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By W. Kunda, J P. Warner

SHERRITT GORDON Mines Limited annually produce 300,-000 pounds of pure cobalt by processing a nickel-cobalt-copper-iron sulphide ore which is mined at Lynn Lake, Manitoba. The ore (containing 0.04 per cent Co) is •concentrated to give a concentrate containing l 1-12 per cent nickel, about 2.0 per cent .copper, and 0.25-0.35 per cent cobalt, which is treated at the Refinery in Fort Saskatchewan, Alberta. The concentrate is leached in an ammoniacal solution of ammonium sulphate using compressed air as the oxidant. After filtration the leach solution contains, besides nickel and cobalt, copper ammines, sulpharnates and salts of various thionic acids. The copper is removed by precipitation as sulphide and the sulphur compounds converted to ammonium sulphate by a combined oxidation and hydrolysis treatment. The purified solution •Containing about 45 gpl of nickel and 1 gpl of cobalt is treated with gaseous hydrogen under pressure at 350°F to precipitate metallic nickel powder, the precipitation 'being terminated when the nickel content of the solution is about 0.8-1 gpl. Only about

Jan 1, 1962

By C A. Fleming, C J. Ferron, P T. OÆ

A new process is described which allows for direct recovery of base and precious metals (Au and PGMÆs) from sulphide concentrates or other feedstocks. The process was developed initially for application to ores or concentrates that are too low-grade for economic treatment by the conventional smelting and matte leaching process (for PGMÆs). It is now believed that the process may be widely applicable to a variety of materials including PGM concentrates containing high amounts of chromium or magnesium. Operating conditions have been identified under which gold, the PGMÆs and the base metals dissolve simultaneously in a pressure leaching autoclave. The product of this leaching process can then be treated by one of several flowsheets, of which the preferred will likely involve solid/liquid separation, followed by precipitation (or adsorption on activated carbon) of gold and the PGMÆs, followed by copper solvent extraction/electrowinning. Depending on the concentrations of nickel and cobalt, the copper raffinate would either be partially recycled to the autoclave to allow Ni/Co concentrations to build up, or it would proceed directly to a Ni/Co refinery. This paper deals with the pressure leaching process and the recovery of gold and the PGMÆs from the pressure leach product.

Jan 1, 2000

By T. Moyo

Ammoniacal solutions are attractive and effective lixiviants for the dissolution of chalcopyrite, allowing for the formation of stable ammine complexes with copper while rejecting iron. Chalcopyrite dissolution in ammoniacal solutions is reviewed from literature, and results from a recently completed study evaluating it through electrochemical and controlled leach studies are discussed. Ammoniacal leaching of copper ores dates back to 1916 in the leaching of low grade copper ores in the Kennecott leaching plant and in the Calumet and Hecla mines, followed by the commercialisation of the Sherrit Gordon process for the ammoniacal leaching of sulphide copper ores in 1948. Since then ammoniacal leaching has received intermittent focus from researchers with the development of the Anaconda’s Arbiter process and the unsuccessful Escondida process. In these leaching processes, the formation of surface deposits has historically been reported to result in the passivation of the chalcopyrite surfaces, whereas oxygen mass transfer in solutions has been the focus of process design improvements. Results of studies on high grade chalcopyrite material is presented in this paper which show that although a surface deposit was formed, it did not result in the passivation of chalcopyrite in ammoniacal solutions, and that while oxygen transfer is critical, it is not the direct oxidant in the reaction. INTRODUCTION Ammonia and its salts have been used in the hydrometallurgical treatment of non-ferrous metals such as copper, cobalt and nickel from oxide ores. Its chemistry allows for the selective dissolution of desired metals, leaving behind gangue including iron. Chalcopyrite is a copper sulphide ore known to be very stable and hence recalcitrant to many hydrometallurgical processing methods. Thus, the dissolution of chalcopyrite continues to be an attractive area of research. The dissolution of chalcopyrite in ammoniacal solutions is possible due to the stabilisation of copper(I) and copper (II) ions by ammonia in alkaline media. Ammonia has become more attractive as a lixiviant during the present socio-economic climate due to its low toxicity, relatively low cost and ease of regeneration by evaporation (Meng & Han, 1996). Ammonia processes are ideal when the gangue minerals are acid consuming e.g. with calcareous or dolomitic gangue (Paynter, 1973).

Jan 1, 2019

By T. Sterry Hunt

WET processes for the extraction of copper from its ores have of late attracted much attention, especially in Europe, where the use of oupriferous iron-pyrites as a' source of sulphur prevails. These processes present an additional interest, when, as is often the case, the copper-ore's to be treated contain one or both of the precious metals, since the complete separation of these from copper, either in the dry or the moist way, is, as is well known, one of the difficult problems of metallurgy. The principal wet processes hitherto in use for the extraction of copper from its ores, may be included under three classes : I. Those in which the copper in snlphuretted ores is rendered soluble in water by calcining them, after a preliminary roasting which removes the greater part of their sulphur, with a portion of common salt, by which the copper is converted into a chloride (Long-maid and Henderson), or with sulphate of soda, by which sulphate of copper is formed (Monnier). Allied to these is the method by which a portion of the copper is got as soluble sulphate by roasting without addition, a process which is sometimes modified and rendered more complete by re-roasting the lixiviated residues with the addition of a portion of raw sulphuretted ore (Bankart). II. Those methods in which free chlorhydric or sulphuric acid is used to dissolve the copper- from oxidized or roasted ores. These, while simple and efficient, are too costly, except in certain localities where chlorhydric acid is a waste product. Related to them is the plan which consists in exposing the oxidized and moistened ores to the slow action of sulphurous acid mixed with air, by which a soluble sulphate of copper is formed. It has been proposed to decompose the solutions of sulphate or chloride of copper by sulphuretted hydrogen, thus separating the copper as sulphide, and using the liberated acid to dissolve fresh portions of oxidized copper. It is, however, the almost universal practice to throw down the copper from its solutions, however obtained, in the metallic state by the use of iron, either scrap-iron or iron-sponge; although in some instances

Jan 1, 1882

By A. D. Pittuck

Refractory minerals .containing uranium and columbium have been successfully treated at Queen's University, Kingston, to produce high-grade concentrates. Recoveries of over 90 per cent of the uranium and columbium have been obtained when treating concentrates containing euxenite, fergusonite, betafite, and pyrochlore. The processes ?consisted -of leaching the columbium with hydrofluoric acid, and the uranium with nitric acid, sulphuric acid, or sodium carbonate solutions. The leach liquors were extracted with organic solvents from which the values were stripped and precipitated, producing final products assaying 45 to 55 per cent Ob2O5 and 60 to 70 per cent U3Os.

Jan 1, 1958

By Richard W. Herzer

Introduction Thermal methods have been predominant in the metallurgy of the base-metals ever since the foundation of the industry thousands of years ago. Development in hydrometallurgy has been confined largely to the past six or seven decades. It was at first necessarily slow, then for a number of years rapid progress was made. At present, the uncertain state of the metal industry has curtailed research on the subject; however, in the next few years one may assuredly look forward to an ever increasing use of wet processes. Leaching methods have attained the dominant position in the case of certain metals, notably zinc. For others, they have made no headway at all. Copper presents a situation all its own. Due to the varied nature of its ores and to the large number of localities in which it is mined, a great number of processes have been developed, each suitable in its own particular field. In the case of this metal, therefore, it is more difficult to arrive at definite conclusions. It is the purpose of this paper, first, to review the outstanding developments of base-metal hydrometallurgy, and, secondly, to consider possible Canadian applications of some of the processes in use. The discussion will be confined to the three metals, zinc, lead, and copper, with particular reference to the last mentioned. Zinc General Note At present, zinc is produced by two principal processes, the older retort process and the electrolytic process. The latter is a development of the last twenty years and, judging by its present progress, bids fair eventually to supplant older methods in a large proportion of cases. It is certain, at least, that in the erection of a new plant, a hydrometallurgical process would be given the most serious consideration

Jan 1, 1934

By B. J. Arnold, F. F. Aplan

"The captive bubble and sessile drop contact angle techniques have been used to evaluate the hydrophobicity of petrographically identified coal macerals. The magnitude of the contact angle on vitrinite is shown to be a function of coal rank ( %C, %0 and % vitrinite reflectance), increasing with increasing rank to a maximum for low volatile bituminous rank coals then decreasing for anthracite. The least data scatter is found when vitrinite reflectance is used as a rank parameter. Contact angles measured or estimated for other discrete macerals, indicate a variation in hydrophobicity based on coal type. These data correlate well with known flotation results. Coal is a heterogeneous mixture of plant remains and associated minerals that have undergone chemical and physical changes by biological and geological processes. These plant remains form coal macerals, discrete organic entities in the coal which have characteristic chemical and physical properties. As the differing chemical nature of each maceral will affect its behaviour in processes such as carbonization the separation of macerals during coal preparation may provide an important asset in coal utilization.Separation of some of the macerals based on specific gravity has been achieved in the laboratory, and by Sun and Cohen , in a study of two Pennsylvania bituminous coal preparation plants. It was indicated that the lighter exinites report almost exclusively to the clean coal while the denser inertinites are concentrated in the higher gravity fractions. The Longwy-Burstlein process6 took advantage of the difference in hardness or friability of the macerals by using impact breakage followed by screening to remove the more friable vitrain,clarain and fusain. The remaining durain, often high in spore content, was then recrushed. The re-mixed fractions produced the optimum size for obtaining the best coke strength for a particular coal.Separation or concentration of rnacerals by froth flotation largely depends on differences in the hydrophobicity of the macerals. Sun and Cohen 5 . indicated an order of decreasing floatability as exinite, vitrinite and inertinite, while a Russian study (as cited by Brown 7 ) showed that vitrains floated more readily than the duller types. More recently, Hirt and Aplan8 showed that 'pseudovitrinite' floats somewhat faster than does vitrinite while fusinite floats very slowly. Olson and A plan found that vitrinite floated easily while substantial quantities of the non-floated coal contained fusinitic material. Burdon noted that vitrinite is more floatable than semifusinite, and Sarker et a/. 11 have shown that vitrinite is selectively concentrated by froth flotation and oil agglomeration. Sun indicated that flotation, based on his calculated floatability index, should decrease in the order clarain, vitrain, fusain and durain; while Horsley and Smith 13 indicated the decrease to be: vitrain, clarain, durain and fusain. These differences can be attributed to the specific coals used in each study and to the individual macerals composing each coal type within a given coal."

Jan 1, 1989

By L. C. Graton

INTRODUCTION CLEAR recognition by Lindgren1 of the important control exerted by physical conditions on the formation of ores laid the foundation for that long-sought attainment, a classification? of mineral deposits based consistently on origin. In probably no other portion of the realm of mineral deposition has the importance of physical influences been so manifest or the validity of Lindgren's genetic classification been so well proved as in that extremely important family of ore deposits which are now universally classed as hydrothermal and generally regarded as of magmatic derivation. While many varieties of deposits, some of them possessing great scientific interest and economic importance, fall outside the limits of this hydrothermal group, inclusion within it of most of the veins and a vast number of the replacement deposits of the world gives to this grand genetic class at once a dominating importance and a powerful scientific appeal. Nowhere are these vein and replacement deposits of hydro- thermal lineage more generously distributed, more extensively exploited or more thoroughly known than in the Cordilleran province of the United States. One of the respects in which these hydrothermal deposits command the interest of both geologists and mine operators is in their relation to depth: the great vertical range through which many of them have already been developed, the disappointing failure of others at relatively shallow levels, and the noteworthy downward changes in character and tenor exhibited by still others. In this hydrothermal category (if the Rand be granted place here) are included the ores of all the very deep mines of the world, ranging from 5,000 to nearly 8,000 ft., as well as of many other mines undoubtedly destined in time to attain corresponding depths.

Jan 1, 1933

The source and origins of the hydrothermal fluid of orogenic, or quartz-vein lode-gold deposits remain uncertain and disputed: Resolution would likely have a significant affect on exploration strategy. The gold-carrying fluid is interpreted to be a low-salinity mixed aqueous-carbonic fluid of near ambient pH and redox state, that may be derived either magmatically through crystallisation of granites at depth in the crust or through metamorphic reactions in a range of rock types. Classic isotope and geochemical tracers of fluid source have not differentiated these possibilities and have produced conflicting conclusions. Mass balance calculations show that fluid composition will be modified by wallrock reaction between source and deposit, and tracers should not be expected to give clear indications of fluid source. Fluid oxidation state can also be modified along flow paths. Tracers should be interpreted with care and with view to possible modification. Modelling of full flow paths and examination of the hydrothermal histories of proposed source rocks may be alternative approaches to determining fluid source.

Jan 1, 2002

By Gerardo Fuentes

The stoichiometry and kinetics of the hydrothermal reaction of enargite in copper sulphate solutions was investigated in the range 140–300 °C, and particle size under 15 and 50 μm. The stoichiometric experiment was performed at 300 °C, using one gram of sample (96% enargite, 4% tennantite) and 100 mL of copper sulphate solution (10 gpl) at 1.2 pH. Enargite reacted completely after 80 min. A solid product was obtained and characterized by XRD. The main product was chalcocite (M) and some traces of djurleite and chalcocite (Q). Tennantite did not react under these conditions. The kinetic experiments were conducted at 140 °C, 190 °C, 250 °C and 300 °C. From fitted data and Arrhenius equation the calculated activation energy (Ea) was 51 kJ/mol. According to the hydrothermal transformations studied, the reactivity of different phases found in the copper concentrates was bornite > chalcopyrite > covellite > sphalerite > pyrite > enargite > tennantite.

By R. W. Raymond

[NOTE.-This paper was presented at the Pittsburgh meeting in a partially completed form, and I fully expected to obtain, before the period of its publication, both the data and the leisure required for its completion. The hygiene of collieries was to have been separately considered, and the sanitary conditions of various metallurgical industries connected with mining were to receive an extended discussion, while at the same time the points treated in the paper as presented at Pittsburgh were to have more ample illustration from American and European sources. Delay in obtaining expected information, and the continuous pressure of professional work, have prevented hitherto the execution of this plan ; and I can now publish only a portion of what I had designed to contribute to the subject. In the meantime this portion, but slightly changed, has been included as a chapter on the Hygiene of Metal Mines in a work entitled Hygiene and Public Health, edited by Dr. A. H. Buck, of New York. An excellent chapter on the Hygiene of Coal Mines, contributed to the same work by Mr. Henry C. Sheafer, of Pottsville, affords me an opportunity to supply one of the deficiencies above alluded to. Mr. Sheafer's chapter contains much material in the way of explanations not required by professional readers; and he moreover devotes a considerable space to the explosive gases, falls of roof, and other causes of accidents to life or limb-a branch of the subject not belonging strictly to hygiene. I therefore make extracts only from his treatise.-R. W. R] IT is convenient to divide mines, with reference to this subject, into two classes, collieries and metal-mines. Subterranean quarries, rock-salt mines, etc., present no conditions requiring them to be separated from the latter class.

Jan 1, 1880

By John E. Appleby

Grupo Acera, HYLSA originated in 1942 when Hojalata y Lamina began operations to supply steel strip for the Cuauhtémoc Brewery, the parent company. Production began in April, 1943 with growth accelerated because of World War II. By 1948 the company was producing its own steel from electric furnaces charged with scrap iron. Further expansion was hampered by the fluctuations in availability, price and quality of scrap iron. As a result, a search was made for a suitable process to enable HYLSA to satisfy its need for a reliable source of feed for its steel-making facilities. Their relatively low iron requirements at the time did not justify the very high capital investments required using the conventional blast furnace method. Processes which by-passed the blast furnaces were very much in their infancy and were not commercially tested. HYLSA then undertook an investigation of the direct reduction processes, knowing full well that much larger companies in other countries had investigated several processes but their work had not established a trend toward direct reduction or any particular process. However, after much effort and persistence HYLSA arrived at a process which suited their conditions. By using Mexican iron ore and natural gas they developed the now well-known H. y L. Process to produce sponge iron. Concurrent with the start of investigations which led to the H. y L. Process, HYLSA began an exploration program for a suitable source of iron ore. Thus in 1951, Cia. Minera Las Encinas, S.A. was formed, culminating in the development of the ore bodies near Pihuamo in the State of Jalisco. Mining commenced and ore was

Jan 1, 1971

By M Fuchs

Roadheaders were introduced to mining and tunnelling in the early-1960s. Since this period their main field of application has been for roadway drivages in coal mines and for mineral production, for instance in coal, rocksalt, potash and gypsum mines. Their main field of application in tunnelling is in softer and partly friable rock formations, a suitable area for non-blasting excavation. However, in longer tunnel sections excavators with high ripping forces have replaced roadheaders in many situations. The main arguments against roadheaders have been and still are:a sharp increase in excavation cost with increasing rock strength and increasing abrasivity; and greater risks in conditions which have not been, or could not be, investigated to a sufficient extent. These arguments have, to a great extent, restricted practical roadheader application in production mining and in a lot of æstandardÆ tunnel projects. To address these limitations, factors which could make roadheaders attractive for such conditions and could overcome existing technological difficulties have been investigated. Extension of the economic field of application in hard rock was found to be the most important and the most effective step forward.

Jan 1, 2002

By G. R. Mansfield

THE geologic structure of the Idaho phosphate field has an important bearing on the classifica-tion and the exploitation of the lands that contain phosphate. Maps and structure sections1 recently made available show that the phosphate is mainly preserved in downfolded or downfaulted areas. Since the phos- phate beds are generally rather soft and lie between harder and thicker formations they have been forced to adjust themselves to these formations during epochs of disturbance of the earth's crust, so that they are frequently affected by minor folds and by faults which may help exploitation by producing undue thickness at certain places or may hinder it by cutting out the phosphate at other places where according to expecta-tion it should be present. A particularly important structural feature which affects both the distribution and the exploitation of the phosphate beds is the great Bannock overthrust by which one or more slices or blocks of the earth's crust have been pushed over other blocks like ice cakes on a pond. Thus in this region a great flap of folded rocks lies across other rock structures for a distance of approximately 35 miles. Most of the commercially accessible phosphate deposits, shown in the maps that accompany the report cited, are in this flap, which con-stitutes the upper block of the overthrust. In some places, however, the phosphate in the lower block is also accessible. In both blocks post-phosphate rocks exposed at the surface here and there indicate that phos-phate beds may lie at greater or less depth below. The presence of the Bannock overthrust beneath much of the area reported is known by the occurrence of windows in the upper block both within the quadrangles mapped and in areas farther west. The portions of the lower block exposed in these windows show stratigraphic dis-cordance with the surrounding formations. The depth of the fault plane at most places, however, can be told only by drilling or actual mining. Some of the folds in the upper block are relatively deep and it seems fairly certain that the phosphate that may originally have been present in their deeper parts has been removed by thrust faulting. The estimates of phosphate reserves in folds liable to be affected by the Bannock overthrust may therefore need ultimate revision depending on the depth of this fault plane, as it may subsequently be ascertained.

Jan 1, 1928

By Frederick Laist

IT IS obviously impossible to design a copper smelting plant which could be considered, ideal under all conditions. For example, a plant properly designed to smelt the concentrates resulting from the treatment of porphyry ores by flotation would not be the kind of plant that one would design for the treatment of a coarse pyritous copper ore. The smelting of flotation or other fine concentrates is generally accomplished in a reverberatory plant, although occasionally the concentrates are sintered and then smelted in blast furnaces. On the other hand, coarse self-fluxing ores, whether pyritic or semi-pyritic, are generally smelted in a blast-furnace plant. Even this statement is only- approximately true, since the choice will be influenced by many factors; for example, the kind of fuel available (coal, oil or coke), the value of power (electric or steam), the uses to which waste steam may be put (heating, drying, etc.), the cost of refractories (silica and fire brick), the permanence and continuity of the ore supply and the relative first costs of the two types of plant (the reverberatory plant is generally more expensive to construct and to shut down), the contour of the site on which the plant is to be built (rugged, rolling or flat), the intelligence and character of the labor avail-able, and many other factors. From these remarks, it is evident that no copper smelter can be ideal for all conditions of ore and locality, and inasmuch as most of the copper now being pro-duced comes from fine or flotation concentrates and these are generally smelted in reverberatory furnaces, I shall confine my remarks to this type of smelt-ing plant.

Jan 5, 1923

By Phillips J. T, Hutton W. A

This paper discusses why Argyle Diamonds adopted thickened tailings disposal in 1985 and how the system was developed over the next seven years. A wide ranging review of the thickened tailings disposal system and the tailings deposit it was producing was carried out during 1992 and 1993. The results of that review led to the abandonment of thickened tailings disposal at Argyle because of serious concerns about the stability of the resultant tailings deposit. The exposure to the risk of liquefaction and an assessment of the consequences if this happened to the thickened tailings deposit are shown. The criteria for selecting the new tailings disposal strategy are then described. The final section of the paper discusses the management strategies that have been put in place. These are designed to: ò reduce the risk posed by the thickened tailings deposit to an acceptable level; ò safely store tailings to the end of mine life; ò produce a land form that can be decommissioned to the satisfaction of Government regulatory bodies.

Jan 1, 1996

By Mine F. J, Woods P. H

Waste waters from mine sites may be screened for all constituents of significance as a first step in establishing water quality criteria for protection of aquatic ecosystems. Inductively-coupled plasma mass spectrometry (ICP-MS) offers the capability to screen water samples and digest solutions for virtually all elements in the periodic table. This technique is therefore able to identify elements whose concentrations are elevated in waste waters and may be rarely present in background surface or ground waters. Such elements may be present in waters at only a few ¦g/L. ICP-MS has sufficient sensitivity to detect sub-¦g/L or even ng/L levels of many elements. Waste waters from most mining activities in the Northern Territory have been examined by ICP-MS. The concentration data provide the basis of a contaminant data base. The data can show specific tracers in the ore or mineral itself, subsequently found in waste water. An example is the presence of rhenium in uranium ore. Such trace elements may be present in soluble or insoluble chemical forms and are more useful as indicators when soluble and exemplifying conservative chemical behaviour. The approach described, based on ICP-MS screening of mine site waste waters, together with general physico-chemical variables, provides the most comprehensive means yet available of selecting variables for monitoring programs and ensures that key elements and/or their chemical species are not excluded from monitoring programs.

Jan 1, 1994

By P. H. Davenport

"The ore-element patterns in regional lake sediment geochemical survey data from four areas of Newfoundland containing granite-related mineralization reveal characteristics which can be used to identify mineralized granitoids. The St. Lawrence Granite with its fluorite deposits, the Ackley Granite with molybdenite mineralization, the Grey River tungsten lodes and the newly discovered Mo- W mineralization at Granite Lake, Central Newfoundland, have all been included in the regional lake sediment geochemical surveys, which now cover about 90,000 km2 of the Island of Newfoundland. The lake sediments over these mineralized granites show diagnostically high values of Mo, F, Pb, Zn and, in three of the four cases, U, together with locally high Ag and Cu values. Furthermore, these ore-metals display a distinctive zoning pattern about the intrusives, which can be related to the classical ore-metal zoning patterns developed about many types of mineralized granitoids.IntroductionGranitoids in Eastern Canada are receiving renewed interest as potential hosts of tin, tungsten, uranium and related mineralization. Although a number of mineral deposits and showings have been known from a few granites (sensu lato) in the region for many years, granites had generally received little attention as potential hosts of metallic mineralization until the recent resurgence in interest in the mineral potential of these rocks. In Newfoundland, where granitoid rocks underlie about 30070 of the land area, petrologic data are available from some granites, particularly those in eastern Newfoundland (Strong et al., 1974; Strong, 1980), but in southern and southwestern Newfoundland the data base is sparse, and indeed in many cases individual plutons are not delineated on the geological maps of the area . Most of the available data on Newfoundland graniotolds have been summarized by Strong (1980), who gives a general assessment of their mineral potential by comparing their physical, mineralogical and petrochemical attributes with mineralized granites elsewhere in the world."

Jan 1, 1982

The uranium ore processed by the McClean Lake Operation is co-mineralized with niccolite, rammelsbergite and gersdorffite (nickel arsenides). During the strong acid leach process, the majority of the reduced arsenic is oxidized and liberated into the lixiviant solution as As+5. Following uranium extraction, the residual raffinate solution, enriched in As+5, reports to the tailings preparation process. In this process, Fe+3 is first added and mixed with raffinate (pH~1) to achieve a Fe+3/Astotal molar ratio of at least 3. This is followed by two stages of neutralization, using slaked lime to achieve a solution pH of 4, and approximately 7.5, respectively. Synchrotron based x-ray absorption techniques (XANES, EXAFS) applied to neutralized raffinate solids and tailings samples collected from the tailings management facility provide evidence that arsenic has been precipitated as a poorly crystalline form of the mineral scorodite (FeAsO4?2H2O).

Aug 1, 2010

By H. Y. Sohn

The ignition and combustion of chalcopyrite particles under suspension-smelting conditions, in which the particle-particle interaction is neglected, has been analyzed by considering the intrinsic kinetics of oxidation and heat and mass transfer between the particle and its surroundings. Reasonable agreement has been obtained between the calculated results and experimental results from a laminar-flow furnace reported in the literature. This computation- al procedure has further been applied to flash-smelting conditions. Under suspension-smelting conditions (Sh = Nu = 2), chalcopyrite particles ignite at around 960 K. The effects of various parameters, such as oxygen concentration, particle size, and gas temperature, have also been evaluated. Comparison between calculated results in this work and experimental observations indicates that SO is the important gaseous product that forms at the surface of the particle.

Jan 1, 1993