Search Documents

By D. J. Salley, M. Burke, G. L. Simard

A PREVIOUS paper1 described experiments on the interaction of dithiophosphate with galena and suggested that the process had characteristics of both chemisorption and surface chemical reaction. In order to add to this knowledge, a study has been made of the competition of a second ion with the dithiophosphate for a galena surface. Use again was made of radiotracer methods. Since radiophosphate was readily available, it was chosen as the competing ion. It is a feature of the tracer method that, in spite of similarity in the chemical behavior of the ions, the measurement of one ion in the presence of another can be made as simply as if only one ion species were present. In order to study the competition effectively, it naturally was necessary to examine the interaction of phosphate with galena alone as well as in the presence of dithiophosphate; consequently, this report consists of two parts. The first is devoted to a study of the rates, the isotherms, the desorption, and the exchange in the phosphate-galena system in the same manner as was done for the dithio-phosphate-galena system. The second portion is concerned with the studies on the competition. Flotation experiments on the three-component sys- tem are also outlined and the relationships involved are discussed briefly. The experimental method was that of the previous paper. The preparation of the galena, the radioactive and the nonradioactive dithiophosphate, and the procedures for radioactive counting, for sorp-tion, and for flotation measurements were identical to those previously employed. The radiophosphate* was used in the form of disodium phosphate solution. In the competition experiments the effect of the phosphate ion in depressing dithiophosphate uptake was followed by holding the initial concentration of nonradioactive phosphate constant while varying the initial concentration of the radioactive dithiophosphate. Counterfashion, the depression of phosphate by dithiophosphate was obtained by keeping the initial nonradioactive dithiophosphate ion concentration constant and altering the initial concentration of the radioactive phosphate ion. All solutions were buffered at a pH of 7.8 a0.1 by 20x10-' mol per cc of sodium carbonate. Measurements with radiocarbonate indicated that little or no carbonate could be sorbed in the presence of

Jan 1, 1952

By H. M. Fisher, R. E. Snow, S. C. Sun

RAPID colorimetric estimation of the amount of transparent minerals has been applied successfully to instream mill products at the cryolite flotation plant of the Pennsylvania Salt Mfg. Co., Na-trona, Pa. Apparatus used was a Bausch & Lomb monochromatic colorimeter. The optical assembly of the colorimeter, shown in Fig. 1, passes light from a 6-v, 32-cp, prefocus incandescent lamp through two condensing lenses, a heat-absorbing filter, an aperture-control diaphragm, an interference filter, and a sample container. The light finally strikes a barrier-layer photocell and the output current of the photocell is measured by a galvanometer. The engraved index line of the collective lens, located between the lamp and the galvanometer, is imaged upon the transmission scale of the colorimeter by the galvanometer mirror. Reflected light from the galvanometer mirror, the position of which is determined by the output current of the photocell, enables reading of percent transmission throughout the mineral sample being tested. The colorimeter, Fig. 1, was standardized as follows: 1—The colorimeter was' connected to a Sola constant voltage transformer, A, of 6-v output operated from a 110-v ac supply. 2—Turning the toggle switch, C, illuminated the transmission scale, B, and the control knob, D, was adjusted with a blank in the optical system until the hairline on the scale read zero. 3—A No. 430 interference filter, E, was inserted into the filter channel, F. 4—A test tube, G, containing 7 cc of distilled water was pressed into the sample space, H, and the fine adjustment knob,

Jan 1, 1955

By A. P. Herring, S. F. Ravitz

Manganese dioxide reacts with sulfur dioxide solution in the absence of oxygen to form manganous ion and dithionate ion, together with a small amount of sulfate ion. The effects of SO, concentration, hydrogen ion concentration, stirring speed, and temperature on the rate of the reaction were studied. If it is assumed that undissociated SO2 molecules react rapidly with MnO, at active surface sites and that bisulfite ions react slowly, an equation can be derived that satisfactorily represents the observed reaction rates. For solutions in which the reaction with undissociated SO2 predominates, reaction rates and activation energies estimated from diffusion theory agree well with those found experimentally. Manganese dioxide dissolves readily in aqueous solutions of sulfur dioxide. The nature of the dissolution reaction was first investigated by Gay-Lussac and welter1 in 1819; they identified a new oxysulfur radical, the dithionate ion, S2O6, as one of the reaction products. Since then, the reaction has become a standard method of preparing dithionates in the laboratory and one of the more promising methods of leaching manganese from low-grade ores. In the "dithionate process" developed by the U.S. Bureau of Mines2,3, manganese-dioxide ore is suspended in a solution of calcium dithionate and leached with gas containing sulfur dioxide. The manganese is dissolved and the sulfur dioxide is oxidized according to the reactions:

Jan 1, 1965

By F. H. Bunge, W. H. Jr. Dailey

This paper covers an investigation of the magnetic roasting characteristics of several iron ores when treated in the form of small pebbles (1/8" to 1/2" in size), and in the form of balled fines. The roasting treatment was carried out with apparatus capable of very close control of time, temperature, face velocity, and per cent combustible in the reaction gas. The effect of rapid initial heating was not studied although it is believed that this has a major effect on the results when processing solid pebbles. The test equipment is a fixed bed type, with such a wide range of gas velocities, temperature, combustible content and etc., that the results obtained can be made to approximate those from various commercial roasting processes. The data presented in this report covers reaction rates, effect of processing variables on silica in the concentrate, flow resistance of beds, and to a limited extent, the effect of particle size. The work is far from complete and raises many more problems than it gives answers for. We believe, however, that much of the information will be useful to those engaged in investigation of magnetic roasting processes. We measured the time required for various degrees of completion of the reaction of hematite to magnetite, for finely pound material in balled form and for carefully sized crushed ore. Measurements were made for a variety of Minnesota and Labrador ores. This report covers an ore requiring fine ginding (-325 M) for liberation (Lower Kevin) and an ore requiring coarser grinding (-100 M) for liberation (Upper Olson). These ores represent approximately the range of characteristics for a group in the west end of the Mesabi Range. Our data shows that the reaction rate is very fast for the finely ground material and that tests are readily reproducible. The reaction rate of the larger size material, in the range of 1/8" to 1/2", is much slower and the results more erratic. We found that for the ores tested, the percentage of silica in the Davis tube magnetic concentrate was a function of iron recovery. Normally the concentrate had a low and relatively uniform percentage of silica up to a critical value of recovery. Above this critical recovery value the percentage of silica increased very rapidly. For example, for one particular ore roasted in the form of 3/8" diameter balls made from -28 mesh fines, the silica remained at about 7% to 8% until the recovery exceeded about 93%. Increasing the recovery beyond 93% resulted in over 25% silica with the additional recovery. The degree of conversion of hematite to magnetite (% completion of roast) required to obtain 93% recovery for balled material, varied inversely with the gas flow rate. Percentage silica in the concentrate resulting from roasting 3/16" and 1/2" solid material was relatively erratic. However, for varying degrees of iron recovery the lowest silica percentages corresponded to those obtainable when roasting fines. It is thought the probable reason for the erratic values was the variation in shattering or cracking of the pebbles during initial heating and roasting. There is some indication that very fast initial heating will greatly increase the permeability of the individual particles of ore to gas flow so that the reaction rate for coarse sizes will approach that obtainable with balled fine material.

Jan 1, 1961

By E. Martinez, M. L. Hollander

Sulfurization of copper oxides was investigated using differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). When mixtures of cuprous oxide and sulfur were heated, the surface of the cuprous oxide was converted to cupric sulfide at the melting point of sulfur (119°C); additional sulfuriza-tion took place when the sulfur vaporized above 200°C. Cupric oxide also reacted with sulfur to form cupric sulfide but the starting temperature was approximtely 225°C. When the cupric oxide and sulfur were physically separated, sulfurization still occurred indicating a reaction with the vapor phase of sulfur. Sulfurization was obtained in runs with cuprite, chrysocolla and basic copper carbonate. Further decomposition of cupric to cuprous sulfide was noted at somewhat higher temperatures. The DTA-TGA data made it possible to set up postulated reactions in each system. These techniques were found to be well suited for studies of this type. Because of the difficulty encountered in flotation of many metallic oxide minerals, efforts have been made to find suitable reagent combinations to sulfi-dize these minerals in the flotation circuit. Recently Bautista and Sollenberger,1 using a different approach, showed that the surfaces of many oxide minerals can be converted to sulfides by heating them with sulfur to temperatures between 150° and 225°C. This technique was effective even with chrysocolla, a mineral which is very difficult to sulfidize. The primary objective of the present study was to obtain information on the reactions between the oxide minerals and suIfur. A better understanding of the mechanisms involved might prove useful in pointing the way to improvements in the process. A secondary objective was to demonstrate the value of thermo-analysis in investigations of this type. Bollin and Kerr2 have reported the use of a modification of differential thermal analysis to study the thermal reactions which occur on heating elemental constituents in a closed system to synthesize copper sulfides and iron sulfides. Differential thermal analysis (DTA) was used to detect enthalpic changes as mixtures of minerals and sulfur were heated. The DTA indicated the temperatures at which reactions started and whether they were exothermic or endothermic. Weight change data were obtained by thermal gravimetric analysis (TGA). These data could often be used to support postulated reactions and to eliminate consideration of other possible reactions. EQUlPMENT Differential thermal analysis (DTA) involves measuring the temperature difference between the sample being tested and a thermally inert reference, generally aluminum oxide, as both are heated at a uniform rate. Series connected thermocouples in the sample and reference are used to detect the temperature differences that occur as the sample undergoes enthalpic changes.3 Crystalline transitions, fusions, vaporizations, dehydrations, decompositions and many solid state reactions can be detected. A DTA unit manufactured by the Robert L. Stone Co., Austin, Tex., was used in this investigation with a temperature rate increase of 10°C per min. The samples and reference were placed in Inconel cups with wells in the bottom into which the differential thermocouples fit, and run in a nitrogen atmosphere. Thermal gravimetric analysis (TGA) measures the changes in weight that a sample undergoes as it is heated. A Chevenard thermobalance converted electronically for graphic recording was used with a heating rate of 5°C per minute. An X-Y recorder plotted temperature on the X axis and weight changes on the Y axis. A flow of 250 ml per min of nitrogen was streamed through the furnace. Samples were placed in a porcelain crucible open at the top, unless otherwise noted. EXPERIMENTAL RESULTS Sulfur: Two sulfur samples from different sources were tested. The DTA of a high purity sulfur and a reagent grade sublimed sulfur are shown in Fig. 1. Nearly all the tests were run with the sublimed sulfur due to the difficulty in grinding and high electrostatic charge encountered with the high purity sample. Sulfur undergoes three endothermal reactions below 200°C. The first is a solid-solid crystalline transi-

Jan 1, 1964

By C. H. Bushell, M. Malnarich

REAGENT control in flotation is more an art than a science. Operators vary the amount of reagents used according to the metallurgy obtained. The amount of collector may be increased, for example, if tailings losses are high and decreased if concentrate grades are low. Often control decisions are based on only one assay, such as the tailings analysis by microscopic or quick assaying techniques. This method of control has given reasonably satisfactory results but is clearly far from perfect. No adjustment in reagent addition can be made until information is received that metallurgy has deteriorated, that is, until the optimum time for adjustment has already passed. The first adjustment probably will be an overcorrection or an undercor-rection, and time will be required to establish the best new addition rates. The seriousness of the problem depends on variability of the ore. In ore of continually shifting composition and complexity, mill results will shift back and forth from excessively high concentrate grades and high tailings losses to excessively low concentrate grades and low tailings losses. Economic incentive for improved control also increases with increasing size of mill. Even a small improvement in percentage recovery in a large mill pays for expensive control equipment. The Sullivan concentrator of The Consolidated Mining & Smelting Co. of Canada Ltd. at Kimberley, B. C., treats 11,000 tpd of complex lead-zinc ore, a scale of operations which justifies an extensive research program on automatic control. Work has followed two separate lines: 1) development of faster analytical methods for pulps, and 2) study of metallurgical control by measurement and control of reagent concentrations in flotation solutions. Although work is incomplete in both phases, a progress report can be given on the development of reagent control. The primary requirement in flotation is selective attachment of particles of one mineral to air bubbles while the other minerals remain detached. Hydrophobicity of mineral surfaces is established by the collector, which coats the minerals by adsorption. The requirement of selective affinity for air bubbles is therefore equivalent to requiring selective adsorption of a collector. Adsorption theory, supported by abundant experimental evidence, shows that the amount of a reagent adsorbed at a surface must be a function of its residual concentration in solution. As the concentration in solution rises, the amount adsorbed increases, usually following some type of logarithmic curve. Selective adsorption must depend on a natural difference in the minerals, but ordinarily the differences in affinity are not great enough to obtain complete coverage of one mineral while the others are completely uncoated. An optimum concentration will exist, below which coverage will not be great enough to float enough of the valuable mineral, and above which too much unwanted mineral will float. All other reagents, except some frothers, affect the amount of collector adsorbing at constant residual collector concentration in solution, either by adsorbing competitively with the collector, or by a

Jan 1, 1957

By E. A. Lowe, C. C. Walden, D. W. Duncan, P. C. Trussell

The industrial importance of bacteriological leaching is finding increasing application in the treatment of low-grade ore and of exhausted mines. Another possible application of bio-leaching is in the extraction of ore concentrates, although here the first uses will likely be for ores that are refractory to conventional metallurgy. Biological leaching studies, which have been in progress for over ten years in this laboratory, are directed towards the more rapid and complete leaching of metals from their sulfides, particularly of copper, zinc, and nickel. The program involves not only laboratory studies, but pilot-plant operations and liaison with mining companies which are underwriting the cost of the investigations. Although the bacterial oxidation of mineral sulfides to the corresponding water-soluble sulfates has been occurring naturally for many centuries, the first documented recovery of solubilized metal from pregnant solution was at the Rio Tinto operation in Spain. At this location leaching variables were studied extensively in an empirical manner some years before the causative bacteria were first isolated from other leaching situations. Documentation of the presence of bacteria in leach waters of the Rio Tinto, Spain, operation was not confirmed until 1962.3 On this continent leaching was pioneered in the early

Jan 1, 1968

By K. W. John, L. Mueller

The rock mechanics approach to stability problems of steep rock slopes dealt with in this article has particular reference to the concept and methods of analysis developed by the Salzburg, Austria, school of rock mechanics. These theories are based on the principal postulate of the anistropy of the technological properties of in-situ rock masses. The present approach, which is more in line with geological engineering concepts, differs distinctively from the soil mechanics and engineering geological concepts commonly practiced in the United States. PRINCIPAL ENGINEERING APPROACH TO GEOTECHNICAL PROBLEMS AVAILABLE ENGINEERING CONCEPTS At the present time two different concepts are available for engineering analysis of underground structures and foundations in different earth material: soil mechanics and rock mechanics. Of these two branches, soil mechanics is, after 40 years of initially revolutionary and more recently evolutionary development, well established. Its principal concept of replacing in-situ soils or rock masses by perfectly homogeneous substitues has resulted in many practicable methods of analysis which, if applied with judgement, have resulted in successful applications throughout the world. Conversely, rock mechanics is still in the early stages of revolutionary development. Few, if any, standard concepts or methods of analysis have been established. Opinions vary widely. At present the professional background of an investigator very often determines the engineering concept to be applied to a particular problem, rather than its actual requirements if considered objectively. This paper stresses the rock mechanics or geological engineering point of view advocated by the Salzburg group. A review of the soil mechanics concepts is considered beyond the scope of this paper. THE "SALZBURG APPROACH" TO ROCK MECHANICS The rock mechanics concept as outlined here was initiated by Stini, based upon work by sander.1 Since

Jan 1, 1963

By B. S. Crocker

Pebble grinding was used at Lake Shore Mines in 1949. A full description of experimental evidence and test plant results was published in 1952 1 and further operating details in 1954 2.' In more recent years the term autogenous grinding has been coined and is now frequently seen in the technical literature. Autogenous crushing and/or grinding covers any grinding mill in which the larger pieces of ore are made to grind the fine pieces; the crushing and grinding may be either dry or wet. This discussion deals only with wet autogenous grinding, and only in the fine grinding part of the circuit, i.e., from 6 mesh to 92 pct —325 mesh. The term pebble milling could apply to fine autogenous grinding in which the pebbles are made from the ore itself, or it could also refer to grinding with pebbles, such as flint pebbles, from a source outside the mine. The first mill to use screened ore to grind the ore (1949) was Lake Shore Mines, Kirkland Lake, Ont. Pebbles were screened from the jaw crusher discharge and used in 5x16-ft tube mills which had been converted to 6 ft 8 in. by 16-ft grate discharge pebble mills. The same general practice was adopted the following year by the Wright-Har-greaves Mine in Kirkland Lake. Shortly after that the Neptune Mines in Nicaragua started correspondence with the writer concerning the possibility of using screened ore for grinding in their plant. They first ran some plant-scale tests which confirmed the Lake Shore findings, with the ultimate result that their plant was successfully converted in 1956 to an autogenous grinding plant. In 1957-1958 the writer was successful in converting the grinding plant of Renabie Mines in northern Ontario to autogenous grinding. In all these plants the existing grinding equipment was converted, usually by expanding the diameter of the shell to make the change from steel grinding to ore pebble grinding mills. In all the above plants primary grinding was done in either small rod mills or ball mills. These mills first reduced the pebble mill feed to about 6 mesh. The final grinding in the pebble mills varied from 75 pct —200 mesh to 92 pct —325 mesh. At Lake Shore three stages of pebble milling were used, with a different size pebble in each stage. In other parts of the world interest was being shown in this form of grinding, and in March 1952 the writer had some correspondence with the mill staff at Outokumpu in Finland, with regard to the pebble system used at Lake Shore. At Stockholm in 1957 the account of the crushing and grinding system at Outokumpu was presented in a paper by H. Tanner and T. Heikkinen.V his is the only ore grinding plant that does not use grate discharges on its pebble mills; trunnion discharges equipped with heavy reverse spirals retain the pebble load. The first new mill designed to use screened ore as a grinding medium was the Bicroft mill at Bancroft, Ont., which began operation in the fall of 1956. In 1957 this plant was followed by similar installations at Faraday Mines, also in the Bancroft area, and the North Rankin Nickel Mines in the North West Territories. In 1958 the Dyno Mill at Bancroft was converted, using the same type of grinding. Since all these plants, designed by Kilborn Engineering (1954) Ltd., were originally planned as pebble grinding mills, they were able to take full advantage of the latest design and technique. They are neat in appearance and very efficient and easy to operate. Effect of Ore Hardness: It is often commented that this type of autogenous grinding applies only to hard siliceous ores. Experience has shown that this is not so. In the past ten years the author has had an opportunity to test more than a dozen ores, several of which were quite soft. One of these was shale. The softer ores actually make better looking pebbles than the hard ores. These soft ores produced no particles in the intermediate size range which could not be ground by the largest pebbles in the mill charge. Slabby ores, such as shales, make a suitable pebble for the —6 mesh grinding. It is not generally recognized that for fine grinding a spherical medium is not essential or more efficient. Another frequent objection to pebble grinding is the belief held by some operators that there is too much variation in the hardness of their particular ore to make it acceptable as a source of pebble media. Experience has also shown that this is not a serious problem at all. The variation in pebble load due to changes in the ore hardness is nothing like as great as most people would expect, and the method used for feeding the ore and controlling the the load in the mills is more than sufficient to take care of daily variations in ore hardness. This point

Jan 1, 1960

By L. C. De Jong, J. F. Shirley, M. L. Campbell

At the Southern Peru Copper operation in Toque-pala, Peru, a new process flowsheet for the recovery of byproduct molybdenite has proved successful in dealing with complex and highly variable ores. This paper describes the process and the functions of each of the chemicals used. The Toquepala concentrator of the Southern Peru Copper Corp., located approximately 720 miles southeast of Lima near the Chilean border, processes a daily 36,500 tons of ore containing between 0.50 and 3.00% copper. The molybdenite content in the ore also varies greatly ranging from essentially nothing to as high as 0.15%. The 1966 yearly averages were approximately 1.20% copper and 0.03% molybdenite. The Toquepala porphyry deposit is a mushroom-shaped ore body consisting of a flat-lying enriched zone of predominant chalcocite with a stem-like extension of hypogene chalcopyrite ore in depth within and around a large breccia pipe.1 The underlying chalcopyrite ore has not been reached; however, due to the presence of a major post orebody fault, predominantly chalcopyrite ore is present in the north side of the pit. Most of the molybdenite occurs either as small, fine granular, pulverulent masses in vugs or empty voids of breccia, or as thin foliated masses and narrow thread-like filling in rock joints and fractures. The normal occurrence of quartz gangue with the molybdenite gives some basis for correlating the molybdenite mineralization to some later periods of silicification. On April 18, 1962, the molybdenite recovery plant at Toquepala went on stream, approximately 27 months after the copper concentrator had been in production. The plant was designed and built by Southern Peru Copper Corp.'s engineering and mechanical departments. The basic method chosen for separating the molybdenite from the copper concentrate was the sodium hypochlorite-sodium ferrocyan-ide process. The selection of this process was based on the results of an extensive testing program performed by the Colorado School of Mines Research Foundation, Inc. and the economic conditions of the area. The originally designed process could never be fully implemented due to a lack of sodium hypo-chlorite and has had to be continually modified to satisfy the requirements of the kaleidoscope of ores from the mine. The three basic problems that plague the operation of any molybdenite recovery plant are all present at Toquepala in large measure. Both the quantity of plant feed and the quantity of molybdenite contained in the plant feed are continuously changing. Even more detrimental to an efficient operation is the fact that the copper mineralization in the feed frequently changes once in every 24 hours and occasionally once a shift. From the point of view of the molyb-

Jan 1, 1968

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By N. F. Schulz, H. A. Lex

Representative samples of typical manganiferous iron ores from the Cuyuna Range, Minn., were reduced to the metallic iron state at temperatures just short of fusion with an excess of coke in the manner prescribed by the R-N Direct Reduction Process. The magnetic separation characteristics of the reduced ores were determined at several grinds by means of the Davis magnetic tube concentrator. The efficiency of metallic iron concentration by magnetic separation was a direct function of the degree of metallization of the ore sample and of the degree of liberation at-tained in grinding. Some improvement of liberation characteristics was observed if a small amount of sodium chloride was present in the ore during reduction. Despite the extremely fine-grained structure of the crude ores, iron concentrates containing more than 90% Fe at iron unit recoveries of 90% or more were obtained from well-metallized samples of three of the four ores tested. The rejection of manganese to the tailings corresponding to such separations varied from 80% to 95%. The manganiferous iron ores of the Cuyuna Range in Minnesota constitute a major potential source of manganese in the United States provided that a technically and economically feasible process can be devised for its recovery. The iron and manganese of these ores occur chiefly as oxides or carbonates in the form of very small, irregular, discrete grains embedded in a matrix of quartz or other siliceous gangue. The present investigation was undertaken as a sequel to an earlier one conducted jointly by the U.S. Bureau of Mines and the R-N Corp. Those results indicated that the R-N Direct Reduction Process might be a means for recovering both a metallic iron concentrate and a marketable manganese by-product from Cuyuna manganiferous iron ores.' The R-N Process is one of many proposed schemes for the treatment of crude iron ores involving the di- rect reduction of the iron content to metallic iron followed by physical extraction of the iron. In general, it consists of the following stepsz2 1) Iron oxide in crushed natural ores or iron concentrates is about 95% reduced to metallic iron at elevated temperatures but without fusion in a rotary kiln in the presence of a solid carbonaceous fuel (usually coke), limestone, and combustible gases. 2) The kiln product, consisting of metallized ore, excess coke, lime, and coke ash is cooled. The excess coke is recovered and recycled to the kiln feed. 3) The metallized ore is ground for liberation and magnetically separated to yield a metallic iron concentrate. 4) The metallic iron concentrate is agglomerated by a briquetting process. The reduction kiln temperatures must be high enough for the desirable chemical reactions to proceed at practical rates without causing fusion of any of the ore constituents. The time of exposure of the ore to the reducing conditions must be ample for 90% to 95% metallization of the iron oxides present. If more than 95% of the iron is metallized, however, difficulties may arise in the briquetting step, presumably because of accompanying higher carbon content of the metallic iron. Although not always necessary, powdered limestone or dolomite may be added at some time during reduction to aid in eliminating sulfur and phosphorus from the final product.3,4 Other additives, notably sodium chloride, when present during reduction, result in improved liberation characteristics of the metallic iron from certain reduced ores.'v6 Under the reducing conditions normally employed, P(co + H2)= 1 atmosphere, and 1830' to 2150°F (1000° to 1170°C), the R-N Process is claimed to be adaptable to a wide variety of low, medium, and high grade iron and manganiferous iron ores. The manganese oxides and carbonates are reduced mostly to nonmagnetic MnO which is rejected to the nonmagnetic tailings. Manganese-bearing pyroxenes of the original ore, however, may be rejected poorly, if at all, apparently because these minerals occur within the iron lattice or are entrapped in the metal particles.7 The investigation was limited to studies of the Cuyuna ores listed in Table I, which were reduced to the metallic iron state under various conditions

Jan 1, 1969

By H. E. N. Stone, B. L. Daniell

This paper examines the cracking behavior on chemical reduction of three types of ore/oxide briquette mixtures. The complex sintering/hardening atmosphere was replaced by 100% oxygen and the blast furnace atmosphere by a 5/95 hydrogedsteam mixture which was used isothermally at 700°C for a constant, arbitrary time. Cracking has been assessed by measurement of reduction expansion or increase in specimen diameter as a result of chemical reduction and a subjective index based on appearance and friability. Oxidized briquettes of iron ore concentrate crack on reduction'-3 in a manner which parallels the degradation of sinter in the Linder test. In the present work, the cracking behavior on chemical reduction of three types of ore/oxide briquette mixture has been studied. Firstly, LD converter fume with additions of silica; secondly, Oxfordshire ore with ferric oxide; thirdly, ferric oxide with separate additions of silica, alumina and lime. Motivation for the first-mentioned experiments lay in the improvement of the fume as furnace feed; for the second, in assessing the tolerance of a finely divided ferric oxide as a base for low melting point additions. The third series provided additional information. In this work, the complex sintering/hardening atmosphere is replaced by 100% oxygen and the blast furnace atmosphere is replaced by a 5/95 hydrogen/ steam mixture used isothermally at 700°C for a constant arbitrary time. The particular reduction conditions were known from previous experience3 to be capable of disclosing any tendency to cracking. Cracking has been assessed by measurement of reduction expansion (i.e., increases in specimen diameter as a result of chemical reduction) and a subjective index based on appearance and friability. Generally

Jan 1, 1970

By S. M. M. Safvi, A. Jowett

Small-scale continuous flotation tests are described in which the influence of 1) pulp density and 2) feed rate on the rate of flotation are investigated. The results provide further evidence of first-order kinetics for flotation and indicate that tests at different feed rates provide a particularly simple method of determining rate constants. Flotation kinetics are discussed in general terms. Over a period of many years numerous research workers have studied the rate of recovery of valuable materials from flotation pulps with a view to deter-mining the mechanism of the froth flotation process. Froth flo ation is fundamentally a collision process involviug at some stage collision and then mutual attachment between air bubbles and mineral particles. It is possible to draw analogies therefore between the flotation process involving mainly macroscopic particles and the ordinary processes of chemical reaction kinetics involving collisions between atoms, molecules and ions. By adopting this analogy we may postulate a general equation representing the kinetics of the flotation process: -dC/dt =K.CC ,Aa (1) The rate constant, K, will be a complex function involving reagent concentrations, particle and bubble sizes, induction times and no doubt flotation cell design. However, these are factors which can be kept constant in any series of experiments, permitting investigation of the exponents a and c, which determine the order of the reaction occurring in flotation. The various research investigations carried out so far have all aimed at evaluating a and c and in the majority of cases only since the concentration of air has usually been kept constant in a series of tests, so that a relationship of the following type has been investigated: -dC/dt = K1 Cc (2) The techniques of both batch testing and continuous testing have been used to investigate this relationship. The technique of continuous testing, in which steady-state conditions are established, permits direct investigation of the above equation. In the case of batch testing, C is decreasing continuously

Jan 1, 1961

By W. F. Rochow, W. C. Burke

Numerous designs of linings and accessories, including dams, lifters, and heat exchangers, contribute greatly to kiln efficiencies. Greater conductivity is achieved with basic brick than with fireclay and high-alumina refractories, depending largely upon the use of steel plates in radial joints between adjacent brick to increase heat transfer through the lining. Refractory lifters expose all portions of the load to radiant or convective heat and raise some portions to higher temperatures by cascading the load down through the path of the hot gases. AMONG the many factors involved in rotary kiln efficiency, advantageous use of refractories is of major importance. So far as kiln lining alone is concerned durability is most significant, as shutdowns for repairs are costly in fuel consumption and production losses. Aside from selection of refractories best suited to the widely varying conditions under which rotary kilns are operated, lining design and accessory constructions within the kiln are highly important. The refractory most economical for any combination of operating conditions can be selected from a wide range of alumina-silica and basic refractories, including insulating, castable, and plastic fire bricks. In several cases silica brick can be used to advantage. Alumina contents of alumina-silica refractories range from 35 to 90 pct in increments of 5 to 10 pct. In general, bricks of relatively low alumina content are best for abrasion resistance, an important property for the brick used in chain sections of cement kilns or in feed end sections of limestone kilns. High-duty fireclay bricks of approximately 40 pct alumina and 55 pct silica are used to line the intermediate zones. Super-duty fireclay brick at slightly higher cost is warranted in operation of kilns with high intermediate zone temperatures and severe spalling conditions. Alumina-silica refractories with an alumina content of 50 pct and higher, termed high-alumina brick, are available in commercial classes of 50, 60, 70, 80, 85, and 90 pct alumina content. These refractories are generally used as high temperature zone linings or placed adjacent to the basic brick-lined sections when the treatment they are required to withstand is excessively severe. Use of basic refractories is confined mainly to zones of maximum temperature and most severe chemical action. Basic refractories ordinarily employed for high temperature zone linings of many rotary kilns comprise the classes shown in Table I. Each class is peculiarly well suited to various particular combinations of working conditions. The approximate chemical analyses of these classes also illustrate differences in compositions with respect to the major components.

Jan 1, 1956

By R. L. Stone

None of the currently used methods of testing bentonites as binders for strengthening iron ore pellets is entirely adequate by itself. Since more knowledge was needed about 1) which properties of bentonite most affect pellet strength, and 2) the precise mechanism by which strengthening is produced, attention was paid to possible new techniques for rating bentonites. With the structural nature of bentonites — in particular, the montmorillonite component — the possibility of rating bentonites on the basis of zeta potential was investigated. These bentonites were also rated by conventional tests and measurements, including colloid content, barrel yield, pH, chemical analysis and physical tests on green and dry pellets. A mechanism for pellet bonding based on electro-kinetic forces between the bentonite and the iron ore concentrate appears to be feasible. In pelletizing, from 6 to 18 lbs of bentonite is added per ton of concentrate to give strength to the green and dry pellets. In some operations even more bentonite than is required for strengthening purposes must be added for moisture adsorption to pennit balling of the concentrate. Since bentonite is thus a major item in pelletizing, more knowledge about the way in which bentonites act as binders and better ways of rating this ability would help in the difficult job of selecting the right bentonite for a given concentrate. The fact of the matter is that although several methods are in current use for testing bentonites, none of these methods are consistently foolproof. For example, a given bentonite may rate high by several standard tests yet turn out to be a poor binder, and of course the reverse also occurs. Actually, more knowledge is needed about 1) which properties of bentonite most affect pellet strength, and 2) the precise mechanism by which strengthening is produced. In this study attention is therefore paid to possible new techniques for rating bentonites. With the structural nature of bentonites as a starting point, we investigated the possibility of rating bentonites on the basis of the zeta potential, i.e., the electrokinetic ootential on the montmorillonite colloids — the main component of bentonite binders. The zeta potential technique was applied to a range of commercial bentonites from several areas in the U.S. and abroad. These bentonites were also rated by conventional tests and measurements, including:

Jan 1, 1968

By David R. Mitchell, Ernest M. Spokes

There is evidence that study of minerals now considered to have susceptibilities too low for magnetic separation should be continued. Present concepts may be false. INFORMATION on magnetic properties of minerals other than magnetite is scanty and sometimes erroneous. Often there is no information at all. W. R. Crane's table of tractive forces published in 1901 is still the best guide in the field, but Crane's data can be used as a guide only, since he ignored such important factors as impurities in the minerals, size, and packing effect. During the 1930's work by the U. S. Bureau of Mines added greatly to knowledge of the behavior of certain minerals in a magnetic field. The USBM bulletin gave for the first time complete magnetic characteristics for magnetite and fragmentary data for a few other minerals. The work reported here covers a study of the magnetic-susceptibility relationships of the wolframite series of minerals and sphalerites as related to their chemical compositions. Theories are proposed to explain divergences found from commonly accepted beliefs. Gouy Method for Determining Magnetic Susceptibility: The Gouy principle has been adequately described in the literature1-3 and applied in various ways.'- It may be summarized by stating that a body of material (gas, liquid, solid, or powder mixed with gas or liquid) of cylindrical shape changes in apparent weight if suspended with one end in a zone of higher magnetic field strength than is found in that portion of the field surrounding the other end of the cylinder. This may be expressed as K2 - K1 A (H2 - H2) - mg. [1] In determining the susceptibility of mineral specimens by the Gouy method, the mineral is reduced to a fine powder and packed in a glass weighing tube. Precise measurement of the bore and depth of the tube is made unnecessary by a calibration procedure. Using the basic expression of Eq. 1, Hoare"

Jan 1, 1959

By A. L. Mular

lnput energy-size modulus relation at small reduc-tion ratios is expressed by Et = (M, -Mt) k;. In some cases 13 is close to Schuhmann's distribution modulus a in value. Experimental results in conjunc-tion with the equation demonstrate the fact that energy input which produced fracture is not expended on unbroken starting material inside the mill at any time. Recently, schuhmann1 derived the Charles equation2 using the postulate that complex com-minution operations can be considered as a summa- tion of many comminution events. The fragments resulting from each event are assumed to follow the Gaudin-Schuhmamn distribution. In his derivation, Schuhmann implies that the size distribution of the total mass of a material is determined by the fraction of that mass actually subjected to comminution. Consider a feed to contain a volume, Vo, of uni- formly sized homogeneous material. After t seconds of grinding, let the volume of broken material be Vo - Vt where Vt is the volume of starting material not yet broken. The size distribution of the broken material is given by: where V is the cumulative volume less than size X, and k, is the size modulus of the broken material at time t. The distribution for the total volume at this

Jan 1, 1962

By A. L. Mular

Two equations express the integral rate at which fines are produced from a homogeneous material: It has been demonstrated that the Gaudin-Meloy distribution for single fracture is equally applicable to multiple fracture and is related to the Gaudin-Schuhmann distribution at least when the distribution modulus a is close to unity. A, xoY was found to be independent of starting mass over the range studied when particle size was held constant. The equations are related to the input energy to the mill. In an earlier paper,' it was shown that when Mo g of uniformly sized homogeneous material is ground dry in a batch rod mill, the energy-size reduction equation is expressed at any time by the mass of unbroken starting material left at time t, ku is the size modulus of the portion Mo - Mt, and P is a constant that is generally equal to the distribution modulus a. If the power to the mill is constant, Et is proportional to time and Eq. 1 is rewritten where (Mo -Mt)/t is actually the integral rate at which fines are produced from a parent material, and A, is another constant. The size modulus, k,, is determined from the Gaudin-Schuhmann equation:2 where y is the cumulative fraction finer than any size x. Recently, Gaudin and Meloy derived a distribution equation for single fracture on the assumption that the random fracture of a homogeneous material produces daughter particles of similar shape. Their equation is given by where y is the cumulative fraction finer than size X, x, is the size of the unfractured parent, and r is designated the size ratio. Although Eq. 4 is derived for a single event, it should also approximately hold for a series of com-

Jan 1, 1962

By Roshan B. Bhappu, Raul A. Deju

A method is proposed by which the chemical reactions between water and a silicate bed can be related to equations of flow. First the reaction mechanism is studied both theoretically and experimentally. Once the reaction mechanism is established, the equations of flow are solved, and a solution (under physically reasonable side conditions) is obtained. Experiments to substantiate the solution are also presented. The knowledge of surface reactions at the water-solid interface is of critical importance in extractive metallurgy. This is especially true in regard to the froth-flotation process of mineral concentration and all phases of hydro metallurgical recovery of metals. Similarly, the problem of ground-water contamination is one of growing concern, not only because of the increasing demands on our sources of supply but also because of the increasing necessity to find vehicles of waste disposal. The hydrochemistry of waste waters must be understood to exploit and develop such resources properly. To understand these problems, it is necessary to obtain a better insight into the reactions that take place between water and a solid packed bed. To do this, we will develop here in a theoretical model that predicts the fate of a chemical species in solution as it reacts with a silicate bed through which the liquid percolates. The treatment will be one-dimensional in nature and will not include dispersion. Experiments will also be presented to substantiate the theoretical model, using the silicate minerals listed in Table 1. This paper is preliminary in nature and is part of a larger study of the surface chemistry of the silicate-water interface. Mechanism of Surface Reactions In a previous paper,' the authors discussed the properties of various silicate minerals from a mineralogical standpoint. Now efforts will be made to study the reactions that occur when these minerals are put in contact with water. The reactions occur almost entirely in the surface layer of the mineralsa and are termed sorp-tion reactions. In this study sorption denotes the composite of all physical adsorption and chemical adsorption occurring between the silicate surface and the water in contact with it. A model of sorption as seen from a chemical-bond standpoint is presented below. On the basis of this model, the surface reactions between deionized water and a silicate mineral can be studied. Sorption Model: It is known that as the oxygen-silicon ratio increases from quartz to the olivines, a greater percentage of the oxygen bonding power is available for bonding to cations other than silicon. Hence, with an increasing oxygen-silicon ratio, the oxygen-to-metal bonding increases. Upon the fracturing of a silicate-mineral crystal, the oxygen-metal bond breaks more easily than the stronger oxygen-silicon bond, resulting in a negatively charged surface. Then, if the mineral is immersed in a liquid containing hydrogen ions, the negatively charged surface should tend to be neutralized by hydrogen ions from solution, resulting in a change of the solution pH. Thus, an increase in the degree of sorption of hydrogen ions must occur as the oxygen-silicon ratio in the crystal structure increases. Also in a previous paper,' the authors found that for a given silicate mineral the change of pH with time is given by

Jan 1, 1971