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By Steve Kral

Mining is a risky business. The cost of exploring for new mineral deposits is high and the success rate of finding one worthy of development is relatively low. And once a deposit is identified, there is no guarantee that it will become a mine. Mining companies have to spend a lot of time and money on due diligence, prefeasibility and feasibility studies, and sustainable development and environmental issues before deciding whether the deposit will become a profitable mine. And then there are the millions of dollars spent in developing the deposit and building a mine. No wonder the mining industry has trouble attracting investment capital. Companies take a large risk in developing reserves and investors take a similar risk in investing in them. So mining companies must be as accurate as possible in determining mineralization that makes it to reserve status. After all, reserves are a company’s most important asset and one of the critical factors analysts and funding sources look at when determining if a mine or mining company is worth the risk.

Jan 1, 2003

By K. L. Paciorek

The primary objective of this program was to determine the toxic fume and fire hazard of elastomeric noise abatement materials with a limited effort expended on the evaluation of specific treatments applied to mine timbers. Under the conditions employed, for a dimethyl silicone rubber, formaldehyde presented the major toxic threat; acrolein was the main toxic species for both polyester and polyether urethanes and an acrylonitrile-butadiene-styrene resin; in the case of flame-retardant styrene-butadiene rubber compositions, hydrogen chloride, and in the case of PFA-treated timber, furfural, were the major toxic compounds. Of the materials studied, the flame-retardant styrene-butadiene rubbers presented the greatest toxic hazard. The generated data have been reduced to toxic load tabulations amenable to employment for potential hazard assessment.

Jan 1, 1982

By Yan Fu, Chuanlin Fan, Xiujing Zhai, Yongfeng Chang, Binchuan Li

"The mechanism of the selective chlorination of nickel and cobalt in reduced limonitic laterite was investigated. Reduced ores were obtained from the reduction roasting of a limonitic nickel laterite. The thermal decomposition of related chlorides and the selective chlorination of the reduced ore were carefully studied using Differential Thermal analysis - Thermogravimetry (DTA-TG), X-Ray Diffraction and chemical leaching methods. In the temperature range of 300 - 500 °C, aluminum chloride hexahydrate was chosen as chloride agent. During the selective chlorination, reduced ore was partly chlorinated by hydrogen chloride generated from the thermal decomposition of chloride agent. Then the oxydrolysis of ferrous chloride re-generated hydrogen chloride was used to further the chlorination. Finally, iron was mostly rejected as hematite, while nickel and cobalt were selectively chlorinated. More than 91 %Ni , 90%Co and about 4%Fe in the chlorinated ore was extracted using a water leaching method.IntroductionLimonitic nickel laterite is an important nickel reserve, which is rich in iron in the form of goethite and has low contents of magnesium, aluminum, silicon, and other elements. Goethite is the main nickel-bearing mineral in the laterite [1]. Nickel is generally present in goethite in three modes: (a) associated with amorphous or poorly crystalline goethite, (b) weakly adsorbed to the crystalline goethite surface, and ( c) as lattice substituents [2]. In addition, cobalt in limonite was found to be associated with a manganese phase rich also in nickel [3]. These results indicate that the extractions of nickel and cobalt are highly dependent upon the degree of the decomposition of the nickel, cobalt-bearing minerals.For the selective extraction of nickel and cobalt from limonitic laterite, the ore is usually processed by hydrometallurgical methods, including high pressure acid leaching (HPAL) [4] or reductive roasting - ammonia leaching (Caron process). However, there are still some disadvantages in these processes, such as the expense of autoclaves and scaling problems in HP AL [ 4, 5], and relatively low overall metal recovery, large energy and reagent requirements for the Caron process."

Jan 1, 2011

By G. Onal, E. Gock, H. Dincer

"During the conventional chromite dissolution process; chromite ore concentrates were ground to at least <90 µm using the ball mill and after adding Na2CO3 and leaning material (instead of CaCO3 also residues from the leaching process) at a ratio of e.g. 1:0.75:1.2, depending on the specific formulation, follows the digestion e.g. in a rotary kiln, heated in countercurrent. The reaction time is 3 to 8 h. The calcine is comminuted and leached with water or a thin chromate solution. Depending on the type of ore used, the yield is 75-85 %. The disadvantage of the above-described process is that it is impossible, using current technology, to achieve a complete digestion of the chromite ores and that the chromium-containing residues have to be disposed of as special waste, involving a large financial effort.The subject of this study was the investigation of the effect of vibratory milling on chromite dissolution. The impact stress in the vibratory mill causes the mechanical activation of chromite, which is characterised by lattice defects. The degree of mechanical activation can be quantified by measuring the relative X-ray diffraction intensities at the parent lattice planes of chromite. Each relative X-ray diffraction ratio correlates with a certain degree of digestion. As opposed to vibratory milling, a complete digestion of chromite using a ball mill is impossible, even at extremely long grinding times.In this investigation a chromite concentrate from Pulpinar/Kayseri with content of 44 % Cr2O3 and the Eccentric Vibratory Mill was used. In order to verify the effect of the vibrating grinding on chromite dissolution, reaction kinetic investigations on digestion and leaching were carried out."

Jan 1, 2003

With a growing concern over mine fires, prevention and suppression are hot topics. NIOSH researchers Michael Trevits, Alex Smith and Edward Thimons discuss their research in the area. The leading causes of US mine fires include flame cutting and, welding operations, frictional heating and ignitions, electrical shorts, mobile equipment malfunctions and spontaneous combustion. From 1990 through 2007, there were 1601 reportable fires that occurred in the US mining industry -an average of 89 fires per year. The fact that mine fires are occurring with alarming regularity reinforces the importance of recognizing and eliminating the potential hazards. These statistics also show the need for improved control and suppression technology to ensure worker health and safety if a mine fire does occur.

Jan 1, 2008

By F. Yang, T. Pan, T. Zhao, Z. Shao, D. Zuo

In this paper, the development and applications of robot technique is summarized, and the status and developing direction of China coal mine robotics. A description of the research meaning, frame and key technology of underground salvage robot is given. Keywords: coal mine, robot, special robot, salvage robot.

Jan 1, 2003

By Stephen H. Stow

Economically important heavy minerals have been studied in the sediments of offshore Alabama and Mississippi. Sandy samples have been collected from a 200 square mile area from Dauphin to Horn Islands. The minerals concentrate in sediments under 10-35 feet of water and may reach 2.4% by weight. The distribution of total heavies is related to the hydrodynamics of offshore Alabama-Mississippi; the greatest concentrations are in the Gulf where the longshore current is diverted from its westerly flow. Preliminary observations indicate that the distribution of individual minerals can be correlated with wave energy, particle size, specific gravity, and morphology of the mineral grains. The coarser, lighter minerals--kyanite, sillimanite and staurolite--are concentrated in relatively low energy areas in the Mississippi Sound and at the end of Dauphin Island in the Gulf; maximum concentration for each is 34.5%, 6.0% and 43.2%, respectively, of the heavy minerals. The smaller, more denseminerals--ilmenite, rutile, leucoxene, zircon and monazite--are concentrated in high energy areas in the Gulf; maximum concentration for each, respectively, is 42.0%, 5.9%, 9.652, 10.7% and 1.3% of the heavy minerals. On the basis of a minimum concentration of 1% heavy minerals, two areas have been identified and the total amounts of different minerals have been calculated to a depth of 25 feet of sediment. The data are preliminary and in no way indicate recoverable reserves, but they do show that on the order of 20 million tons of heavies exist. These deposits should be termed "indicated resources".

Jan 1, 1975

By H. R. Hanley

The characteristics of alloyed anodes and their influence on the products of electrolysis and power consumption have been noted previosly ously in the literature.&apos; This paper presents data in continuation of the subject given in an earlier paper by the authors.2 As noted in the authors1 previous paper, the objectives of this work are the reduction in the cell voltage and the purity of the cathode product resulting from the use of alloyed lead anodes. The fundamental facts developed in the previous paper were: (I) that the presence of a small amount of calcium lowers the anode potential approximately 50 per cent. below that of pure lead; and (2) that thallium when present to the extent of 4 per cent. rendered the anode practically stable, thereby lowering the contamination of the cathode by lead to a very insignificant amount. The specific features determined in this recent work are as follows: 1. The lowest limit of calcium in the lead which will still give maximum lowering of the anode potential. 2. The characteristics of the thallium-calcium-lead anode. 3. The characteristics of the silver-calcium-lead anode. 4. The characteristics of the thallium-silver-calciurn-lead anode. It was determined that 0.1 per cent. calcium in the anode was sufficient to lower the anode potential approximately 50 per cent. Table 1 shows the characteristics of alloyed anodes tested. The anode composed of 0.1 per cent. calcium, 4 per cent. thallium, and 95.9 per cent. * Professor of Metsllurgy, Missouri School of Mines. t Professor of Metallurgy and Ore Dressing, Hea.d of Department, Missouri School of Mines. Assistant Professor of Metallurgy, Missouri School of Mines. 1 Among others, see C. G. Fink and Li Chi Pan: Insoluble .Anodes for the Electrol-ysis of Brine [Trans. Amer. Electrochem. Soc., (1926) 49, 83-1341 and other papers by these authors. U. CJ. Taintori, 8. G. Taylor arid H. P. Ehrlinger: Lead Alloys for Anodes in Electrolytic Production of Zinc of High Purity. Truns. A. I. M. E. (1929) 192-200. 2 H. R. Hanley, C. Y. Clayton, D. F. Walsh: Investigation of Anodes for Production of Electrolytic Zinc. Trans. A. I. M. E. (1930) 275

Jan 1, 1931

By E. Pothas, P. R. Jochens, L. B. McRae, D. D. Howat

Discussion H. J. S. Kriek*: During 1956 Iscor became interested in the acid smelting in a blast furnace of siliceous ores containing 15 per cent silica. This interest was as a result of the successful acid smelting in Germany at a CaO/SiO2-ratio of between 0.4 and 0.8. Acid smelting can become an economic proposition due to a reduction in slag volume accompanied by a lowering of the coke rate. Paschke and Hannel1 found that a free flowing slag (i.e. viscosities of less than two poise) would not be obtained at C/S-ratios lower than 0.82 and that the addition of 3 to 4 per cent alkalies became necessary. In titanium bearing slags it was established that maximum fluidity is obtained with 10 per cent TiO2. Acid smelting resulted in increased sulphur pickup with decreasing O/S-ratio and desulphurization with soda-ash outside the blast furnace was necessary. At Iscor Mr B. B. Segal established that, for slags with (CaO+MgO)/SiO2-ratios of 0.5, 0.6 and 0.7, the addition of alkalies and TiO2 resulted in a substantial decrease in melting point and also of viscosity. At a (CaO+ MgO)/ ratio of 0.53 the addition of 10 per cent TiO2 is as effective as the addition of 3.75 per cent alkalies. This means that a blast furnace can be run without the addition of alkalies and consequently the detrimental effect of alkalies on the lining can be avoided; also that fluid slags with 5 to 15 per cent MgO and 8 to 10 per cent Al203 can be obtained with 10 per cent TiO present. (CaO+ MgO)/SiO2-ratios smaller than 0.5 the addition of titania resulted in a decrease in melting point and also in an increase in viscosity. In ceramic crucibles the slag did not become more viscous with time but thickened appreciably when held in graphite crucibles. No titanium carbonitride could be detected and only Ti20a was found. Prof Howat also did not obtain the TiC-TiN solid solution and therefore it appears that the thickening of the slag is not solely due to the formation of the titanium carbonitrides. This is in agreement with the work of Michailov and Belyakova2 who found that TiO2 decreased the viscosity of acid slags and increased the viscosity of basic slags; that titanium sesquioxide (Ti203) and titanium monoxide (TiO) increased the viscosity of both acid and basic slags and that titanium carbide, which is found in larger amounts in basic slags than in acid slags, considerably increases the viscosity of both acid and basic slags. Patzak et al3 describes the following oxygen deficient titanias, viz Ti2O, TiO, Ti203, Ti3O5, Ti4O7, Ti5O9, Ti6O11- Ti10O9,Ti7O13,Ti8O15 and Ti9O17. Until about 1959 a fair amount of data was available on the viscosity of normal slags to which titania was added2, 4, 5, 6, 7, 8, 9. At Iscor an electromagnetic viscometer based on the viscometer developed by Bockris10 was built. Progress was very slow as our main attention was centred around problems associated with the operation of the Rotor process. During 1962 and 1963 two papers containing fundamental work on the viscosities of slags in the systems CaO-SiO2-TiO211 and CaO-SiO2-TiO2 at 0 per cent A1203, 10 per cent Al203 and 20 per cent Al20312 appeared. The work of Ross and Ohno obviated the necessity of any further work by Iscor on the viscosity of blast furnace slags containing titania. Ross gave the following optimum compositions for blast furnace slags containing titania. Titania % Lime % Silica % Lime/Silica-ratio 20 30-42 38-50 0.6-1.10 30 27-36 34-43 0.63-1.06 40 24-32 28-36 0.66-1.14 The optimum temperature is 1,500°C because at this temperature the largest workable composition range exists. Higher temperatures result in narrower optimum composition ranges and in increased rates of thickening. The factors affecting the rate of thickening were also investigated by Ross and Ohno, as the main problem with slags containing titania is not finding compositions which are fluid at operating temperatures but the thickening due to a lack of oxygen. With these slags when there is insufficient oxygen to co-ordinate each silicon atom with its own individual shell of four oxygen ions, some sharing of oxygen ions must occur, resulting in the formation of large silicate ions and a rise in viscosity. The formation of oxygen deficient titania can be prevented as long as iron oxide or manganese oxide remain in the liquid slag. Iwase13 and also Viens, Camp bell and Rogers14 obtained in the experimental smelting of ilmenite ores fluid slags with more than 3 to 5 per cent manganese oxide and ferrous oxide in the slags. It appears that titaniferous ores can be smelted in blast furnaces under the following conditions: (1) Acid slag operation, i.e. at C/S-ratios from O.6-1.0. (2) Dolomite can be used as a flux as MgO is a network modifier. (3) The Al2O3-content must be less than 15 per cent as alumina promotes thickening of the slag and decreases the optimum composition range. (4) The hearth temperature must not exceed 1,500°C. According to Ross a definite amount of energy is required for each ton of iron produced and in order to increase production an increase in heat

Jan 6, 1969

By ?. ?. Zimin, B. A. Kutlin, X. Baddamsuren

"The ore features of Mongolian fluorspar deposits are shown in the paper. The possibilities of the technological of ore processing on the Bor-Undur combine scheme improvement are analyzed there. Some recommendations are done for scheme development: spare output technologies, the usage large-sized ore separation, flotation by columnar flotation cells, leaching and concentrate agglomeration.New evaluation and usage of mineral sources approach was developed for Mongolian fluoride ores. The conception allows using maximally available mineral resources for getting of metallurgical and acid fluoride concentrates while mining and technological processes: mining (as well as quarrying) -storing, blending-crushing, grinding-combined processing methods-metallurgical using preparation It has been established mining and processing technological schemes should include:• spare blasting technologies providing minimum effect on fluorite crystals which follows ore selective excavation;• large-sized ore separation methods resulted in maximum recovery for these operations in respect to metallurgical fluorite concentrates;• grinding and flotation regimes providing high-quality flotation concentrates getting for their usein chemical industry;• fluorite concentrates agglomeration technologies for providing ability of their use in metallurgical and cement industriesThe results of industrial usage of jugging machines, columnar and pneumomechanical flotation machines for fluospar dressing.The technological approaches implementation in “Bor-Undur” plant maintenance within 1993-2002 years leaded to:• change plant status from just flotation one to gravitational-flotation characterized by technological process stages which are traditional for metallurgical industry and plant capacity increase from 400 to 600 thousand tones of treatment ore as well as concentrate production increase from 120 to 150-170 thousand tones including 120-140 thousand tones of acid grade concentrates with 97% CaF2 ."

Jan 1, 2003

By Robert A. Kinzie

PAGE Introduction.............334 Climate.............335 History............335 Geology.............341 Mining..............343 Shafts............343 Stations and Ore-Bins....345 Levels. Drifts. Cross-Cuts, Winzes. etc........347 Stoping Surface-Pits,......... 351 Underground Stoping........... 332 Chutes. Drilling. etc.....354 Tramming............. 358 Sampling. Maps and Assaying.........360 Labor.............361 Machinery.............362 Milling,..............362 Crushers.............366 Battery-Foundations...........368 Power (Water—Coal)...........373 Methods of Catching the Free Gold...374 Screens..............376 Apron-Plates,.......376 Concentration..........378 Clean-Up,.............378 Labor..............382 Costs..............382 Introduction. The Treadwell Group comprises five mines. four of which are situated on the northeast side of Douglas island. directly opposite the town of Juneau, and one about four miles back of Juneau. These mines are owned by four separate companies. but are all under the same general management. Those situated on Douglas island are : The Treadwell mine. owned by the Alaska Treadwell Gold Mining Co.. the Mexican mine. owned by the Alaska Mexican Gold Mining Co.; and the Ready Bullion and Seven-Hundred-Foot mines. owned by the Alaska United Gold Mining Co. The Juneau mine. owned by the

Jan 1, 1904

By J. M. Houston

"In October 1982, Rio Algom Limited, purchased a tin prospect near East Kemptville, Nova Scotia, from Shell Canada Resources. This purchase led to the development of the first substantial tin mine in North America. This paper deals with the drillingand blasting operations at Rio Algom, in particular the use of emulsion explosives technology.This presentation describes what emulsion explosives and Heavy AN/FO are, and the support system used to deliver these products to the blast site.HistoryThe East Kemptville mine site was discovered by Shell Canada Resources (SCR) in July 1978. SCR started diamond drilling on the tin prospect in March 1979. After 8220 m of BQ drilling, it was estimated a reserve of 56 000 000 tonnes of 0.165% Sn was present.The property, which is located 54 km northeast of Yarmouth, Nova Scotia (Fig. 1), showed, after a feasibility study that, given an ore treatment rate of 9000 tonnes per day, a mine life of 17 years could be expected.GeologyThe tin deposit of the East Kemptville Mine, occurs in granitic rocks at their contact with metasedimentary rocks. The mineralization is in the form of cassiterite with associated sulphides of copper and zinc. The tin occurs in a stockwork of tin bearing Greisen alteration zones.East Kemptville MineThe mine planned an open pit operation with the first five years being mined at a rate of 14 000 tonnes per day, of which 5000 tonnes will be waste and low grade ore. Eventually, the total mining rate will be 21 000 tonnes per day, including 12 000 tonnes of waste and low grade ore. Figure 2 shows the general layout of the mine site.East Kemptville Chooses Emulsion ExplosivesIn December 1984, East Kemptville Mine awarded the Explosives Division of C-I-L Inc. a contract to supply explosives to the borehole. The system chosen gives East Kemptville the on-the-bench choice of regular AN/Fa, or a high density, Heavy AN /Fa product called SuperAN· or a water resistant emulsion explosive called MAGNAFRAC· for holes that cannot be dewatered.Emulsion ExplosivesAn emulsion is an intimate mixture of two liquids that do not dissolve in each other. Emulsions have contributed to our daily lives in such products as insecticides, salad dressings and cosmetics."

Jan 1, 1987

By J. Fletcher Smoak

Consumption of abrasive materials in the United States in 1981 was approximately $340 million, of which 37 % was industrial diamond (natural and synthetic), 39% manufactured abrasives, and 24% natural abrasives. Production and shipments of natural abrasives, excluding emery and industrial diamond, decreased in quantity 9% and 7%, respectively, when compared with that of 1980. Emery showed the largest change in output, decreasing 30% in quantity and 20% in value. Production of nonmetallic .manufactured abrasives material plus shipments of metallic abrasives material decreased 5% in quantity but increased 4% in value. Nonmetallic manufactured abrasives consisted of aluminum oxide (fused) and crude silicon carbide produced in the United States and Canada and accounted for 63% of the value of all manufactured abrasives. Metallic abrasives shipments included chilled and annealed iron shot and grit, steel shot and grit, plus cut wire, aluminum, and stainless steel shot and equaled 37 % of the value of all manufactured abrasives.

Jan 1, 1982

By Yajie Wang

High horizontal stress and its adverse effects on longwall gate entries have been witnessed in many US coal mines for the past years. It can cause cutter roofs and leads to roof falls either in the headgate or tailgate entries depending on the relative orientation of the panel to the maximum principal horizontal stress. Underground observations have shown that if the maximum principal horizontal stress is perpendicular to the panel entries, the roof is in the worst condition and if the maximum principal horizontal stress is parallel to the entries, the entry is in the best condition. However, little has been done about the high horizontal stress effects on panel entries subject to different orientations of high horizontal stress. In order to design a longwall panel and control the roof effectively, the redistribution and concentration of high horizontal stress field must be fully understood. Using 3-D finite element analysis, this paper presents the high horizontal stress redistribution and concentration for a retreat longwall face with different orientations of high horizontal stress. Based on the results, possible roof failures for different orientations of high horizontal stress are analyzed. A panel design criterion is recommended.

Jan 1, 1996

By R. T. Stain

Dynamic methods of load testing piles using a drop hammer to generate the input force have been in use for many years. These systems are based on the wave equations developed at the Case Institute of Technology, Cleveland, Ohio in the late 1960&apos;s and are well documented. Such tests are generally restricted to preformed driven piles, the driving hammer being a convenient means of generating large input forces. In 1979 the CEBTP (France) was sponsored by the FNTP (Fédération Nationale des Travaux Publics) to undertake research into existing classical methods and to develop a test that would be equally applicable to bored cast in place piles. The result of this research programme is a methodology known as SIMBAT, and which was included in a series of class A prediction trials in Belgium in 1987 (Refs 1 & 2, Paquet 1987, 1988). The purpose of this paper is: a) to examine the type of &apos;dynamic&apos;/&apos;static? load test correlations that can be obtained. b) to highlight the differences between the classical&apos; methods and the &apos;SIMBAT&apos; technique.

Jan 1, 1989

By D. N. Ferguson

The separation of certain minerals by electrostatic techniques can be difficult due to their similar electrical conductivity, and any technique to improve this can be useful in certain difficult separation applications. Triboelectric differences between mineral scan result in certain minerals acquiring small positive or negative charges prior to inductive electrostatic separation. By changing the polarity of inductive separation, relative changes in mineral recovery can be observed, which may be beneficial. In this brief study a JKTech mineral liberation analyser (MLA) was used to evaluate the effect of polarity change on selected minerals present in the Richards Bay Minerals deposit, and an approximate triboelectric series based on this behaviour is presented.

Jan 1, 2010

By P J. Terbrugge, O K. H Steffen

Open pit slope design has conventionally been effected as a bottom up function utilising available geotechnical information. This results in a decision criterion based on Probability of Failure and Factor of Safety with a risk assessment being carried out on the proposed design slopes. The design approach recommended reverses the process using fault event tree decision methodology. The consequences of this approach are that acceptable risk criteria have to be determined by mine owners. Slopes are then designed by the technical staff to achieve these corporate goals. Benefits that arise from the process are that the owners take a proactive decision on the risk benefit relationship allowing the technical staff to optimise the geotechnical exploration programme and design.

Jan 1, 2006

By J. H. Swisher

When an Fe-Mn alloy is internally oxidized, the inclusions formed are MnO which contains some dissolzled FeO. In the internal oxidation reaction, not all of the manganese is oxidized; some remains in solid solution as a result of the high Mn-0 solubility product in iron. Taking these factors into consideration, the rate of internal oxidation of an Fe-1.0 pct Mn alloy is computed as a function of temperature, using available thermodynanzic data and recently published data for the solubility and diffusivity of oxygen in iron. The predicted and experimentally determined rates for the temperature range from 950 to 1350°C are in good agreement. ThE rates of internal oxidation of austenitic Fe-A1 and Fe-Si alloys have been studied extensively.1"4 Schenck et al. report the results of a few experiments with Fe-Mn alloys at 854" and 956C, and Bradford5 has studied the rate of internal oxidation of commercial alloys containing manganese in the temperature range from 677" to 899°C. When Fe-Mn alloys are internally oxidized, the inclusions formed are solutions of FeO in MnO, the composition depending on the experimental conditions. Since the thermodynamics of the Fe-Mn and FeO-MnO systems have been investigated,6"9 and since the solubility and diffusion coefficient of oxygen in y iron have been determined recently,&apos; it is possible to predict the rate of internal oxidation from known data. The calculations used in predicting the rate of internal oxidation will first be outlined, then the results of the prediction will be compared with the experimental results of this investigation. PREDICTION OF PERMEABILITY FROM THERMODYNAMIC AND DIFFUSIVITY DATA Oxygen is provided for internal oxidation in these experiments by the dissociation of water vapor on the surface of the alloy. The dissociation reaction is: + H2(g) + [O] [1] where [0] denotes oxygen in solution. The equilibrium constant for this reaction is known as a function of temperature:&apos; log As oxygen diffuses into the alloy, oxide inclusions are formed which are MnO with some FeO in solid solution. The reactions occurring are: [Mn] + [0] = (MnO) [31 and [Fe] + [0] = (FeO) [41 where [ Mn] is manganese dissolved in iron and (FeO) is iron oxide dissolved in MnO. The overall reactions may be written as follows: [Mn] + HOte) = (MnO) + H2(£) [5] and [Fe] + H20(g) = (FeO) + Hz(R) [61 The standard free-energy changes and equilibrium constants for Reactions [5] and [6] are known.6 Therefore the equilibrium constants for Reactions [3] and [4] may be obtained by combining known thermodynamic data for Reactions [I], [5], and [6]. For Reactions [3] and [4]: K = and For the present purpose, both the Fe-Mn7,8 and FeO-~n0&apos; systems can be considered to be ideal, i.e., [amn] = [NM~] and (aFeO) = (NM~~) = 1 - (NFeO) where the Ns are mole fractions. These relations, together with Eqs. [I] and [8], permit us to compute both the oxide and metal compositions as a function of temperature and oxygen potential at any point in the specimen. For cases where the oxygen concentration gradient between the surface and the subscale-base metal interface is linear, the kinetics of internal oxidation is an application of Fick&apos;s first law: where dn/dt is the instantaneous flux of oxygen into the specimen, g-atom per sq cm sec; 6 is the instantaneous thickness of the subscale, cm; Do is the diffusion coefficient of oxygen in iron, sq cm per sec; p is density of iron, g per cu cm; h[%O] is the oxygen concentration difference between the surface and sub-scale-base metal interface, wt pct. B6hm and ~ahlweit" derived an exact solution to the diffusion equation for systems in which there is a stoichiometric oxide formed. They showed that the oxygen concentration gradient is given by a rather complex error function relation. For the Fe-Mn-0 system and for most other systems that have been studied, however, variations in oxide compositions are small and rates of internal oxidation are sufficiently slow that the deviation from linearity in the concentration gradient of oxygen is negligible. The mass of oxygen transported across a unit area of the specimen for the total time of the experiment is given by the mass balance equation:

Jan 1, 1969

By S. Masambi, S. M. Bradshaw, C. Dorfling

The presence of iron in base metal leach solutions complicates nickel and copper recovery processes. Leach solutions are commonly purified by precipitating iron as hematite, jarosite or goethite. Fast oxidation of ferrous to ferric is an important step to remove iron by precipitation at low pH values. This study reviews the ferrous oxidation mechanism and reaction rates in chloride media using molecular oxygen. This work contributes to the development of a better understanding of iron precipitation processing steps, which continues to be an important matter in hydrometallurgy. The specific objectives were to investigate (1) the effect of key process variables (acid concentration, copper concentration, and temperature) on the rate and extent of ferrous oxidation by dissolved oxygen and (2) the oxidation mechanism by varying chloride and hydrogen ion concentrations separately. Batch oxidation tests were performed using synthetic chloride solutions containing 45 g/L Fe and 3 g/L Ni at 50 and 80°C, dissolved Cu concentrations (0.3 to 3 g/L) and HCl concentrations (2 to 4.6 M). Regression analysis indicated that the significance of the effect of different variables on the extent of oxidation achieved after five hours decreased in the following order: HCl concentration > Cu concentration > temperature. Increasing the HCl concentration from 4 M to 4.6 M caused the degree of oxidation achieved after two hours to increase from 58.3% to 98.1%. The catalytic effect of dissolved copper on ferrous oxidation was more significant at low acid concentrations. Experimental data suggest that rapid oxidation of ferrous is possible at high HCl concentrations for two main reasons. First, the formation of ferrous chloride complexes allows ferrous oxidation to proceed without the formation of superoxide as an intermediate product (the rate limiting step in dilute acid solutions). Second, the formation of ferric chloride complexes stabilises the oxidation product.

Jan 1, 2016

By A. D. Pelton

The evolution of thermodynamic models of molten salts and slags, starting with Temkin in 1948, has culminated in today?s sublattice models. These modern models have permitted the development of large thermodynamic databases of model parameters, which can be used to calculate and predict accurately the properties of multicomponent solutions. The model parameters are mainly obtained by the critical evaluation and optimization of available thermodynamic and phase equilibrium data for binary and ternary subsystems. The models are then used to estimate the properties of multi component slags and salts from these assessed binary and ternary parameters. When used in conjunction with thermodynamic databases for solids and solid solutions (salts, oxides, sulphides, metals, etc.), other liquid solutions (such as liquid metals),and gases, these databases permit the calculation of complex multiphase equilibria in multicomponent systems, with applications in a wide variety of industries. These calculations can be performed with general commercial software for Gibbs energy minimization, which can also be incorporated into process simulation software packages. Keywords: slags, salts, solution models, thermodynamics, phase diagrams, oxide systems, molten salts

Jan 1, 2004