Search Documents

By H. M. Garrett

Dr. Ft: L. Handy of Iowa State University's Engineering Experiment Station very handily reduced the relatively complex cement making process to ultimate simplicity: "Take two cups of crushed limestone; add one-half cup of clay or pulverized shale (plus perhaps some sandstone or iron ore; depending); mix thoroughly and grind up fine. Bake in white-hot oven at about 2,600 F. Cool. Add a tablespoon of gypsum: and, again, grind very fine. If you can do this for about a penny a pound, you can almost compete.' Before Dr. Handy's time, is the early days of the U. S. Cement Industry, from 1817 when the Erie Canal was started to 1895 when the first rotary kiln was built, this was the recipe, more or less. Dr. Handy's raw materials were deposited in a natural "cement rock" found in the Hudson, New York and Lehigh, Pennsylvania valleys. At first the natural rock was pulverized (not ground fine) and interlayed with wood for "baking" at about 1,800° F in a batch process vertical shaft kiln or beehive oven; then the calcined pozzalonic (lime and silica) product was cooled and ground fine for shipment in wooden barrels. Later, in 1871, true portland cement was made. The same natural rock was ground fine to make briquettes; interlaced with wood or coal in the "Beehives"; and baked at a clinkering temperature of about 2,6000 F to produce clinker (calcium silicates) which was ground fine for sale. In those days, the crude cements dropped in price from over 1 cent to 1/2 cent per pound. In 1895 the rotary kiln introduced-continuous instead of batch processing. The countercurrent process allowed the same raw materials (now finely ground to about 80% minus 200 mesh like face powder) to be fed at one end of the rotary kiln. Heat from a coal burner was introduced with combustion air at the other end to produce the requisite clinkering temperature. The clinker product was discharged for atmospheric cooling, mixed with gypsum, and again ground fine as face powder. The cement was sold primarily in barrels, later in jute or cloth bags, then paper bags, and now primarily in bulk truck and barge carriers. The price dropped to a 1/4 cent per pound by 1910, but with recent inflation it now sells for 2 cents per pound.

Jan 1, 1978

By S Dunbar, S Harrison, J Cilliers

‘Those who cannot remember the past are condemned to repeat it.’George Santayana (16 December 1863 in Madrid, Spain to 26 September 1952 in Rome, Italy). Philosopher, essayist, poet and novelist.The minerals industry requires specific technical expertise in mining and mineral process engineering, geoscience, environmental science and social sciences to maintain and operate its current business and to develop future projects. A steady supply of new graduates is essential to replace not only retirees, but also those who left the industry during the recent downturn. The supply of trained graduates is however not instantaneous, and there is a four-year lag from entry into university to graduation. It is essential that this lag in human resource supply and a likely shortage of knowledgeable, qualified talent is taken into account when considering risks to the business of mining. The shortage of skilled professionals has in the past resulted in unsustainable salary expectations and is likely to pose financial and execution risk to new projects. This paper will review some recent statistics of students in training from around the world and explore what action is required to not only sustain the training courses that provide future ready talent to the mining industry, but also ensure that graduates from these courses are equipped to create change and be leaders in the industry. The future prosperity of the industry is dependent on delivering sustainable stakeholder returns through the cycle. This will be achieved by developing capable individuals that can respond to the complex, adaptive challenges facing the industry.*This is an abstract only. No full paper was prepared for this abstract.*CITATION:Plint, N, Cilliers, J, Dunbar, S and Harrison, S, 2018. Back to basics: resourcing the mining industry, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 33–34 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By V. I. Ryaboy, V. P. Kretov, I. V. Ryaboy

Attachment of collectors containing thioamide groups to sulfide minerals is interpreted as interaction between NH-acids and sulfides acting as solid bases with the formation of a donor–acceptor reagent–mineral bond. As the acid function of the reagents increases, their sorption optimum is observed to shift to a less alkaline region. The developed concept has enabled us to propose a method for calculating optimal pH values for the sorption of reagents on mineral surfaces depending on the acidity of the reagents and the pH value of the medium. Our estimates correlate well with the experimental data. We studied the spectra of the reagents containing thioamide groups and the spectra of copper salts and their surface compounds on an oxidised copper film, chalcopyrite, and pyrite using the infrared (IR) spectroscopy. Some effective applications are described of reagents containing thioamide groups in the flotation of copper, copper-molybdenum, and gold ores, both separately and in combination with other collectors. It is shown that the flotation performance is maximised by adding two- and three-component collector combinations, including collectors containing thioamide groups. In this case, the increase in the recovery of copper, molybdenum, and gold is up to 2-5% compared to conventional process conditions. A new, high-performing reagent containing a thioamide group is proposed. A commercial pilot batch of the reagent was produced and sent for flotation tests in gold ore applications. Keywords: Thioamide reagents, dialkyldithiophosphate, collectors, flotation, arsenopyrite, chalcopyrite

Jan 1, 2020

By K. BRUCE DAVIDSON

-The North Central Texas area includes the following counties in Railroad Commission of Texas District 7-B: Brown, Callahan, Coleman, Comanche, Coryell, Eastland, Erath, Fisher, Hamilton, Haskell, Hood, Jones, Lampasas, Mills, Nolan, Palo Pinto, Parker, S

Jan 1, 1946

By W. J. Dolinar

The U.S. Bureau of Mines implemented the Fe2+ /Fe3+-SO2 anode reaction in a Cu electrowinning cell with full-size electrodes. Electrowinning was conducted using industrial electrolyte at 258 A/m2 (24 A/ft2) and 40 'C. Manifold injection of electrolyte provided circulation between the electrodes. Five-day operation using an optimized manifold gave an average cell voltage of 0.94 V at 89% current efficiency, which amounted to an energy savings of 2.82 MJ/kg (0.355 kW-hr/lb) compared to conventional electrowinning. Acid misting was completely eliminated, and SO2 concentration 23 cm (9 in.) above the cell averaged 0.06 ppm.

Jan 1, 1995

By H. Kolcz, G. S. Dobby, R. Stratton-Crawley, S. W. Wilson

"The Matte Separation plant at Inco's Copper Cliff Smelter Complex is described. The continuing implementation of flotation column technology in the flowsheet in various applications is reviewed. Comparative studies on 0.038, 0.10, 1.1 and 1.8 m diameter columns have been carried out. It was demonstrated that consistent metallurgical performance was obtained with both pilot- and plant-scale units. Grade-recovery relationships for plant columns can be predicted with confidence from data obtained from test equipment. The work indicates further areas of the flowsheet in which columns present an attractive alternative to conventional cells. INCO’s MATTE SEPARATION PROCESSNickel concentrate from Inco’s Sudbury area mills has been processed in the Copper Cliff smelter for sixty years. Prior to that, Sudbury ores had been smelted directly. Historically, copper-nickel separation from smelter converter matte was achieved by the classical Orford process. In the mid-1940s Inco developed an improved separation scheme based on mineral processing methods, the Matte Separation Process. A generalized Copper Cliff Smelter Complex flowsheet is shown in Figure 1.Iron sulfide and gangue contained in the nickel concentrate are rejected in the smelting-converting process. The molten product, typically containing 50% Ni, 25% CU and 22% s, is cast at about 1000 ºC, into floor moulds and allowed to slowly cool for four days. During this period nickel sulfide (heazlewoodite, Ni,S,), copper sulfide (chalcocite, cu,S) and a copper-nickel metallic alloy precipitate and form a coarse grain structure that is amenable to separation by physical means. The phase chemistry of this system has been described previously by Sproule et al (1). This matte then becomes the feed to the Matte Separation Process.A simplified block flowsheet of the Matte Separation Process is shown in Figure 2. Matte ingots, typically weighing 20 tonnes each, are broken and crushed in a three stage crushing plant to approximately minus 0.01 m and conveyed to the Matte Separation building. Grinding is carried out in two parallel circuits each consisting of a rod and ball mill in series producing a flotation feed which is 65% minus 44 microns. Typical throughput of each grinding circuit is 650 tonnes/day. The metallic fraction of the feed is recovered by magnetic separation. This product, containing about 66% Ni and 15% CU, is a major component of the feed to the Copper Cliff Nickel Refinery."

Jan 1, 1990

By H. F. MILLE, Charles A. Kumke

This paper describes mining operations at the Iron King mine of the Shattuck -Denn Mining Corp. in Yavapai County , Ariz . It is one of a series being prepared by the Mining Division of the Bureau of Mines on mining practices and methods in various mining districts of the United States . The mine and surface plant of the Iron King mine are situated in secs . 15 and 21 , T. 13 N. , R. 1 E. , about 20 miles by State highway 69 east by southeast from the city of Prescott and 1 mile southwest of the mining town of Humboldt on the Prescott -Mayer branch of the Atchison , Topeka & Santa Fe Railroad (fig . 1 ) . A l-mile truck haul brings the mill concentrates (lead and zinc ) to bins on a siding of this line at Humboldt ; the silver -gold bullion is handled separately . Some of the oldest residents in the vicinity of Prescott claim that the original location for the Iron King mine was made in 1880. The earliest recorded production was in 1906-1907 . During the first World War the mine produced fluxing ore for the Humboldt smelter . In 1936 the property was purchased under tax sale , and a mill (140-ton capacity ) was erected in 1938. The capacity of this mill was increased to 225 tons the following year , and a cyanide plant was added in 1940 to treat mill tailings . In 1942 the Shattuck -Denn Mining Corp. purchased the property . A new working shaft was sunk , mine development was speeded , and mill capacity was again increased . The predominating locally formation of this district is a pre -Cambrian schist known as Yavapai schist . The Iron King deposit , which carries gold, silver , lead , zinc , and copper , occurs as a series of lenses in a sheared zone of the pre -Cambrian schist . Down to the 600 -foot level the ore was mined and drawn by the shrinkage method . The cut -and -fill method was adopted for the lower levels . Waste rock for stope fill was provided by block-caving an area on the 600 level . Main haulage drifts were driven in the firm footwall , and crosscuts were driven across the mineralized zone at 100 -foot intervals . Grab samples taken from each car of every working place and channel samples cut every 5 feet in development work are considered sufficient for estimating ore reserves .

Jan 1, 1950

By Wilbert J. Huff

The Explosives Division of the United States Bureau of Mines makes certain laboratory and field investigations that deal with explosions, explosives, ignition, and combustion, together with many related problems, such as the cause and prevention of fires, and certain fundamental scientific questions bearing upon the determination of the chemical and physical mechanisms underlying some of the important phenomena encountered. Its work on explosives has been particularly directed toward determination of the chemical and physical characteristics of explosives designed for use in coal mines and in a more limited way to those used in metal mines, tunnels, quarries, and other engineering operations. This testing is handled by two units of the Explosives Division, as follows:

Jan 1, 1937

By A Trench, S Ulrich, S Hagemann

A synthesis of reported All-In Sustaining Costs (AISC) for 57 Australian gold mines in the period 2014–2016 reveals for the first time the clear influence of grade upon cost performance, most evident in gold-only underground mines. Past syntheses of costs on a single quarter basis have viewed grade as potentially subordinate to other factors such as production scale, mine geometry, mineralisation style and a combination of other mining-related costs. Better understanding of the grade to cost relationship on both an individual mine and whole-of-portfolio portfolio basis is needed. Insight into the roles and trade-offs between other cost-determining factors will aid the future sustainability of the gold mining sector in Australia and assist in targeting prospects for future resource growth and development. While the overall trend is toward higher-grade mines displaying lower costs and vice versa, the grade to cost functions appear radically different between mines based on time-series data. The median cost of Australian gold production in the final quarter of 2016 was A$1081/oz. The threshold for lowest cost quartile production was A$965/oz. Underground production now dominates Australian gold output – with 71 per cent of ounces originating from operations that are either entirely underground or hybrid pit/underground mines. Scale economies are evident, with smaller mines predominately among the higher cost producers. For mines of equal grade, open pit mines are lower cost than underground operations. Reserves are modest, with a median mine life for underground operations of 4.7 years based on current production rates. Reserve additions at mines are likely constrained not only by the finite nature of mineralisation at each mine but also by the geometrical difficulty of accessing suitable drilling locations to test for reserve extensions.CITATION:Ulrich, S, Trench, A and Hagemann, S, 2017. Towards an understanding of production costs at Australia’s underground gold mines, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 241–250 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By M Runkel, J Hammerschmidt, A Holmstr÷m, A Charitos, J Gnntner

"Roasting of sulfide concentrates and refractory ores is an important process step in the production of metals and chemicals. Sulfides of iron, zinc, copper and other metals are oxidised with air between 600 and 1000¦C into metal oxides or sulfates and gaseous sulfur dioxide. Roasting of gold concentrates and ores is a typical processing step used to produce a porous calcine suitable for cyanide leaching. During roasting, elements posing problems are removed through: oxidation (carbon, sulfur, arsenic and antimony) or volatilisation (mercury, chlorides and tellurides, as well as arsenic and antimony).This paper focuses on the environmental aspect of roasting using fluidised bed technology. The approach is to show that roasting is an attractive solution for refractory ore pretreatment since, on the one hand, it is an efficient technique yielding high gold recovery rates, while on the other it is able to comply with current stringent international environmental regulations. In that respect, roasting is compared to alternative processes, ie pressure oxidation (POX), bio-oxidation (BIOX) and ultra-fine grinding (UFG). Roaster high gold recovery rates are ensured through roasting-leaching test work, which allows the derivation of respective roasting parameters that are easily achieved in an industrial plant. On the other hand, key elements of the roasting process are analysed, which make it an efficient and sustainable technology, ie the possibility for energy recovery, the removal of off-gas impurities during the different gas cleaning steps (handling of arsenic, mercury emissions and further hazardous dust and gas emissions) and the treatment of the SO2-gas in order to produce sulfuric acid, and finally the treatment of the effluent produced. The processÆs environmental aspects are summarised by listing roasting emission values that comply with international regulations.CITATION:Charitos, A, Runkel, M, Gnntner, J, Holmstr÷m, A and Hammerschmidt, J, 2013. Roasting - a study on environmental aspects, off-gas, effluents and residue treatment, in Proceedings World Gold 2013, pp 509-518 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Sep 26, 2013

By C. D. Harrison

The 1990 Clean Air Act Amendments specify national objectives for reducing sulfur dioxide emissions from coal-fired power plants System-wide reductions can be achieved by combining various emissions controls technologies and allowances, or compliance credits, that can be purchased, sold, or traded. As part of the Coal Quality Expert development project funded by U.S. Department of Energy and the Electric Power Research Institute (EPRI), Black & Veatch has developed an interactive computer tool, the Acid Rain Advisor (ARA). ARA allows the user to develop integrated system-wide compliance plans by comparing the costs and emissions reduction benefits for emissions allowances and various SO2 control technologies such as coal cleaning, fuel switching, and scrubber retrofits.

Jan 1, 1992

By B. Pesic

The recovery of silver and gallium from flue dust with acidified thiourea solutions was examined. The studied parameters were temperature, particle size, thiourea concentration, acid concentration, and the presence of oxidants. The study of the extraction of these two metals was undertaken in sulfuric acid medium. Silver dissolution was dependent on all examined parameters, whereas gallium extraction was only dependent on the initial acid concentration. The experimental results suggest a new very efficient method for recovery of both silver and gallium. Almost complete recovery of silver and gallium can be accomplished within 90 minutes. A technological scheme for recovery of both metals is also proposed.

Jan 1, 1986

By VIVEK KASHYAP, EVODY TSHIJIK KARUMB, PATRICK TAYLOR

Zinc ferrite is considered an integral phase of zinc residues as their presence can render the processing of zinc and critical metals difficult. It is a refractory phase, formed at high temperature in the presence of zinc (Zn) and iron (Fe) during oxidative roasting in zinc processing plants. Because it is refractory to efficient leaching, unreacted zinc ferrite gets concentrated in the final zinc residue, making it a viable source of zinc and indium extraction. The refractory nature of zinc ferrite makes the extraction of associated critical metals difficult. In order to explore a potential economic and carbon-neutral zinc ferrite processing technique, our previous study proposed the partial reduction of zinc ferrite followed by sulfuric acid leaching of the roasting product [1-3]. As demonstrated in our previous studies, partial reduction of zinc ferrite results in a mixture of ZnO, Fe3O4 and FeO, which can be further subjected to leaching for efficient extraction of zinc and critical metals such as gallium and indium.

Dec 1, 2022

By K. O. Bennington

[Thermodynamic properties of two amphibole asbestos minerals were determined by the Bureau of Mines. The enthalpies of formation were determined by the hydrofluoric acid solution calorimetry method. Heat capacity measurements were made with an adiabatic calorimeter over the range of 5 to 315 K. High-temperature enthalpies above 298 K were determined with a copper-block drop calorimeter to 852 K for amosite and to 902 K for crocidolite. Thermodynamic data determined for amosite of the composition Fe5+2.2766Mg1.5209 Fe+03.12Mn0.0825) Si7.9163Al0.0837022.0363)(OH)1.9637 are: S°298 = 168.8 ± 5 cal/deg-mole; ?Hf°298-15 (elements) = -2419.24 ± 2.95 kcal/mole; ?Hf298.15 (oxides) = -30.92 ± 2.47 kcal/mole; and ?Gf° -2260.1 kcal. Corresponding data found for s3 ,1s = crocidolite of the composition (Na1.9Ca0.1)(Fe2+.024Fe2.5 01Mg0.475) (Si7.9713024Fe+2 2.501Mg0.475) (Si7.971 Al0.029022.095) (OH)1.905 are: S2.98-15 - 161.1 ± 4 cal/deg-mole; ?Rf° 2 298.15 (elements) = -2433.56 ± 2.99 kcal/mole, 10298.1s (oxides) = -87.69 1.70 kcal/mole, and ?Gf°298.15 = -2269.8 kcal.] Calorimetric values from this investigation were combined with other data from the literature to calculate the Gibbs energies of formation and equilibrium constants of formation over the temperature range of the measurements. Tables of enthalpies of formation and Gibbs energies of formation are given as a function of temperature both from the elements and the constituent oxides.

Jan 1, 1978

By Michael W. Roper

McDermitt Mine is located in northern Humboldt County, Nevada, approximately 5 miles south of the Orgeon-Nevada line (Figure I), within the Opalite Mining District. The operation is a joint venture between Placer Amex Inc., San Francisco, and Mineral Exploration Company Ltd., a limited partnership. The McDermitt orebody possesses a discontinuous, gently dipping blanket-like form within 150 feet of the surface over an area 2200 by 2500 feet. The ore zone averages 20 feet in thickness. The ore minerals consist of cinnabar (HgS) and corderoite (Hg3S2C12) deposited parallel to bedding within argillized, Miocene age, tuffaceous lacustrine sediments. Exploration indicates reserves of approximately 3,000,000 tons ore at a grade of 10 pounds mercury per ton (0.5%) An open pit mining operation utilizing scrapers and bulldozers supplies a concentrator with 700 tons of ore per day. The milling operation consists of autogenous grinding, flotation concentration, concentrate dewatering, and tailing disposal, with a design capacity of 100 tons per hour. The furnace section consists of a concentrate filter, vertical six hearth furnace, mercury condensing system, and exhaust gas scrubbing systems. It has a design capacity of 1000 pounds of dry concentrate per hour. The McDermitt orebody is situated about 2000 feet north of the Cordero Mine which, with adjacent properties, produced over 115,000 flasks (76 pounds per flask) mercury from 1941 to 1972' Two other major mercury producers in the district have been the Bretz and Opalite Mines, located in southern Malheur County, Oregon, within 5 miles of the Oregon-Nevada line. Bretz and Opalite have produced about 14,800 and 12,370 flasks mercury respectively, from 1926 to 1965.1/

Jan 1, 1976

By Haitao Wang, Shen Wang, Dongyin Li, Yongjia Xu, Jinzhao Liu, Xiangjun Chen

The safe and efficient driving of the roadway along the gob is becoming an increasing challenge due to the influence of the integrity and bearing capacity of the side coal and rock mass in the gob under the roof cutting mining mode. The mechanism of protection of the gob-side roadway is shown by modeling the influence of roof cutting and the application of a destress technique on the lateral coal and rock mass zones. The distribution characteristics of stress, displacement, and fracture in the overburden are analyzed by numerical simulation method under the three working conditions. The results indicate (1) construction of row holes and blasting can form artificial structure’s fracture plane above the roadway, which can effectively cut the connection of overburden. The artificial structure’s fracture plane decreases the peak value of the lateral coal body’s support pressure. The peak point and the critical location of the stress reduction area are both transferred to the deep part at the same time, providing adequate space for the gob-side roadway. (2) The stress arch foot in the lateral rock mass rises to the rock layer at the bottom of the roof cutting plane after top cutting, as does the high-pressure area of deviator stress. The fracture plane prevents lateral coal and rock mass instability and limits the negative effects of roof deformation above the gob-side roadway. (3) The triangle slip fracture zone moves to over the roof cutting drilling zone due to the fracture plane, showing a high position compared to its original zone.

Jul 11, 2023

By G. W. Reimers

Exothermic oxidation reactions of sulfide minerals can produce self-heating situations in underground mines. If the self-heating is unchecked and the resulting heat is not sufficiently dissipated, the result can be spontaneous combustion fires. This U.S. Bureau of Mines report describes laboratory tests conducted to measure the reactivity of pyrite following treatment with additives to counter its self-heating tendencies. Thermal gravimetric analysis (TGA) and isothermal oxidation procedures were used to obtain information on pyrite ignition and the behavior of pyrite during low-temperature oxidation to ferrous sulfate. The determination of treatment effectiveness was evaluated by adding 5 wt pct of the compound to the pyrite before milling. The resulting samples were then heated in moistened air and after 24 h, the amount of pyrite oxidation determined. Results of these tests indicated that several phosphate compounds reduced the oxidation of pyrite to sulfate to about 50 pct or less compared with the oxidation of untreated sample at 200° C. Cupric sulfate reduced oxidation to about 20 pet untreated sample at 200° C. Sodium, calcium, and potassium chloride reduced pyrite oxidation to levels of approximately 15 pct or less of the untreated sample. Using only 1 pct of sodium or calcium chloride was effective in slowing pyrite oxidation to less than 20 pct of the untreated sample.

Jan 1, 2010

By D. R. Tesarik, T. C. Trancynger, W. G. Pariseau

The Davis detector cavern was excavated on the 4850 Level of the former Homestake Mine in the mid- 1960s to house a novel neutrino detector developed by Nobel Prize winner Dr. R. Davis. The original cavern was about 9.1 x 16.8 x 9.7 m high, excavated in a Precambrian amphibolite known as the Yates unit and enlarged in 2010. In 2009, the National Science Foundation selected the site for development as a national Deep Underground Science and Engineering Laboratory (DUSEL). This retrospective rock mechanics study of the Davis cavern is intended to inform design of the much larger caverns and laboratory rooms being considered for the same 4850 Level. Provision for rock mass variability, sequential excavation, jointing and wall damage from blasting are included in analyses that were done using two different computer codes and methods (FLAC3D and UTAH3). The results are in agreement with observations and show that cavern stability is indeed the case. The main conclusions are: (1) if conventional scaling of laboratory rock elastic moduli and strengths is done, then 0.25 and 0.50 scale factors are reasonable, (2) measured statistical and spatial variability in rock properties should be incorporated into geomechanical analyses to better reflect reality, (3) discontinuities (“joints”) are essential to reliable analysis of excavation safety and when properly taken into account eliminate the need for empirical “scaling” of laboratory rock properties, (4) major geologic features such as rhyolite dikes and breccia zones may be important to excavation safety at Homestake and should be explicitly represented in analyses and (5) quantifying wall blast damage is important to engineering design of underground excavations.

Jan 1, 2012

By Thomas A. Davidson

The Bureau of Mines Helium Field Operations has developed a method for determining the helium-3 CHe) content of natural gas and mixtures eontaining 0.5% to 100% helium. The helium is preconcentrated by passing samples through anhydrone, a liquid-nitrogen-cooled chareoal trap, and a titanium absorption trap. The mass spectrometer used to analyze the samples is tuned to 3He at mle ::=; 3.016. Hydrogen deuteride (HD+) interference is minimized by the titanium absorption trap and high instrument resolution. Sample measurements are compared with identically made measurements of a gravimetrically prepared primary standard. The method has a limit of detection of 0.15 ppb with a relative deviation of 13% below the 1~ppb level. The relative deviation is less than 7% for gases eontaining over 100 ppb 3He. The helium isotope ratios of natural gases from the Bureau's Tuck-Trigg Dome Red Cave, the Bush Dome Red Cave, and the Bush Dome Brown Dolomite Formations are distinguishable from one another and from the isotope ratio of crude helium produced from the Panhandle-Hugoton Gasfield.

Jan 1, 1990

By Zizheng Zhou

The separation of cobalt from other base metals can be achieved using diffusion dialysis in a two-compartment reactor separated by an anion exchange membrane. In this work, a Neosepta AFX membrane was used to study the separation of various metals in an acid-chloride solution. The separation of one metal from another was investigated on the basis of the various chloro-complex stabilities. The metals investigated included cobalt, nickel, copper(II), iron(III), aluminum, zinc and cadmium. The impact of several critical experimental conditions, including the initial metal and chloride concentration in the feed solution and their initial concentration in the receiving solution, on the separation factor and the metal transfer rate is reported in this work. Effective separation of different metal pairs was achieved under the experimental conditions.