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By S. Korman, S. B. Tuwiner

THE purpose of this study is to investigate the effect of agitation upon the flotation of chal-cocite. It is believed that the concentration of collector which is required to produce a water-repellent surface upon the mineral is determined among other things by the rate of collector diffusion through an impeding film of more or less stagnant fluid. Its thickness depends upon the intensity of agitation. In the first part of this study a method is developed for determining the thickness of the fluid film upon a polished surface of synthetic chalcocite. The film thickness is then determined as precisely as possible under three readily reproducible condi-tions of agitation intensity. These determinations serve as a quantitative evaluation of agitation. In Part 2 it is demonstrated that in solutions of lime at PH 10.5 and in the absence of air or dissolved oxygen the concentration of collector required to develop water-repellency upon the surface in any given time is a function of the fluid film thickness, and that after about 10 min the relation becomes linear. This points to the possibility that the collector ions may be involved in a diffusional process through the fluid film. In Part 3 it is shown that where air is present more collector is required to produce water-repellency than for the case in which air is excluded. Next, it is shown that for the latter case the most minute amount of sodium sulphide suffices to preclude development of water-repellency. Both oxygen and sulphide are in a sense, therefore, depres- sants for chalcocite in a lime solution at PH 10.5. When oxygen and sulphide are present simultaneously there is a region of minimum critical collector concentration for water-repellency which is very markedly lower than the concentration where oxygen alone or where sulphide alone is present. Moreover, this minimum appears to be identical with the collector concentration which is required where sulphide and oxygen are both excluded. This would appear to indicate that in certain ranges of concentration of oxygen and sulphide, each cancels the effect of the other. The range of concentration of solution for this effect in solutions saturated with oxygen is a function of the agitation intensity. Thus we are led to believe that a chemical reaction involving sulphide and oxygen occurs at the surface of the chalcocite. Part 4 of this study is a further development of this idea and some of its consequences from an electrochemical standpoint, while in Part 5 the disappearance of sulphide from pulps of several minerals is noted with an indication that both dissolved oxygen and the mineral are necessary, the reaction being catalytic or pseudocatalytic. Part 1. Determination of Fluid Film Thickness The Fluid Film Concept: The hindrance to the diffusion of collector to the surface of the chalcocite is a barrier of stagnant fluid in the vicinity of the surface. It is the same barrier that limits the rate of transfer of heat, the rate of solution of dissolving crystals, or any of the other rates pertaining to chemical engineering unit operations in which the film resistance to diffusion plays a principal role.' Fluid films are present upon the surface to an extent which varies with the geometry and intensity of agitation. They result from the fact that in the region within the fluid near a surface, motion is entirely parallel to it. From the standpoint of transfer to or away from the interface the effect is as though the fluid were stagnant. Outside of the film, movement of the fluid may be streamline or turbu-

Jan 1, 1951

By M. E. Volin, D. W. Fuerstenau, A. L. Mular

This paper presents the results of an investigation of the effect of feed size on the integral rate at which feed material is dry ground in a laboratory rod mill. The data are interpreted in terms of the Gaudin-Meloy distribution equation. The equation, Mo — Mt/t = Aoxo r-1, is found to hold for feed sizes smaller than some critical size or for grinding times larger than some critical time. It is shown that a feed grindability is related to this critical size, which appears to be the transition between predominately impact and predominately attrition grinding. Detailed research on feed size effects in comminution has been rather limited. Gaudin1 ground single-sieve fractions of quartz for extended time periods (four hours) in a laboratory ball mill. His work, which was designed to study the nature of size distributions, showed that there is a critical feed size above which comminution occurs primarily by attrition. For his system, the critical size was about 2-mm when using 22-mm balls. In a more recent investigation2 single-sieve fractions of quartz were ground in a laboratory rod mill for constant time periods (15 min) to determine how feed size affects the grindability of the material. This work showed that grindability is essentially independent of feed size in the rod mill after this moderately long grinding period. To further clarify this observation, a study in which feed size and grinding times are varied over a fairly wide range, and particularly, a study of the dependence of the integral rate of grinding on feed size during the initial stages of rod milling is necessary. For an analysis of such data, it is desirable to use a parameter which is very sensitive to change during the initial grinding period when reduction ratios are

Jan 1, 1964

By A. L. Mular

The flotability of crushed zinc oxide pellets which were doped to produce more n-type or less n-type (more p-type) properties was studied with a Halli-mond tube. Flotation data are presented to show that the pH at which 50% recovery occurs and the collector concentration for 50% recovery depends upon the concentration of doping agent added substitu-tionally for zinc in zinc oxide. The time of sintering of pellets has a slight effect on flotation behavior. A speculative model is proposed to explain the observed behavior in light of the fact that zinc oxide is a semiconductor and has a space charge region at the surface. Experimental verification is necessary before the proposed model can be established as correct. In recent years, the attachment of flotation reagents to minerals has been shown to depend upon the behavior of an electrical double layer which exists in the liquid near the mineral-liquid interface.' Thus, a mineral placed in water will undergo a surface re-arrangement, most likely be preferential dissociation of lattice atoms or by surface adsorp-tion-desorption reactions, to form a dissociation electrical double layer with the solution. Among the many factors which influence adsorption processes, the past history of the mineral itself is one of the most important. Because of trace impurities, minerals from different localities may exhibit marked differences in flotation behavior. This is not unexpected, since investigators who have studied the detailed structure of the double layer's3 have known for some time that the impurity content of many minerals may exert strong influence on double layer variables. For this reason, their studies were usually made with minerals of high purity. The precise relationship between impurities in a mineral and double layer parameters such as capacity and electrostatic potential has been obscure. However, impurities in a semiconductor type mineral which is in contact with a solution have been recently shown to produce changes in the properties of a space charge layer which extends into the bulk of the semiconductor to a distance of one micron or more.4 The double layer is influenced by the space charge layer, whose origin is best discussed in terms of the band theory of solids. A qualitative description of the behavior of impurities in the bulk of a semiconductor, from a chemical point of view, is useful at this point. Briefly, impurities or defects (lattice vacancies, interstitials, etc.) in semiconductors such as PbS, ZnS, FeS2, Cu2S, NiS, NiO, Fe2O3, ZnO, CdS, MnO2 and many others, behave somewhat like acids or bases in water. For example, water dissociates into hydrogen and hydroxyl ions according to

Jan 1, 1965

By S. Usui, I. Iwasaki

The effect of pH on the adsorption of dodecylamine at the mercury-aqueous solution interface was investigated by differential capacity and electrocapillary measurements. With dodecylammonium acetate, the differential capacity curves showed two desorption peaks in the cathodic branch with their relative intensities varying with the solution pH. With dodecyltrimethylammonium chloride, only one cathodic de-sorption peak was observed in the same pH range. Through thermodynamic analysis of the electrocapillary curves, the adsorption density of undissociated amine was evaluated separately from that of aminium ion. The adsorption densities of the un-dissociated amine and of the total amine increased with increasing pH. The ratio at the interface of undissociated amine to aminium ion was several orders of magnitude greater than the ratio in the solution and increased with increasing pH. The potential at the closest distance of approach of counter ions to the mercury surface was compared with values of zeta potential on quartz previously reported. The most important variable in the flotation separation of minerals is probably the pH of the pulp, and a number of theories have been proposed to explain its effect on the condition of the mineral surfaces, on the dissociation of collectors and of inorganic and organic species (accidentally present or intentionally added) in the pulp, and on the mineral-collector interaction. In the development of a theoretical background for oxide flotation systems, an experimental approach based on electrokinetic measurements has been of much value, although the effect of pH becomes confounded since it governs both the electrochemical conditions of the oxide surface and the dissociation of the collector. For investigation of the adsorption behavior of long-chain collectors on oxide minerals, however, electrokinetic potential measurements are the most widely used technique. Hydrogen and hydroxyl ions are found to be the potential determining species, thereby governing the interfacial electrical conditions. The electrostatic interaction between the charged mineral surfaces and ionized collectors is regarded as the driving force for the adsorption of the collectors. An association of alkylamine collectors adsorbed on quartz surfaces has been postulated from streaming potential measurements, and a term "hemi-micelles" has been proposed.' The possibilities of coad-sorption of undissociated amine along with aminium ion has been inferred from contact angle measurements? and from adsorption studies.~ Electrochemical titration as applied to silver sulfide provides a more quantitative approach to the analysis of the electrical double layer at an ionic solid-solution interfaceqG and the electrochemical evidence for the adsorption of amine at pH 4.7 indicates a specific affinity of dodecylammonium ion towards silver sulfide surfaces, whereas at pH 9.2 the adsorbed species might be free arnine." A combination of differential capacity and electrocapillary measurements on a dropping mercury electrode was reported to be a sensitive method of provid- ing reliable information on the adsorption behavior of dodecylammonium acetate (DAA) at a natural (near neutral) pH.? It was also shown that there were striking similarities in the properties of the double layer and in the adsorption behavior of the amine on mercury and on such ionic solids as quartz, silver sulfide, and silver iodide. The effect of pH on the differential capacity curves at a mercury-sodium fluoride solution interface has been investigated by Austin and Parsonss who reported that between pH 7 and pH 11 there was very little effect. In the present paper, the adsorption behavior of DAA was investigated as a function of pH through differential capacity and electrocapillary measurements and the information gathered was correlated with that available in literature on quartz and silver sulfide. Experimental The apparatus and the method used for determining the differential capacity and the electrocapillary curves were identical to those described previously.' The ionic strength of the supporting electrolyte was fixed at 0.1 M with potassium fluoride, and the pH of the solution with potassium hydroxide. Only the neutral to alkaline range was covered in order to avoid the dissolution of the glass vessel with hydrofluoric acid. Results In Fig. 1 the differential capacity has been plotted against the applied potential at a DAA concentration of 10-' M at three different pH values. The curves are characterized by one capacity peak in the anodic branch, by two capacity peaks in the cathodic branch, and by a marked depression in capacity between the peaks. The depression indicates an adsorption of the arnine in this potential range. One of the cathodic peaks appears at pH 7.3 near -1.4 v and decreases with increasing pH. The other appears at pH 8.9 near —1.2 v and increases with increasing pH. At pH 9.6 only the latter peak is observed. Beyond the cathodic peaks, all the curves tend to converge with the curve in the absence of DAA, implying that two different species are being desorbed in this potential region. The anodic peak near 0.0 v increases markedly with increasing pH. The well-defined anodic peaks at pH 8.9 and 9.6 were accompanied by an appreciable increase in the current flow (in excess of O.luA), and, therefore, is a "pseudo-capacity"'" due to a

Jan 1, 1971

By S. R. Mitchell, M. C. Fuerstenau, A. M. Gaudin

The two principal attributes of a thickener pulp are its settling rate and the ultimate pulp density of the thickened mud. Testing for evaluation of thickening attributes of a pulp has usually been done batch-wise by placing the pulp in a one-liter glass graduated cylinder and noting the position of the boundary between clear liquor and mud as a function of time. Refinements in testing have included repulping a thickened mud to various dilutions and evaluating the settling rate on these various repulped muds. This is considered to make possible an estimate of the settling rate as a function of pulp density and therefore to pinpoint the slowest settling pulp which ought to be critical from the standpoint of thickener plant design. Finally, tests are sometimes made with model-size continuous thickeners of arbitrary size. In batch tests it is generally accepted that the behavior of a pulp is described by a curve such as the one shown in Fig. 1, in which the position of the boundary between clear liquor and mud (the mudline) is plotted against time. It has also been assumed that the initial settling rate (the slope of the line segment AB) is descriptive of the pulp and independent of the size and shape of the container. Where the settling rate is slower and clearly a function of time (portion CD of the settling curve), it has been assumed that the data are not of much use because of the intrusion of compacting phenomena. The next important datum is the height. PE, which is a measure of the so-called ultimate pulp density. This quantity is assumed to be characteristic of the pulp and independent of the container. When actual data are recorded, results are rarely exactly according to this theory and numerous deviations may be noted. For example, some pulps seem to require an appreciable delay before beginning to settle, as in line AF. This is spoken of as representing the need for an induction time for flocculation. It has seemed desirable to explore thickener data with a more sensitive charting tool than is afforded by the coordinates of Fig. 1. Instead of relating the position of the mudline to time, it has seemed desirable to relate the rate of descent of the mudline to time. This, in effect, amounts to charting the slope of the line in Fig. 1, or the derivative of its ordinates, against the same abscissas. Fig. 2 shows the new chart for the curve ABCDE of Fig. 1. When this new tool was applied to data obtained, for example, with sized clay pulps, astonishing differences appeared. Apparently the traditional tool conceals differences instead of revealing them, whereas the new method emphasizes them. It seems, also, that both charting methods have their uses. In this article attention is focused on the effects of pulp density and container height as they affect the settling behavior of flocculated suspensions of a sized kaolin. The thesis by M. C. Fuerstenau' gives more details than this summary, as well as some information about the effect of other variables. Sized fractions of kaolin were used, supplied through the courtesy of S. C. Lyons of the Georgia Kaolin Co. Large batches were made by agitation in a flotation machine and in a glass jar equipped with a mixer. The settling vessels were 2, 4, and 6-in. diam glass cylinders of various heights. Pulp Dilution: The effects of pulp dilution on sedimentation behavior are shown in Figs. 3, 4, and 5. Fig. 3 shows data on pulps ranging from 2.44 to 16.7 pct solids initial pulp density, flocculated with Separan and charted in the traditional manner. At least two of the lines appear nearly straight in their early sections, but neither of the others does. Each of the curves in this figure includes a region labeled

Jan 1, 1960

By F. A. Forward, A. H. Ross, J. Halpern

IN treating carnotite ores by carbonate leaching it is often necessary to subject the ore to a prior roast. Among the advantages that may result from roasting are: 1) improvement in settling and filtering characteristics of the ore; 2) destruction of carbonaceous impurities that would otherwise contaminate the leach solution and consume carbonate reagent; and 3) substantial increase in extraction of vanadium, which is usually low when the un-roasted ore is leached. Recovery of vanadium may be an important factor in processing carnotite ores, and the fact that vanadium is usually recovered in good yield by acid leaching has often been an economic advantage over results obtained with carbonate leaching. It has been found that response of an ore to carbonate leaching, following roasting, depends primarily on its composition and on the conditions of the prior roast treatment, particularly with regard to the roasting temperature, the atmosphere maintained above the charge during roasting and cooling, and the presence of certain reagents. These factors, especially the last, virtually determine the extractions of uranium and vanadium attainable by carbonate leaching, as well as the reagent consumption on leaching, whereas adjustment of the leaching conditions has only a limited influence. These variables were examined systematically to elucidate roasting and carbonate leaching of carnotite ores and to establish a treatment yielding high recoveries of both uranium and vanadium. Tests were conducted on samples of four different Colorado Plateau ores, described in Table I. Roasting Procedure: The roasting tests were conducted in an electrically heated furnace in which the temperature was controlled to +15oC. The ore (usually a 500-g charge) was ground to —100 mesh, mixed with the desired reagent, and spread in a silica tray to a bed depth of about 1/2 in. During the roast the charge was rabbled at intervals of 15 min. To provide free access of air the furnace was usually ventilated; however, in a few experiments a controlled atmosphere such as carbon dioxide or hydrogen was maintained above the charge. Leaching Procedure: The pulp, generally comprising 500 g of ore and a 2000-cc solution containing 80 g per liter of Na2CO3 and 20 g per liter of NaHCO3, was agitated in an autoclave at 120°C, under 30 psi O2. Samples were withdrawn periodically, and the solution and ore residue were analyzed. Under the conditions employed, maximum attainable extractions were generally achieved in 4

Jan 1, 1958

By W. A. Hockings, M. E. Volin, A. L. Mular

Batch grinding tests at short times were made in a laboratory rod mill with 10 x 14 mesh quartzite in corn syrup-water mixtures of varying viscosity. The weight fraction broken and size modulus were found to be independent of viscosity up to 20 cp, but at higher viscosities breakage of the feed decreased while the size modulus increased. The mechanisms of the effect of viscosity are discussed on the basis of particle dynamics. The effect of viscosity of the suspending medium on grinding has not been accorded much attention, although both wet and dry grinding are used in bene-ficiation processes and the viscosities of water and air differ greatly. schweyerl investigated the rates of grinding of quartzite in a pebble mill with air, water and glycerol as the suspending media and found the rate in terms of new surface developed per thousand revolutions of the mill to be constant and dependent on viscosity up to 20,000 rev, but to decrease and become independent of viscosity for longer grinding times. Based on considerations of the drag forces exerted on suspended particles and on the grinding media under turbulent and laminar flow conditions, viscosity can be expected to inhibit the rate of grinding in a given system as it changes the flow conditions from turbulent to laminar. The objective of this study was to determine the effect of viscosity of the suspending fluid in the batch grinding of a homogeneous feed of uniform size for short residence times, a region of more practical interest than the long times investigated by Schweyer. In actual grinding operations, the viscosity of the fluid does not change, but the consistency (apparent viscosity) of the pulp increases with additions of fines. In this study the consistency of the pulp was not measured because it was thought that the grinding times were too short to alter the apparent viscosity appreciably.* *Minus 200 mesh quartz at 65% solids has an apparent viscosity of about 25 cp as measured by a simple consistometer. Personal communication from D. F. Kelsall to A. L. Mular. For the purposes of this study, the effect of viscosity on grinding can be shown sufficiently well on the basis of the cumulative weight fraction finer than the feed size at any time, and the size modulus, ku, of the finer than feed sizes. The parameter, ku, is obtained from a form of the Gaudin-Schuhmann equation. METHOD AND MATERIAL The grinding tests were accomplished in a laboratory rod mill 10-1/2 in. long and 8 in. bore with a 20-lb rod charge consisting of two 1-in., eight 3/4-in., eight 1/2-in., and eight 1/4-in. diam rods. The mill speed was 43.2 rpm. Charges of 400 g 10 x 14 mesh Wisconsin quartzite** were ground for times of 60, **Courtesy of Minnesota Mining and Mfg. Co. 120, 180 and 240 sec in 500 cc of fluid made up of corn syrup and water in measured proportions. The temperature of the pulp was measured immediately upon completion of grinding, and the size distribution of the quartzite was determined by a wash-wet-dry screening technique. Corn syrup was chosen because of its high viscosity and ideal viscous behavior, and also because its density is not greatly different from that of water. The viscosity of each fluid mixture was measured with an Ostwald viscometer at 25 C. and a correction was made for the temperature of the pulp according to tabulated values for sucrose solutions. The density differences were so small that they probably did not significantly affect breakage. RESULTS AND DISCUSSION The conditions of the grinding tests, computed viscosities, weight fractions finer than the feed size and values of a and ku are listed in Table I. The values of a and ku were taken from the linear portions of the log-log plots of cumulative weight fraction finer vs size for the various grinding times. The plots for tests in fluids of average viscosity 8.6 cp are shown in Fig. 1. In general all of the sets of curves were linear throughout most of the size

Jan 1, 1965

By T. D. Teimann, R. L. Stone

The rate of extraction of silica from certain Wisconsin and Minnesota taconites was investigated by autoclave digestion in sodium hydroxide solutions over the temperature range from 296° to 416°F. Rates obtained were from 7 to 254 g mol per cm2 per sec giving activation energies from 15,000 to 19.000 cal per mol. Activation energy was found to increase with increase of sodium hydroxide concentration. Digestion of different forms of quartz indicated the rate of extraction to be solely a function of surface area under similar conditions of digestion. The results of an experimental investigation on the caustic extraction of silica from certain Wisconsin and Minnesota taconites were presented in a recent publication.' It was shown that residual concentrates containing as much as 65% iron can be produced by this method. The occurrence of silica as a micro-crystalline variety of quartz with high specific surface area gave correspondingly high mean rates of extraction in solutions of sodium hydroxide. The present paper extends this previous work by presenting the results of a study of the specific rates of extraction of silica from these ores as related to temperature, sodium hydroxide concentration and variety of quartz. Calculated activation energies for the dissolution of silica are presented and some possible mechanisms are discussed. TACONITES The low-grade siliceous iron ores investigated included taconite from the Wisconsin Gogebic range and a semi-taconite2 from the Minnesota Mesabi range. These are adequately described in the former paper1 and by the U. S. Bureau of Mines. The principal iron minerals are hematite (Fe2O3) and goethite (Fe2O3 . H2O), while the gangue is essentially a microcrystalline variety of quartz. Minor amounts of other minerals are present. Typical analyses show the Wisconsin taconite to contain approximately 28% Fe and 54% SiO2 and the Minnesota semi-taconite 39% Fe and 39% SiO2. Analyses of the actual samples used are given in Table I. The taconite samples were prepared for digestion by reducing the original ore to minus 200 mesh in a Braun pulverizer. QUARTZ Three varieties of quartz were studied in the investigation: Ottawa sand, chert and chalcedony. The Ottawa sand was obtained from the Ottawa Silica Co. The chert and chalcedony came from Wards Natural Science Establishment. A minus 200 plus 400 mesh fraction of the Ottawa sand, used as a surface area standard, was repeatedly washed in hot concentrated hydrochloric acid followed by distilled water before autoclaving. The other two varieties were used directly after grinding through 200 mesh in a laboratory hammer mill. Analyses of these samples appear in Table 11. Microscopic examination of the silica samples by the oil immersion method showed that the Ottawa sand consisted of clean, uniform spheres of quartz with little deviation in size from a mean diameter of 82µ determined by a distribution count. The other two varieties consisted almost entirely of microcrystalline aggregates of quartz crystals; the aggregates ranged from 10 to 70µ. The individual crystals in the chert aggregates were estimated to be of the order of

Jan 1, 1964

By H. S. Choi, I. Iwasaki

The effect of temperature as a parameter in ore flotation has not been systematically studied, although for some ores it has been known for many years that selectivity and grade of concentrates can be improved by conditioning or flotation at moderately or substantially elevated temperatures. In 1934 Coghill and Clemmer' stated that "elevated temperature seems to make hard water more tolerable" in soap flotation, and a cursory examination of the literature indicates that some earlier investigators were aware of beneficial results obtained by floating above room or mill-water temperatures. Tartaron,2 Mitchell et aL.,3 and particularly Hamilton et al.' have shown that high pulp temperatures give high recovery and selectivity in soap flotation of fluorite from a variety of gangue minerals. As a result of the work of these and of other investigators, high-temperature flotation has been successfully applied to industrial flotation of fluorite. Falconer" stated: "The use of high temperature conditioning (above 35°C and preferably above 60°C) in connection with the soap flotation of nonsulphide ores, as covered in B. Kalinowski's French Patent No. 847,-215, Dec. 7, 1938, is claimed to provide more intense activation, better separation of values from gangue, and reduction in quantity of reagents." The various investigations quoted above clearly demonstrate the improvements to be realized by increasing the temperature at which fluorite is floated, and it is surprising that little or no information has been published on systematic investigation of the same method as applied to other ores. The present article describes comparative results obtained at room temperature and at elevated temperatures by otherwise conventional flotation of iron ores, by single-mineral flotation in the Hallimond tube, and by contact-angle measurements. It is not the purpose of this article to evaluate the commercial feasibility of high-temperature flotation of iron ores, although preliminary calculations show that the cost should not be excessive and that the gains to be realized may well outweigh the extra cost of heating the pulp. LABORATORY FLOTATION TESTS Experimental: All flotation tests reported in this article that use a laboratory Fagergren cell were made on ore A (either containing the primary slimes or deslimed at 20µ as specified) following the preparatory procedures described elsewhere- or this particular ore. Ore A contained hematite, goethite (limonite), quartz, chert, and minor quantities of magnetite. Confirmatory tests made on other ores show that the same general principles apply to ores of similar mineralogy and structure. Conditioning was performed by adding the collector, as a soap, to the vigorously stirred and heated pulp. Conditioning time was 5 min, the pulp contained 60 pct solids, and the equivalent of 0.5 Ib of fatty acid was added per ton of original (unde-slimed) ore. The pH of the pulps ranged from 6.8 to 7.1 (25°C measurement) during conditioning. The pH was adjusted in the flotation cell by addition of either NaOH or H2SO4, the values reported being those of the pulp cooled to 25°C. During conditioning and flotation the pulp temperatures were controlled within l° to 2° by immersion of Variac-controlled Calrod elements in the conditioning beaker and in the cell, and by the use of water, heated to the appropriate temperature, for transferring pulp, adjusting pulp volume, and washing. Selectivity indexes have been calculated on the same basis as those reported in the article by Cooke, Iwasaki, and Choi." Results: Table I gives the results obtained by floating deslimed ore A with pure oleic acid at 25" and 50°C, and Fig. 1 shows the selectivity indexes plotted against pH of flotation. Inspection of Table I

Jan 1, 1961

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of cassiterite and fluorite with oleic acid as collector and with alizarin dyes as modifying agents were studied by means of small-scale, vacuum-flotation tests. Artificial cassiterite was prepared fot these tests by high-temperature methods, especially by crystallization from borax melts. FLOTATION test work on Bolivian tin ores some years ago led to the development of several workable schemes of selective soap flotation, but left unanswered many questions of soap-flotation chemistry.' Accordingly, the study of the activation and soap flotation of quartz just reported? was undertaken as a start toward the elucidation of some of the unanswered questions. The experimental work on quartz flotation consisted primarily of small-scale, vacuum-flotation tests in which activator and collector additions and pH were varied between wide limits. In the work to be described in this paper, the soap flotation of cassiterite and fluorite was studied by the vacuum-flotation technique developed in the previous study of quartz. Using oleic acid as collector, conditions for floating cassiterite and fluorite in the absence of modifying agents were first determined. Then the effects of barium, calcium, and alizarin dyes on the flotation of cassiterite, fluorite, and quartz were studied. Alizarin dyes were chosen for study because of the promise they have shown as selective modifying agents in the flotation of non-sulphide minerals.12,1 Also, they form a family of reagents in which there should be a systematic correlation of chemical properties with composition. As a preliminary to the study of cassiterite flotation, an investigation of the preparation of synthetic cassiterite and X ray studies of natural and synthetic cassiterite were undertaken. Experimental Methods The flotation technique and most of the reagent preparations were described fully in the previous paper.' In addition, for the present work it was necessary to prepare satisfactory stocks of clean cassiterite and fluorite and to obtain samples of alizarin dyes. Cassiterite: High-grade cassiterite from the Netherlands East Indies was crushed, sized, and cleaned

Jan 1, 1951

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of cassiterite and fluorite with oleic acid as collector and with alizarin dyes as modifying agents were studied by means of small-scale, vacuum-flotation tests. Artificial cassiterite was prepared fot these tests by high-temperature methods, especially by crystallization from borax melts. FLOTATION test work on Bolivian tin ores some years ago led to the development of several workable schemes of selective soap flotation, but left unanswered many questions of soap-flotation chemistry.' Accordingly, the study of the activation and soap flotation of quartz just reported? was undertaken as a start toward the elucidation of some of the unanswered questions. The experimental work on quartz flotation consisted primarily of small-scale, vacuum-flotation tests in which activator and collector additions and pH were varied between wide limits. In the work to be described in this paper, the soap flotation of cassiterite and fluorite was studied by the vacuum-flotation technique developed in the previous study of quartz. Using oleic acid as collector, conditions for floating cassiterite and fluorite in the absence of modifying agents were first determined. Then the effects of barium, calcium, and alizarin dyes on the flotation of cassiterite, fluorite, and quartz were studied. Alizarin dyes were chosen for study because of the promise they have shown as selective modifying agents in the flotation of non-sulphide minerals.12,1 Also, they form a family of reagents in which there should be a systematic correlation of chemical properties with composition. As a preliminary to the study of cassiterite flotation, an investigation of the preparation of synthetic cassiterite and X ray studies of natural and synthetic cassiterite were undertaken. Experimental Methods The flotation technique and most of the reagent preparations were described fully in the previous paper.' In addition, for the present work it was necessary to prepare satisfactory stocks of clean cassiterite and fluorite and to obtain samples of alizarin dyes. Cassiterite: High-grade cassiterite from the Netherlands East Indies was crushed, sized, and cleaned

Jan 1, 1951

By Brahm Prakash, R. Schuhmann

Maurice Rey—It may be interesting to note that depressing effects can also be obtained from cyclic compounds other than dyes. One such compound which is a dispersing agent for carbon, pigments and other compounds is known in France by the trade name "dispergine" and in the United States by the name of Lomar P W (Jacques Wolf & Co.) Dr. Rey admits that the compound is not very selective and is similar in its action to starch and dextrin.

Jan 1, 1951

By Brahm Prakash, R. Schuhmann

Maurice Rey—It may be interesting to note that depressing effects can also be obtained from cyclic compounds other than dyes. One such compound which is a dispersing agent for carbon, pigments and other compounds is known in France by the trade name "dispergine" and in the United States by the name of Lomar P W (Jacques Wolf & Co.) Dr. Rey admits that the compound is not very selective and is similar in its action to starch and dextrin.

Jan 1, 1951

By J. J. Strohl, H. J. Schwellenbach

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By H. J. Schwellenbach, J. J. Strohl

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By J. F. Myers, F. M. Lewis

The purpose of the mill tests reported herein, was to determine the relative power efficiency of fast and slow rod mill speeds on the ores of the Tennessee Copper Co. The tests were carried out at the Isabella Mill of the Tennessee Copper Co. under normal operating conditions. Every effort was made to keep the ore supply and other conditions constant, but it will be noted that some slight variation did occur. The open circuit 6 by 9 rod mill, and ore, used in these tests have been previously described.' Four motor drive pinions were available and the tests were performed in the following manner: each pinion was used to drive the mill for a period of one week. Constant sampling prevailed during the test run, and a composite sample for each speed test was made. Realizing that some slight liner wear took place during the four-week test-run, and that some slight variation of the ore may have taken place, a second and then a third series of four-week test-runs were made with the four different pinions. The screen analysis of each of the three-week run samples at the various speeds were composited and the results are shown in Table 2. The authors feel that the 22 weeks of operation consumed during these test-runs would represent fairly the effect of each mil1 speed. The data of line 37 show that within the error of sampling and plant operation, the tons of minus 65-mesh material generated per horsepower hour was the same at all four speeds. Obviously, rod and liner consumption could not be determined for each mill speed in so short a time. The rod load was maintained at a uniform level throughout the series of tests. The authors have operated rod mills at various speeds over the past several years, and it is their observation that rod consumption is directly proportional to the mill speed. Likewise, that the rod consumption per unit of work performed is a constant, providing that all other operating factors remain the same. It is a well known fact that the efficiency increases in a ball mill as the speed decreases. This is due to lateral segregation of the large and

Jan 1, 1950

By J. F. Myers, F. M. Lewis

C. G. McLACHLAN*-I have read this paper with considerable interest and wish to congratulate the authors on the care with which they carried out their experiments and for the detailed sizing data they have presented. 011 the other hand I do not feel that these data establish their contention that "with other factors the same, the work accomplished in a fine crushing rod mill is directly proportional to the rotating speed." My reason for making this statement is that their conclusion is based on the percentage of minus 65 mesh material produced at each of the speeds at which their tests were run. The rod mill is, however, not a fine grinding machine and therefore figures based on minus 65 mesh production are figures outside the range at which it operates to best advantage. Further, if rod mill performance is to be judged on the production of such material it should have been established that a rod mill will produce it as effectively as a ball mill. To check this last point I have gone over some of the rod and ball mill grind-ing figures for our Waite-Atnulet operation and find that the ball mill at that property produces more minus 65 mesh material per horsepower than the rod mill. Moreover, this comparison is even more favorable to the ball mill if it is limited to minus 100 mesh production. It therefore seems to me that the most satisfactory basis on which to try to com-pare the performance of the rod mill in the present case is by taking a weighted average of all the new surface produced per horsepower. W. H. Coghill3 has published data which can be applied to the sizing figures presented by the authors, and if this is done the units of new surface per horsepower produced in the Tennessee tests are as follows: Admittedly the increment of increase in new surface per horsepower shown in the foregoing figures is not great, but its trend is nevertheless definitely in favor of slower speed operation, for the conditions under which these particular tests were run. I also feel that some information regarding liner contour should have been given as I think it will be found that a mill which is equipped with liners which lift the rod load can be run at a slower speed than a mill in which the liners are comparatively smooth. J. F. MYERS and F. M. LEWIS (authors' reply)—The position taken by Mr. McLachlan is to be respected even though we are all aware of the faults of surface considerations. As he points out. the increment of increase is very small. It is so small in fact that for all practical purposes it does not conflict with the conclusions of the paper. The authors appreciate Mr. McLachlan's comprehensive study of the data.

Jan 1, 1950

By S. R. B. Cooke, H. S. Choi, I. Iwasaki

Oleic acid is the chief ingredient of fatty acids used as collectors in nonsulfide flotation. With a few notable exceptions, the various quantities of saturated and other unsaturated acids comprising the remainder are usually unspecified. There have been occasional investigations to determine the collecting ability of these other acids and their effect on collection with oleic acid, but much remains to be done in this field. Hukki and Vartiainen, on the basis of tests using palmitic, oleic, and linoleic acids and an impure linolenic acid, concluded: "The collecting power of fatty acids used in flotation increases with increasing unsaturation of the hydrocarbon chain." ' Their work was concerned mainly with the response of ilmenite to the foregoing collectors, although tests were also made with magnetite, pyrite, hematite, and rutile. Cooke and Nummela, employing an almost identical technical procedure to evaluate flotation of hematite, magnetite, and goethite with various fatty acids, including oleic, linoleic, and linolenic acids of the highest purity obtainable, found: "The effectiveness of 18-carbon unsaturated fatty acids, as collectors for hematite, decreases with increase in the number of non-conjugated double bonds." 2 Vn the foregoing investigations, collector efficacy was measured on the basis of reagent addition in moles per liter. Sources of fatty acids which may be used as collectors are legion, but provided the oleic acid content is high enough, cost is normally the determining factor. Usually this restricts severely the purity of the collector in terms of its oleic acid content. Within the past year or so users of crude oleic acid have become somewhat more discriminating, apparently realizing that some of the organic acid impurities may exert deleterious effects upon certain separations. Marine fish oils have been used as collecting agents, particularly in the earlier days of froth flotation. Currently their use is restricted in this field to relatively minor applications, although they are cheap sources of fatty acids. In an effort to increase the range of uses for fish oils, the U. S. Fish and Wildlife Service has awarded contracts to study these materials as sources of new products suitable for commercial development. Funds for these investigations were provided by the Saltonstall-Ken-nedy Act of 1954. One such contract was awarded to the University of Minnesota School of Mines and Metallurgy for the purpose of investigating the use of fish oil fatty acids and their derivatives in ore flotation. Besides the esters of various saturated aliphatic acids and of palmitoleic and oleic acids, fish oil also contains esters of monoethenoid and polyethenoid acids with 20, 22, and 2 carbon atoms. The basic purpose of the investigation was to determine whether the longer-chained unsaturated acids possess collective properties superior to those of the C,, monoethenoid oleic acid, the C16 monoethenoid palmitoleic acid, or the saturated C,, stearic acid and its shorter chain homologues; whether highly unsaturated acids have merit as collectors; whether rela-

Jan 1, 1960

By M. Takeda, H. Majima

Pyrite has been subjected to xanthate adsorption experiments in order to determine the nature of adsorption. The identification of adsorbed species by means of IR spectroscopy, as well as UV, shows that the dixanthogen adsorption is undoubtedly predominant on pyrite surface. The rest potential measured in a xanthate solution coincides with the redox potential of the xanthate-dixanthogen couple. Considering all the factors together, the adsorption mechanism of dixanthogen on pyrite has been proposed as This mechanism is consistent with the "passive" character of pyrite. From the measurement of the redox potential of the xanthate-dixanthogen couple in the homologous series, a rule similar to that for surface tensions proposed by Traube is also applicable. The polarization experiments were conducted on pyrite in a solution containing xanthate. In the anodic polarization, xanthate is observed to oxidize to dix- anthogen, which depresses oxygen evolution on py rite at higher potential levels. In spite of the importance of the presence of pyrite in the flotation of sulfide minerals, not many studies have been made on the xanthate-pyrite system in the absence of an activator or a depressant. In an early study by Gaudin and Wilkinson,' it was pointed out that no ferric xanthate was formed on a pyrite surface, while in another study Gaudin and his collaborators2 proved the formation of a very small amount of ferric xanthate. Afterwards, Gaudin et al.3 studied the pyrite-xanthate system with a radiometric technique and found indications of the presence of adsorbed dixanthogen on pyrite surfaces. Mamiya4 studied the flotation behavior of the pyrite-xanthate system and also suggested the possibility of dixanthogen. Recently, Fuerstenau and his coworkers at The Colorado School of Mines5 studied the role of dixanthogen in the xanthate flotation of pyrite and discussed the problem in terms of such properties as oxidation potential, zeta potential, flotation recovery, pH of solution, etc. However, the mechanism of dixanthogen adsorption at the pyrite-water interface is not yet understood. This paper is concerned with an electrochemical in-vestigation through which the mechanism of dixanthogen adsorption on pyrite is interpreted. EXPERIMENTAL MATERIALS AND TECHNIQUES Pure potassium xanthates were synthesized by an

Jan 1, 1969

By J. M. W. Mackenzie

Recent successful applications of emulsion flotation suggested that a detailed study of some of the physicochemical aspects of the process was warranted. The objective of the current research was to provide general information on the electrokinetic properties of emulsion drops of the type commonly used in oxide mineral flotation. Combinations of the parafin oil Nujol and several long-chain reagents in water were studied with respect to the effect of pH on the surface and zeta potentials. It war concluded that for emulsion drops in the presence of a long-chain reagent but in the absence of any other electrolyte which is adsorbed at the oil-water interface, the surface potential will be some two to four times the value of the zeta potential. During the last decade there has been a revival of interest in the so-called agglomeration, or emulsion flotation (in this paper the term emulsion flotation will be used), process. In this flotation technique, a neutral hydrocarbon oil, together with longcnain reagents, usually of a polar-nonpolar structure, are added to the pulp to function as collectors for specific minerals. The combination of oil and longchain reagent is either emulsified prior to addition to the pulp, or in the pulp itself, and the resulting collector species is an oil-in-water emulsion of a neutral hydrocarbon oil emulsified by polar-nonpolar long-chain reagents. Emulsion flotation has been used successfully in the treatment of ores containing iron, manganese,' and titanium oxides. 2 Two other major and long established applications of emulsion flotation are the flotation of coal and Florida phosphate ores. In recent years, emulsion flotation has been particularly successful in the treatment of ores containing oxide minerals which slime readily and, consequently, are difficult to treat by conventional flotation processes. The recent successful

Jan 1, 1970