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By Andrew S. Lain

Before you begin mining aggregates in Oklahoma, you must first look at the market and ask these questions: (1) Is there a need for additional aggregate production? (2) Can we produce the aggregate efficiently enough to compete? (3) Is the deposit close enough to the market? If you can answer "yes" to each of these three questions, then you are ready to start the process for mining aggregates in Oklahoma. The first consideration to be made is whether the aggregates produced will need to meet state specifications. If you do not want to meet state durability standards, there are numerous formations you can mine and produce aggregates. We will deal only with the formations that will meet state durability standards. These formations are located in the southern half of Oklahoma and in the Tulsa area. Some of the formations that contain limestone, dolomite, and sandstone that will meet state durability tests are West Springs Creek, Kindblade, McKenzie Hill, Butterly-Royer, and Rush Springs Formations. Granite and rhyolite also are found in Oklahoma and can be used as construction aggregates. Once an area has been identified as having durable material, and drilling has verified its properties, a permit must be secured from the Oklahoma Department of Mines, Oklahoma Department of Environmental Quality, and any other applicable agency before construction of a plant occurs. There are many different brands of crushers available, but all work on one of two crushing principles-either compression breakage or impact breakage. Each type has its application in different industrial-mineral types and the desired end product. A typical small crushing plant would start with a jaw crusher for primary reduction, and then employ one standard cone and one shorthead cone for secondary reduction. One grizzly would be in the circuit with one finished-product screen and three to four finished-product stackers. Natural sands and gravels are found in rivers and old river channels. These too need to be tested to see if spec product can be produced. All of our sand operations have very little gravel; thus, we do not process the gravel. The sand is processed through classifying tanks for gradation.

Jan 1, 1999

By L. B. Sand

Zeolite ores and protores occur in extensive deposits in the western United States. A recent paper1 describes the general geology and mineralogy of these deposits and their geographical distribution. The zeolites are alkali-and silica-rich varieties of mordenite, erionite, chabazite, phillipsite, ferrierite, and clinoptilolite. The zeolites were formed primarily by the alteration of Tertiary pyroclastics and in some cases by the alteration of Tertiary siliceous flows. The altered pyroclastics are of principal economic significance. The volcanic ash and tuff were deposited in restricted basins where saline lakes developed. The pyroclastic layers often are of uniform thickness and macrotexture for miles from the center of the basin and interfinger with detritals toward the edge of the basin. Thus, in addition to unaltered volcanic glass, the gangue minerals in these deposits include some combination of primary minerals in the pyroclastics, alteration products of the glass other than the zeolite of interest, and detrital minerals. The gangue minerals include quartz, cristobalite, tridymite, opal, orthoclase, micro cline, montmorillonite, hornblende, calcite, gypsum, thenardite, and iron oxides. The presence of one or two other zeolites in the layer in some cases also is considered as gangue. There are no quantitative mineralogical descriptions of these deposits in the literature; this: study was undertaken to obtain this information. As the zeolites occur as single crystals, crystal aggregates, zeolite-glass particles and in cemented aggregates, a method of study was chosen which would give the size distribution of the zeolite in these types of association. For this reason dispersion of the rock was achieved with a minimum of grinding action. This excluded those zeolite ores, silicified by hydrothermal altera¬tion which require grinding. A relative index of the ease of disaggregation of the zeolite ores thus is obtained, but the technique does not give absolute size distributions of the zeolite crystals. Samples were selected for analysis which do not necessarily represent the best ore horizon, but do represent a spectrum of zeolite occurrences from significant deposits. They are mordenite, erionite, and phillipsite from Rome, Oregon, clinoptilolite and chabazite from Reese River, Nevada, chabazite (and var. herschelite) from Bowie, Arizona, erionite from Jersey Valley, Nevada, and erionite from Pine Valley, Navada.

Jan 1, 1967

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 S. Upadhyay, A. M. Afrapoli

"Fluctuation in market price of mineral products has enforced mining companies to cut down their operating costs. One useful way of reducing operating costs in mining operations is to make optimal operational decisions, incorporating efficient mining fleet management systems (FMS). Over the last 50 years, several types of FMS have been developed and introduced in the mining systems around the world. However, there are two major shortcomings in existing systems. Firstly, none of the existing systems consider effects of the short-term plan on dynamic truck assignment. Secondly, most of the existing FMSs make decisions in a way that optimize a specific objective while ignoring others. Herein, we propose a multi stage and multi objective FMS that deals with aforementioned shortcomings of the currently available systems. The developed FMS, in its upper stage, links semi-dynamic operational decision making process to the short-term production plan by simultaneously assigning shovels to mining faces and allocating trucks to the shovels by implementing its first multi objective decision making model. Then in its second stage, in a real time dynamic decision making process, it assigns trucks to shovels and corresponding destinations by optimizing its second multi objective decision making model. The developed fleet management system has been verified using an Iron ore case study and the results are presented in this paper. INTRODUCTION Open pit mining method is the most common method to mine mineral from the earth. Based on Wetherelt and van der Wielen (2011) there exist 2500 open pit metal mines that are categorized as industrial scale mines. Over all of those, the amount of material handled using truck and shovel material handling system is more than the amount handled by all other systems together (Humphrey & Wagner, 2011).It has been proven by several studies that more than 50 percent of operating cost in a truck and shovel based open pit mining operation is accounted for material handling (Moradi Afrapoli & Askari-Nasab, 2017, Alarie & Gamache, 2002, Ahangaran et al., 2012). Thus, introducing even 1 to 2 percent cut in its material handling cost saves huge amount of money for the stakeholders. Over the last 50 years, two major areas have been identified to cut this cost. On one hand mining industries have been implementing equipment with higher production capacities, and on the other hand, researchers have been developing operations research techniques to make dynamic decisions as close to optimal as possible to deduct operational costs in truck and shovel based mining operations."

Jan 1, 2018

By O. J. Malacarne

The use of a series of thickeners for-washing of the slime fraction of a leached ore is not uncommon in the uranium industry; however, few ore processing mills have used a thickener circuit for washing leached total ore. The, common, practice is to wash the sand fraction in classifiers and, the slime fraction only in thickeners. At the Uravan mill of Union Carbide Nuclear, Company, a total ore' acid leach is employed to extract uranium values from sandstone ores. These leached ores are then washed at a nominal rate of from 1,000 to 1,300 tons per day in a series of washing thickeners. A typical screen. analysis of the ore being washed in these thickeners, is 25 percent -200 mesh, 3 percent +35 mesh and about, one' percent +l4 mesh., The amount of slimes (-200 mesh) will vary from 20 to 30 percent s the mill receives ore from 150 different mines The washing liquor is acidic, from a pH of about 2 in the first thickener to 0.6 pH in the last thickener and of course ally acid duty equipment must be used. The Uravan mill of Union Carbide Nuclear Company is located in Montrose County, Colorado and is physically located on a canyon wall along the San Miguel river. The thickeners referred to in this, paper are located on top of the mesa above the canyon. For purposes of identification, these thickeners are numbered 1 through 8 from south to north.

Jan 1, 1962

By Peter Jackson

Over a quarter of a century ago a New Zealand physicist suggested the possibility of detecting coal seam discontinuities by a seismic method carried out underground in the mine workings (Evison 1955). Twenty years ago initial experiments were performed in Germany (Krey 1963). Since then research has continued, notably in Britain and Germany, in order to provide a reliable fault and washout detection system. The National Coal Board of Great Britain has now carried out more than 250 surveys in Britain, France and Australia. The first NCB survey in the U.S., a successful application of the technique to an Ohio coal seam, will be described and the strengths and limitations of the method explained.

Jan 1, 1984

By M. D. Ross

Complex combinations of technical, economical, safety and environmental factors are causing mining engineers everywhere to review their operations. Naturally, some of the traditional methods still have a large role to play because they are, in certain circumstances still adequate and are well understood. But, as circumstances change, a new approach is required. The combination of technical and economical reasons for introducing Long- wall mining in some seams in the United States is associated with strata control. Other reasons include simplifications and improvement of ventilation, reduced coal dust, more output from each unit and crew, increased percentage of recovery from each area of the coal deposit, and reduction of rock dusting. The innovations which transformed Longwall mining from a cyclic to a continuous mining system were the armored conveyor and power loading machines, but the crucial breakthrough for application to high production was the development of the powered roof support, which at once gave the system the advantages of mastery of strata control, increased rate of advance, significant reduction of manpower, and increased safety.

Jan 1, 1974

By E. H. Cho

Pittsburgh #8 coal cleaned by the conventional froth flotation was oxidized in a 2-liter Parr Stirred Top Bench Reactor. The oxidation time was varied up to 5 hours; the temperature, from 150 to 225°C, the oxygen pressure, from 15 to 500 psi; and coal dosage, from 15 to 60 grams in 600 ml of distilled water. The oxidation of coal was evaluated by measuring total acidic group formed on the coal surface as well as by measuring the adsorption of calcium ion from a solution under standard conditions. It was found that the oxidation increased with increasing oxygen pressure without much leveling-off effect. However, the oxidation of coal increased sharply from 150°C to 175°C and then tended to level off above 175°C. The adsorption of calcium was reached more than 3% on the coal oxidized under favorable conditions. The adsorption of calcium on coal increased sharply with increasing equilibrium pH. The BTU loss was sensitive to oxidation temperature. This study concludes that the optimum oxidation of coal should be conducted at high oxygen pressures around (e.g. 500 psi) 170°C for 3 hours. This study also concludes that the oxidation in an autoclave is far superior to that in a fluidized bed in terms of calcium adsorption, BTU loss and reaction time.

Jan 1, 1994

By Alex Stewart

Most iron bearing ores treated by the R=N Direct Reduction Process yield products lower in phosphorus than the initial ores. The sulfur is also removed and generally is in the range of 0.02 to 0.05 percent in the final R=N product, independent of the sulfur content of both the ore and carbon fuel used. The mechanism and thermodynamic justification of the principal reactions leading to the removal of both phosphorus and sulfur and the control of sulfur in the iron concentrate produced are discussed and supported by the results of laboratory experiments. In substantially all iron ores (containing up to 4% sulfur as sulfate) and solid carbon fuels (containing sulfur from 1 to 8%) tested, the rejection of sulfur has been in the order of 90-95%, with not more than 0.02 0.05% sulfur in the final metallic iron product. With respect to phosphorus-bearing iron ores ranging up to 0.70% P205, the rejection to tailings has been in the order of 70-95%, resulting in not more than 0.03 - 0.07% phosphorus in the final product.

Jan 1, 1958

By A. Pomianowski

A spectrophotometric study of xanthate solutions alone and their mixtures with equimolar amounts of alkyl trimethyl ammonium bromide (CnTAB) has established the additivity of ultraviolet spectra (190 - 400 my) for potassium ethyl xanthate (KEtX) and C12TAB. No obvious effect of C12TAB addition on the decomposition rate of KEtX was observed. The kinetics of collector abstraction by copper, nickel and sphalerite powders were studied in 10-5 - 10-3M solutions of KEtX alone and in equimolar mixtures of KEtX and C12TAB. Presence of C12TAB caused a 2 - 3 fold increase in the amounts of xanthate (equivalent to many monolayers) abstracted by copper and, nickel powders but had no effect on the abstraction of xanthate by unactivated sphalerite. The adsorption of C12TAB itself from the mixed solutions was very low, at most ore monolayer only, and was not influenced by the nature of the solid powder used. The results are interpreted in terms of molecular associations and synergistic action of xanthates and quaternary amines.

Jan 1, 1963

By Keith G. Wallace

In 1982 the U.S. Congress passed the Nuclear Waste Policy Act (NWPA) to provide for the development of an underground repository for spent nuclear fuel. This development will be managed by the United States Department of Energy. In 1986, the President selected three areas for site characterization to determine their suitability for the development of an underground repository; those sites were: 1) A site in volcanic tuff located at Yucca Mountain in Nevada, 2) a site in bedded salt located in Deaf Smith County in Texas, and 3) a site in basalt located in Hanford, Washington. At present conceptual repository designs are being developed for each site. A key element of a repository design is the underground ventilation system required to support construction, nuclear waste emplacement, and potential waste retrieval. This paper describes the preliminary ventilation system designed for the repository in tuff. The concept provides separate ventilation systems for the construction and waste emplacement activities. The paper further describes the means by which acceptable environmental conditions will be re-established to allow re-entry into previously closed rooms for the purpose of inspection, maintenance or retrieval.

Jan 1, 1987

By Michael M. Kimberley

Gold is sufficiently valuable that concentrations as small as 1 p.p.m. may support low-cost mining. Mindful of this, Sunoco Energy Development Company (Sunedco) obtained fire-assay analyses on some sulfide-rich sandstone encountered during routine drilling. Initial fire-assay values from independent laboratories indicated some significant gold values and independent analysis by atomic-absorption spectroscopy also indicated detectable gold at economically interesting concentrations. Sunoco geologists realized the potential for spurious results inherent in both fire assaying and atomic-absorption spectroscopy. Fire assaying that relies on gold measurement by weight requires perfect separation of gold from all other elements, Imperfect separation may artificially enhance apparent gold concentrations. Atomic-absorption-spectroscopy (AAS) measurements are potentially subject to interferences from other elements. For example, Beamish and Van Loon (1977, p.4) report AAS measurements of 4 p.p.m. Au to be characteristic of completely barren soil samples when samples are prepared by standard dissolution in HC1 to form solutions of about 10,000 p.p.m. total cations (mostly Na, Ca, Mg, Al, and Fe) .

Jan 1, 1981

By Greg Sanders, Michael Gilbert, James G. Shaughnessy

"A new raw water intake from the South Holston River to the Kingsport water treatment plant is under construction. The exploration for the 335 meter long horseshoe shaped tunnel design took advantage of the site topography along the alignment to better define the complex geology consisting of Knox Limestone and Sevier Shale formations separated by a fault zone. To assess the geological conditions and to determine the extent of the fault zone, a 183 meter long horizontal boring and geophysical investigation was performed along a portion of the tunnel’s horizontal alignment. Explorations defined the fault zone width at approximately 5 meters and packer testing along the alignment provided a means of estimating groundwater inflows. As the tunnel excavation advances, comparisons of exploratory data versus encountered conditions will be made. This paper provides details on the geotechnical investigation, lessons learned, and the influence of the geology on the tunnel design. PROJECT OVERVIEW The Kingsport WTP is supplied with raw water via the South Holston River. The City of Kingsport (City) owns and operates the Kingsport Water Treatment Plant (WTP). The existing raw water intake system, consisting of pumps, power supply, raw water piping and appurtenant facilities has been determined to be vulnerable to failure. The original Raw Water Pump Station (RWPS) was constructed in 1928 as part of the original Kingsport WTP project. The 1928 RWPS operated until 1970 when it was replaced by the existing RWPS. This existing system still utilizes the 1928 piping and piping tunnel. Improvements to the RWPS hydraulics and capacity are required to meet the City’s future water treatment needs. The system extracts water from the South Fork of the Holsten River and coveys it via three pressure pipelines to the WTP. The conveyance piping exits the RWPS, ascends approximately 23 m up a near vertical rock outcropping then enters a 120 m long tunnel. Upon exiting the tunnel the piping continues underground to the WTP. The drill and blast tunnel is horseshoe shaped and nominally 3.3 m high by 3.3 m wide. The easternmost 62.3 m of the tunnel is without any type of support. The remainder is lined with concrete. According to the City of Kingsport, the concrete lining was installed in the 1970’s as part of upgrades to the facility. An evaluation of the current facilities concluded that the conveyance piping was in poor condition and required replacement."

Jan 1, 2016

By John E. Good

A new coal boom is under way in western Canada. As recently as the mid-1960's, the coal industry in western Canada was considered stagnant and, by some people, dying. But Japan's growing need for coking coal, which became more pronounced in 1966, sparked great interest in western Canadian reserves.' As a result of a rapidly expanding export market and growth in domestic demand, Canadian coal production spurted to 16.6 million tons in 1970. Production in 1965 was 11.5 million tons. Today, about 60 percent of Canada's output comes from western Canada, but by 1975, 87 percent (23.1 million tons) is expected to come from this area, and 96 percent (68.3 million tons) by 1990. Opening, developing and operating new coal mines to meet the needs of a rapidly expanding market for western Canadian metallurgical coals pose some unique and difficult problems. Covering nearly every aspect of opening properties in remote areas, the problems are met in exploration, transportation, construction, mine operations, coal preparation, manpower and ecology. Paul Weir Company has been involved as coal consultants on various projects in western Canada for some years. In presenting this paper, we make no attempt to present an in-depth or detailed analysis of any one specific situation. Each has, or would have, its own peculiarities or characteristics, both physical and economic. Instead, the objective is to provide for those not intimately

Jan 1, 1972

By N. W. Johnson

Two analysis tools which have been applied to problems in industrial Flotation processses are the application of kinetics and detailed mineralogical or liberation data, both on a size-by-size basis. The paper describes practical examples of the application of kinetics to examination of the entrainment recovery mechanism, the retardation of valuable minerals at the commencement of flotation sections and the resolution of problems related to valuable species with reduced rate constants at the end of sections. Examples of the usefulness of size-by-size liberation data for monthly composite samples of products from a concentrator are also described.

Jan 1, 1985

By J. J. Sammarco

Computer-aided control of a mining machine pl aces its operator in a safe, remote location, thus reducing health and safety risks, Sensing systems that provide machine position (xy coordinate) and heading (yaw), are extremely useful in aiding remote positioning of mining machines. The U. S. Bureau of Mines has developed two sensing systems for position and heading information: A scanning laser system based on the Lasernet sensor and a mechanical position and heading system based on linear-position transducers. The Bureau is also investigating a system from Honeywell, the modular azimuth and positioning system that employs ring-laser gyroscopes. This paper presents a brief background and status of these sensing systems as applied to the remote positioning tasks of a continuous mining machine. Included are field test results of each system operating in a room-and-pillar research section at a cooperating mine in West Virginia.

Jan 1, 1992

By C. D. Broadbent

Open pit blasting can be a low cost routine or a high cost bottleneck depending on geology, environment and the operator's ability to master site conditions. Because blasting is a repetitive operation, there should be ample opportunity to optimize the process; but ordinarily this optimization is rarely achieved and then only if conditions remain consistent and predictable. Often, however, and particularly with porphyry coppers, hidden differences in rock characteristics nullify the operator's best efforts. At a porphyry copper deposit being mined by Kennecott Copper Corporation, there are all of the following rock types: oxidized and fresh porphyry, altered and silicified limestone, explosion breccia, rhyolite, sandstone, shale and several others. The units can cover large areas or be limited to a few tens of feet in maximum dimension. They are rarely visible on the bench surface and often do not show up in the bench slope. Even though the blastability of these rocks varies substantially, a generally-uniform blasting pattern was formerly used with disappointing results. Secondary blasting was often required and high bottoms occurred frequently; overblasting wasted drilling time, and explosives caused excessive throw. Several possible methods for improving the blast design were considered and a geophysical method was chosen for testing. There had been reports in the journals that seismic methods were being used in Europe as aids for blasting design. A literature search turned up work by Haxby 1/ on use of a refraction seismograph for measuring pit bench rock strength. Using his work as a springboard, equipment was purchased and a program was planned and carried out for developing, evaluating and implementing the seismic method for routine mine use. Following are discussions of the development program and results.

Jan 1, 1974

By M. M. Fine

One of the most persistent topics at technical gatherings and in the recent scientific press has been the reduction of iron ore. To cite some of them, in May 1960, a symposium on the subject was sponsored by the Electrochemical Society. The proceedings, comprising two dozen papers, were later published.3/ More recently, November 1966, a similar gathering took place at Michigan Technological University, Houghton, Mich. The latest meeting was the International Congress on Production and Utilization of Reduced Materials, Evian, France, May 1967. If the volume of literature, symposia and conferences is a measure of industry concern, then current interest must be at a high level. To the casual observer of the metallurgical scene, the abiding and recurring attention to iron ore reduction outside the blast furnace must be somewhat puzzling in that so much research endeavor has resulted in so little commercial activity.

Jan 1, 1968

By Charles S. Spirakis

The consistent occurrence of organic carbon early in the paragenesis of Mississippi Valley-type sulfide deposits suggests a genetic link between organic carbon and the sulfide ores. Reduction of sulfate in the mineralizing solution by organic carbon at the sites of mineralization is commonly proposed as the precipitation mechanism for these deposits, but inorganic sulfate reduction is too slow to be important to the chemistry of the ore zone, and bacteria, which catalyze sulfate reduction at low temperatures, are not known to survive at the temperatures of mineralization. Sulfate reduction with or without bacteria is not consistent with the sulfur Isotope evidence. Sulfur isotope studies also indicate that the derivation of sulfur from hydrocarbon at the sites of mineralization is unlikely. Organometallic complexing may be involved in the transport of metals, but no plausible mechanism for breaking these complexes at the sites of mineralization has been proposed. Reduction of partly oxidized sulfur species in the mineralizing solution by organic carbon at the sites of mineralization is a possibility which deserves further scrutiny.

Jan 1, 2013

By D. Naden

Sulphuric acid leaching is the most common method of recovering uranium from its ores and in the following discussion various alternative low cost routes are investigated for recovering uranium from milled ore acid leach liquors. Historically uranium has been recovered from dilute impure leach liquors by packed bed ion exchange (IX), solvent extraction (SX) or a combination of both. All three process routes as commonly practised require complete separation of leached solids and liquids, as well as extensive solids washing in order to reduce soluble uranium losses to a minimum. Separation and washing are costly and often require equipment covering large areas of land. The size and capital cost of the post leach solids washing operation is controlled by the ease of washing the slimes fraction of the ore and the key to the problem of cost reduction, as many others have realised, lies in the ability to separate the slimes fraction of the ore from the more easily washed sands, to wash the sands in low cost equipment and to recover uranium from the slimes stream. Sands and slimes are usually defined in this context as the material larger than, and smaller than, 325 mesh (44 microns). The proposed flowsheets which minimise solid-liquid separation and washing have usually incorporated ion exchange (Refs 1,2,3), and resin in pulp is in fact practised in several mills in the USA (Ref 1). The resins used in these flowsheets and [cranium] recovery throughout the world have been strong base resins which load rapidly and are moderately selective for the uranyl sulphate complex anion over the range of pH which is conveniently achieved in an acid leach operation. Strong base resins are, however, not as selective as weak base resins and weak base liquid reagents owing to the adsorption of competing anions, some of which find their way into the product while others act as resins poisons. Ion exchange using strong base resins must therefore be followed by a more selective weak base SX purification step if pure yellow cake is required.

Jan 1, 1976