 percent, respectively. Moreover, antimony extraction rose to 43.10 percent, from 23.73 percent without roasting.IntroductionGenerally, gold ores can be classified as “free milling” or “refractory,” depending on how they respond to cyanide leaching (La Brooy et al., 1994; Celep et al., 2009). While high gold recoveries (> 90 percent) from free-milling gold ores can be readily achieved, refractory gold ores are often characterized by low gold recoveries (50-80 percent) within a conventional cyanide leaching process (Rubisov et al., 1996; Adams, 2005; Celep et al., 2009). The presence of preg-robber components such as carbonaceous matter, sulfide minerals and aluminosilicate can further decrease gold and silver dissolution rates and recoveries. With the negative effect of the presence of preg-robbers, gold dissolution is considerably diminished in solutions with high cyanide concentration (Tan et al., 2005). In such refractory gold ores, gold particles may sometimes be occluded or included in the sulfide minerals, and pretreatment becomes necessary to break down the mineral structure to liberate gold for subsequent recovery. These ores are usually pretreated with an oxidizing process, after which gold and silver can be recovered by a standard cyanidation process. The main oxidizing processes available for refractory gold ores are roasting (LaBrooy et al., 1994; Angelidis and Kydros, 1995; Li et al., 2009; Karimi et al., 2010; Kyle et al., 2012), biological oxidation (Lawrence and Bruynesteyn, 1983; Komnitsas and Pooley, 1990, 1991; Deng and Liao, 2002; Ciftci and Akcil, 2010, 2013) and pressure oxidation (Mason, 1990; Murthy et al., 1992; Qingcui et al., 2009). In addition, ultra-fine grinding is being used in some plants for the liberation of gold and silver particles that are interlocked in the sulfide or silicate minerals (Celep et al., 2011)"

Jan 1, 2015

By T. Grammatikopoulos, C. Gibson, M. Aghamirian

"Lithium is a silver-white alkali metal that has recently become a target commodity for many junior mining companies. Lithium has a wide range of applications including ceramics and glass, lubricating greases, air treatment, continuous cast mold flux powders, polymer production, and primary aluminium production, but the current popularity of lithium can be attributed to an anticipated increase in demand for lithium ion batteries. The most common lithium-bearing minerals occur globally in pegmatites and granites. Spodumene [LiAl(Si2O6)] is the most desirable economic lithium-bearing mineral due to its high lithium content (theoretical Li2O content of about 8 wt%). However, lithium can also be hosted in petalite, micas, tourmaline, and phosphate minerals. In Canada there are a number of spodumene deposits in Ontario and Quebec at development stage that have the potential to help narrow the forecast lithium supply gap stemming from the widespread acceptance and use of electric vehicles. This paper discusses the beneficiation of hard rock lithium deposits, including an in depth review of the use of flotation to upgrade spodumene. A review of the use of dense media separation (DMS), magnetic separation, and roasting as means to process spodumene bearing ores is also discussed. INTRODUCTION Lithium is a soft, silver-white alkali metal with atomic number 3. As of 2015, the end use of lithium was estimated to be as follows: lithium-ion batteries (35%), ceramics and glass (32%), lubricating greases (9%), air treatment (5%), continuous casting mould flux powders (5%), polymer production (4%), primary aluminium production (1%) and other uses (9%) (Jaskula, 2014). Lithium is found in brine lake deposits containing lithium chloride and in hard rock pegmatite deposits. It can be sold in a range of product types; as mineral concentrate, brine, compound (lithium carbonate or lithium hydroxide) and metal depending on its intended end use (Goonan, 2012; Jaskula 2014). In the 1990s the United States and Chile were the largest producers of lithium globally, primarily from brine operations (Ober, 1994). Most of the lithium compounds produced over this period were consumed in the glass, ceramics and aluminium industries. In the 1980s and 1990s there were two major lithium operations in the United States: brine operations at Silver Peak, Nevada, and a spodumene mine and lithium carbonate/lithium hydroxide plant in Kings Mountain, North Carolina. The Kings Mountain operations ceased production in 1991 and were eventually dismantled in 1994. Brine operations were also active in Argentina and Chile, while the largest lithium ore deposit in the world was operated at Greenbushes, Western Australia. During this period it was predicted that major shifts in demand for lithium would come from the success and widespread use of electric vehicles (Ober, 1994)."

Jan 1, 2017

By R. R. Beebe

Marcona Corporation produces about 3.5 million long tons of iron ore pellets annually on its two travelling grate pelletizing machines at San Nicolas, Peru. During 1969, these pellets will be shipped to customers in four marketing areas--Japan, the United States, Europe, and South America. All of these sales are on the open market, and product quality is a major concern. Typical chemical, structural, and quality test results for Marcona pellets are given in Table 1. [ ]

Jan 1, 1969

By R. L. Loofbourow

Most wet mines face unquantified water costs as well as the obvious wing costs. Both kinds of cost are multiplied by work at greater depth and by higher unit prices. This intensified challenge merits renewed or perhaps altogether new response. Any practicable way to reduce inflow tackles both obvious and unquantified costs. Ten groups of possible ways to do it are cited. Most have been demonstrated. Two are known from work other than mining. Surely new means can be found. Here are opportunities for operators of wet mines!

Jan 1, 1980

By Rudolf E. Greuer

Modeling a mine ventilation system is a formidable mathematical problem requiring the solution of twice as many equations as there are airways with half of these being quadratic equations. The addition of a mine fire to the problem adds many more complexities. The U.S. Bureau of Mines and Michigan Technological University attacked this problem in stages. During the 1970's and 80's, a series of computer simulation programs were developed which offered progressively increasing levels of modeling capability and accuracy. The current version of this software, MFIRE 2.0, was released in 1993. This programs provides for full dynamic state mass flow-based modeling of a mine ventilation system, and includes natural ventilation, mechanical ventilation, air reversals and recirculation, methane and fire contaminant spread, and moisture condensation and evaporation in the mass flow and heat exchange calculations. This program is a useful tool for routine ventilation planning and can significantly improve mine safety by enabling detailed quantitative analysis for mine emergency preplanning. This paper reviews the evolution of the MFIRE program from its infancy as a simple digital network simulator replacing analog computers to the present version.

Jan 1, 1994

By C. B. Kettering

In the early operations of the Inspiration Consolidated Copper Company, which extended from 1914 to 1926, the ore as mined contained principally sulphide copper minerals, manly chalcocite (Cu2S), This ore was milled in a 20,000-ton concentrator and the copper was recovered as a flotation concentrate which, was smelted at the nearby plant of the I ternational Smelting and Refining Company. In those days no attempt was made torecovercopper from any oxide minerals present. As mining operations progressed it soon became evident that the proportion of oxide minerals present would steadily increase, and. it was realized that some method of treatment would have to be developed whereby copper could' be recovered from both oxide and. sulphide minerals contained in the ore. A research program was instituted and after some years of experimental and, pilot plant work a new hydro-metallurgical process was developed whereby it became possible to extract copper from both the oxide and. sulphide minerals in a single leaching operation. The solvent adopted consisted of a mixture of sulphuric acid and ferric sulphate,' Following dissolution, the dissolved copper would be recovered by electrolysis. Also in the process of electrolyzing, spent solutions of sulphuric acid, H2SO4, and. ferric sulphate, Fe2(SOk)3, would be regeneratede and. somade available for continuous reuse. The process appeared to be a success and the decision was made to build a full-sized leaching plant for the future treatment of I spiration mixed ore. Construction was started and the leaching plant was completed in 1926, With the start of ferric-sulphate leaching the Inspiration Concentrator was shut down and, with the exception of a few short time runs has not operated since that time. The ferric sulphate leaching process served. Inspiration well for some 30 years and during that time won a name for itself as a classic example of hydro-metallurgical treatment of mixed sulphide oxide ores. Originally the plant was designed. to process some 7,500 tons per day of a mixed ore containing about 1,2% copper, of which 60% was present as oxide and. 40% as sulphide, Down through the years various improvements were made for the purpose of increasing plant capacity and as well to change the limiting 'factor of sulphide-oxide ratio, The final result was a leaching capacity of 9,600 tons per day when treating a 1% copper ore with a reversal of the sulphide-oxide ratio to 60% sulphide and 40% oxide.

Jan 1, 1958

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Z. H. Xie, D. S. He, W. M. Xie

China has rich phosphate rock resources, but with their continuous depletion, the rich ore resources have become increasingly scarce, which makes the exploration of collophane more important. Medium- and low-grade phosphate rock cannot be utilized until their impurities have been removed through mineral processing. However, due to the similar physical and chemical properties of dolomite and phosphorus minerals, the separation of dolomite from phosphate

By John C. Cook

There is pressing need for better means of exploring ground conditions ahead of mining. Exploration drilling has the drawbacks of high cost and a seriously restricted sampling zone. Traditional geophysical exploration methods including gravity, magnetic and resistivity techniques have been used with some success underground, but all give poor target delineation relative to range, since they depend upon minor distortions of a potential force field. Much greater ratios of resolution to range can, in principle, be obtained by employing electromagnetic or seismic radiation as the exploring agent. This is because propagation time delays can be used to measure range accurately, and to separate the small effect of a distant target from the much larger effects of nearer targets and of the transmitter. Seismic waves have long been used with great success in exploration for petroleum, and recent research on their application underground is promising. Electromagnetic (radar) waves have been very successfully used for detection and location through the air and outer space, and have also explored through polar icecaps more than a mile thick.l It is not generally realized that below the soil many rocks are also somewhat radar-transparent, and that radar can be used for exploring such rocks, once a quarry, a mine, a borehole, or other means of access has been opened. Radar waves are reflected by any boundary or object where they encounter a large change in electrical properties.2 The usual cause of such a change is a difference in moisture content. Radar waves passing

Jan 1, 1972

By K. C. Liddell

Chemical processing research made notable advances during 1986. This overview is necessarily brief. It is intended, though, to give the flavor of selected areas where recent activity has been high. At the 1986 Annual Meeting of The Metallurgical Society, a symposium was held on hydrometallurgical reactor design and kinetics. R. G. Bautista, R. J. Wesely, and G. W. Warren edited the resultant proceedings volume. This is the first available volume on this subject. It includes 26 papers on fundamental kinetic studies, modeling of reaction kinetics and reactor performance, and pilot and production-scale operations. Metals discussed include silver, gold, platinum, titanium, cobalt, nickel, zinc, copper, manganese, and iron. Agitated tanks, heaps, dumps, pachucas, pressure auto-claves, electrolytic reactors, liquid membranes, and fluidized beds are among the reactor types covered. An international symposium dealt with iron control in hydrometallurgy. It was sponsored by the Metallurgical Society of CIM, the Institution of Mining and Metallurgy, The Metallurgical Society of AIME, and Gesellschaft Deutscher Metallhutten and Bergleute. J. E. Dutrizac and A. J. Monhemius edited a proceedings volume containing 41 papers. These covered many areas of process chemistry, solvent extraction, precipitation, treatment of pickle liquors, impurities, residues and the environment, and process selection. Precious metals Concerning precious metals processing, the US Bureau of Mines IC 9059 is noteworthy. It includes papers presented at a USBM briefing at the Western Mining Conference in Denver. Topics covered included ion exchange, staged heap-leaching direct-electrowinning, and mercury precipitation during cyanide leaching of gold ores. Also dealt with were carbonaceous gold ores, carbon adsorption and desorption, heap leaching, the carbon-in-pulp process, and precious metals recovery from electronic scrap and solder. R. C. Sandberg and J. L. Huiatt (Journal of Metals, June 1986, and USBM RI 9022) developed a method to recover silver, gold, and lead from a complex sulfide using ferric chloride, thiourea, and brine-leach solutions. Gallium Gallium recovery was also a subject of considerable interest. Much information on this topic is proprietary. But two papers described solvent extraction of gallium. V. P. Judin and R. G. Bautista (Metallurgical Transactions B, 1986) developed an equilibrium model to separate gallium chloride from aluminum chloride. Tributyl phosphate was the extractant studied. T. Sato and H. Oishi (Hydrometallurgy, 16, 1986) investigated gallium extraction from sodium hydroxide by using Kelex 100. Data were given on the equilibrium distribution and the extraction kinetics. Galena Interest continues in recovering lead by hydrometallurgical processing of galena. J. E. Dutrizac; S. H. Kim, H. Henein, and G. W. Warren; and M. C. Fuerstenau et al. (Metallurgical Transactions B, 1986) all investigated leaching of PbS by ferric chloride. Dutrizac reported parabolic kinetics and reaction control by outward diffusion of lead through a porous layer of elemental sulfur. Kim, Henein, and Warren, however, reported that two leaching reactions occur. A nonoxidative reaction produces H2S and there is an oxidative leaching by ferric ion. Acid dissolution was found to predominate when the acid activity to ferric activity ratio is high. Surface chemical reaction was reported to be controlling. Fuerstenau et al. also reported that mass transfer through the sulfur layer is rate determining. But they believe that the species transferred are chloro complexes of ferric ion. In other work on lead, A. Y. Lee, A. M. Wethington, and E. R. Cole, Jr. of the USBM described an environmentally acceptable hydrometallurgical alternative to the smelting of lead concentrates (RI 9055). Uranium In spite of depressed prices, much work has been reported on uranium processing. R. G. L. McCready, D. Wadden, and A. March-bank (Hydrometallurgy 17, 1986) described the nutrient needs for in-place leaching by T. ferrooxidans. They reported the optimal conditions for uranium solubilization. L. E. Eary, H. L. Barnes, and L. M. Cathles (Metallurgical Transactions B, 1986) carried out an experimental and modeling study of uraninite dissolution. They concluded that ferric ion preferentially leaches uraninite in pyritic ores. Hydrogen peroxide was found to be less selective. A. Vuorinen, P. Hiltunen, and O. H. Tuovinen (Hydrometallurgy 15, 1986) studied redox and precipitation reactions of iron and uranium in leach liquors. P. T. Chiang (Hydrometallurgy 17, 1986) reported on the effect of uranium loading in the DEPA-TOPO process to separate uranium and iron from wet-process phosphoric acid. F. J. Hurst (Hydrometallurgy 16, 1986) conducted a fundamental study of the separation of uranium from phosphoric acid by DEPA and TOPO. D. E. Chia and W. C. Cooper (Hydrometallurgy 16, 1986) reported on bench and pilot scale work to recover uranium by the HIMIX process. They reported that the acid consumption to produce an eluate suitable for yellowcake

Jan 5, 1987

By Robert M. Abrams

The passage of an ore pulp through the froth flotation process results in the production, after dewatering, of a finely divided concentrate. This intermediate product Trust be handled to the subsequent process facilities via belt or other types of conveyors, hoppers, bins, chutes and feeders, and the like, Flotation concentrate has some difficult handling properties such as abrasiveness, sluggishness when moist, or fluidity when dry, These characteristics require the judicious selection or design of handling, equipment and the careful engineering of transfer chutes. It is to the special problems which are encountered in this in-plant handling procedure that this paper is addressed. The methods which will be discussed have beer. developed primarily for the handling of the froth flotation concentrates of copper, leap, and taconite iron ores, but may be generalized, with appropriate modifications, to apply to other similar metallic concentrate&. The controlled feeding of concentrates frog: bins and hoppers is covered alone with some of the techniques used to transfer concentrates from ore belt conveyor to another. Methods are described for unloading belt conveyors into bins or storage tiles, The reclamation from storage piles is also treated.

Jan 1, 1976

By C. Bryce

Rhenium is primarily recovered as a byproduct from molybdenum roasting circuits. However, rhenium is also present in copper heap leach solutions at very low concentrations, and this represents a potential new rhenium source. Freeport-McMoRan (FCX) has developed a new process using carbon adsorption followed by ion exchange or solvent extraction to concentrate rhenium from less than 1 ppm to 15 g/L. The product is compatible as a feed to an existing rhenium processing plant. A pilot plant operated at Freeport-McMoRan?s Sierrita Mine has demonstrated the technical and financial feasibility of this new process. Copper heap leaching dissolves more than just copper from the copper ore. Several metals can be identified in leach solutions at concentrations of financial interest, rhenium in particular. Rhenium spot prices ranged from a yearly average of $1,070/kg ($483/lb) in 2005 (Magyar, 2008) to $10,400/kg ($4,706/lb) in 2008 (Polyak, 2013). The spot price in September 2013 was approximately $3,000/kg ($1,357/lb) of rhenium (Metal Bulletin, 2013). Several facilities at Freeport-McMoRan have very low rhenium concentrations in their copper leach solutions, ranging from 0.1 to 0.7 mg/L. The large quantity of solution available, if coupled with an attractive rhenium price, means that the rhenium contained in leach solutions may represent a significant value if a method of extraction could be developed. Testing has confirmed that rhenium adsorbs onto activated carbon much more selectively than other metals in the pregnant leach solution (PLS). A 100-gpm pilot plant was constructed to verify the lab-scale results on a larger basis and establish operating parameters for a full-scale operation. This report summarizes the initial technology development and pilot plant operation.