Search Documents

By Donald R. Rathburn, Jung K. Mok

An historical review and a survey of current practice in the application of fine dry grinding in the iron ore industry are given. Both closed and open-circuit applications are reviewed, and the advantages, disadvantages, and process considerations for these two systems are discussed. The ground ore temperature rise, and the cool-ing of ground ore and mill ventilation are also discussed, along with the agglomerating characteristics of dry ground ore. Screening during the production of high-grade lump ore for blast-furnace burden, particularly from recently developed ore deposits in Australia, South America, and Africa, has resulted in large tonnages of fine iron ore. Certain owners have decided to stockpile these fines until some later date. Others have either sold these fines "as is" for agglomeration at the steel plant or have established their own pelletizing plants to produce a premium quality agglomerate. In almost all cases some grinding is required prior to pelletizing. Most of these ores occur in the weathered state which may result in difficulty in the thickening and/or filtering stages when ground in a wet system. The resulting filter cake would probably require partial drying to provide a product suitable for the subsequent balling operation; however the partial drying and subsequent material-handling steps often produce random agglomeration. Normally, these agglomerates will cause problems in the green ball formation process. Thus dry grinding has developed as a solution to this problem. Most of the dry iron ore grinding installations utilizing the standard-type rod or ball mill are grinding high-grade natural ores that do not require beneficiation. Although power consumed per ton of final product is 120% or greater in dry grinding as compared with wet grinding, the additional power required for dewatering and handling a wet ground product can easily exceed the initial power savings. Dry grinding of raw feed and clinker has been practiced in the cement industry for many decades but its major application to the iron ore industry has taken place only within the past ten years. The successful application of dry grinding to iron ore has required the solution of several unique design and operating problems. The purpose of this paper is to review recent developments in fine dry grinding of iron ore for pelletizing and to discuss the various designs and operating parameters involved.

Jan 1, 1971

By R. T. Hukki

IT is no great exaggeration to say that present grinding practice and economics are largely determined by lining design. A record of outstanding liner wear can be achieved with any liner surface pattern that will positively lock the outer layer or layers of grinding media. With no slippage, lining wear is bound to be slight. At the same time, popular practice calls for tumbling loads of about 50 pct of mill volume to obtain maximum grinding capacity. Innumerable parallel grinding investigations have verified that optimum speed for such a mill lies within 70 to 85 pct of theoretical critical speed. If the mill is speeded up to 100 pct of critical, little or no grinding can be accomplished. Earlier Work on Supercritical Grinding: First investigations concerning grinding at supercritical speeds seem to be very old. Remarkable work on the subject has been performed by White,&apos; and his experiments have been described and summarized by Richards.&apos; White&apos;s contributions seem to have passed unnoticed by Fahrenwald,2,3 whose extensive experiments have been well presented, yet apparently very little appreciated. Additional work on grinding at supercritical speed has been reported, e.g., by Lewenson and Tscherny,&apos; USSR; Anselm and Grunder,&apos;, Germany: and Rose and Evans." Great Britain. Subcritical and Supercritical Speeds: In the formula of the critical speed given in the textbooks of mineral dressing, no factor indicating the coefficient of friction is generally included. If this factor = 1.0, which is equivalent to grinding conditions in a mill provided with heavily ribbed lining, the formula of the critical speed holds as such and grinding is possible at subcritical speeds only. In a mill equipped with a smooth or relatively smooth lining, the numerical value of the friction factor in the denominator of the formula becomes <1.0, indicating that grinding at supercritical speeds should be possible in such mills. It has been recently shown by the author&apos; that a wide supercritical speed range will become available for grinding—and especially for fine grinding—if the basic conditions within the mill have been selected properly. Mathematical analysis of mill dynamics at supercritical speeds&apos; has indicated that the mill speed may be increased if: 1) total mass of grinding medium decreases, 2) mass of the individual grinding piece increases, and 3) the coefficient of friction between the outer layer of medium and the mill lining decreases. It is obvious that the mass of the individual grinding piece will be affected by its shape and by its specific gravity. The coefficient of friction decreases with: 1) increasing smoothness of liner surface, 2) increasing roundness of the grinding piece, 3) increasing fineness of material to be ground, 4) decreasing pulp density, and 5) decreasing hardness (abrasiveness) of the mineral to be ground. In a mill equipped with heavily ribbed lining, practically no size reduction will take place between the lining and the outer layer of tumbling medium, because slippage is prevented. In the overwhelming majority of today&apos;s mills the grinding accomplished is by virtue of cataracting and/or cascading media with some action within the tumbling charge. Sub-critical speeds only can be applied. In grate-type or peripheral discharge mills equipped with a smooth lining a tumbling load of any kind of common medium occupying about 50 pct of mill volume will behave in such a way that practically no slippage will take place. This has been verified in experiments run by the author in the laboratory and in pilot plant mills&apos; equipped with smooth lining. Again, the operation is limited to the subcritical range. In a mill with a smooth lining, grinding will be possible either at subcritical speeds or within a wide supercritical speed range as soon as the basic requirements for a desired speed are fulfilled. In a mill operated at supercritical speed, any point on the liner surface proceeds at a speed greater than that indicated by the formula of the critical speed, while any grinding piece situated in the outer layer of the medium against the lining proceeds in the same direction at a speed less than that indicated by the critical speed. This speed difference produces a very effective attrition grinding zone between the liner surface and the outer layer of the medium. The share of attrition grinding of the total grinding accomplished increases rapidly with increasing speed in the supercritical speed range. The smooth surface of the mill lining may be well illustrated by the surface of a bucking board. The outer ball layer may be similarly represented by the lower surface of a muller. If the bucking board, the material to be ground, and the muller all proceed in the same direction at the same speed, no grinding will result. This is the general situation in the mills of today. If, however, the muller is pulled with respect to the board, moving or stationary, grinding will be accomplished effectively. Although grinding in today&apos;s mills is primarily the result of cataracting and/or cascading media, the principal place of grinding at high supercritical speeds will be the attrition zone.&apos; In the mineral dressing laboratory of the State Institute for Technical Research. Helsinki, Finland, large quantities of different ores have been ground in a pilot plant ball mill (3x3 ft) at a top speed about 230 pct of the critical. With the same mill, theoretical investigations concerning the grinding characteristics of ball mills equipped with a smooth lining have been carried out up to the speed of 313 pct&apos; of the critical. With a smaller mill, the grinding characteristics with a variety of grinding media have been investigated at speeds up to 2000 pct of the

Jan 1, 1959

By R. T. Hukki

John F. Myers (Consulting Engineer, Greenwich, Corm.)—Since the art of comminution has lain practically dormant for many years, it is very interesting that R. T. Hukki approaches the subject with a new concept. One is reminded of the research carried on by A. W. Fahrenwald of Moscow, Idaho, a few years ago. Fahrenwald mounted a steel bowl on a vertical shaft. The balls and ore placed in the bowl were rotated at fast speeds, thus simulating the supercritical speeds used by Hukki. The rolling action of the balls against the smooth shell liner has pretty much the same effect. The action is horizontal in one case and vertical in the other. Both researchers report good grinding activity. It is also constructive that such able investigators give to the students of comminution their interpretation of their laboratory results in terms of large-scale operation. History shows that it takes a lot of time for such radically new ideas to be absorbed by the industry. Typical of this is the present-day activity of cyclone classification in primary grinding circuits. The idea of cyclone classification has been kicking around for 30 or 40 years. Certainly we all suspect that the ponderous grinding mills of today, and their accessory apparatus, large buildings, etc., will ultimately give way to small fast units, just as this has occurred in other industries over the past 50 years. At the moment there is no evidence that ball and liner wear is prohibitively high. In fact, at the time Fahrenwald was demonstrating his high-speed horizontal machine at the meeting of the American Mining Congress, several years ago, he assured this writer that the balls retained their shape much longer than they do in conventional tumbling mills. Rods and balls that slide (as some operators in uranium plants are experiencing) get flat. Apparently the balls have a rolling action. Mr. Hukki&apos;s references to the processing capacity of the Tennessee Copper Co. mills is adequate. Those studying this subject will be greatly interested in the paper presented by Richard Smith of the Cleveland-Cliffs Iron Co. at the annual meeting of the Canadian Institute of Mining and Metallurgy in Vancouver April 24, 1958. This paper will be published during the latter part of 1958 in the Canadian Institute of Mining and Metallurgy Bulletin. Hukki&apos;s pioneering spirit is to be commended. R. T. Hukki (author&apos;s reply)—It has been heartening to read the objective discussion by J. F. Myers. The sincerity of his opinions is further strengthened by the fact that the article he has discussed contradicts in a major way the parallel achievements of his life work. Myers is right in his opinion that in general it takes a long time before new ideas are accepted by the industry. On the other hand, revolutions usually take place at supercritical speeds. There are many indications at present that both the unit operation of grinding and the related subject of size control are now just about ripe for a revolution. In grinding, brute force must ultimately give way to science. Rapid progress can be anticipated in the following fields: 1) Autogenous fine grinding at supercritical speeds will be the first advance and the one that will gain recognition most easily on industrial scale. At this moment, little Finland appears to be leading the world. Crocker recently made a statement that in nine cases out of ten, your own ore can be used as grinding medium more effectively and far more economically than steel balls. This is true. The present author would like to introduce a supplementary idea: In eight cases out of the nine cited above, it can be done at the highest overall efficiency in the supercritical speed range. Fine grinding must be based on attrition, not impact. The path of attrition may be vertical, horizontal, even inclined. 2) In coarse grinding, the conventional use of rods is sound practice. However, even the rods can be replaced by autogenous chunks large enough to offer the same impact momentum as the rods. To obtain the momentum, the chunks must be provided with a free fall through a sufficient height in horizontal mills operated at supercritical speeds. Coarse grinding must be based on impact. Detailed analysis of the subject may be found in a paper entitled "All-autogenous Grinding at Supercritical Speeds" in Mine and Quarry Engineering, July 1958. 3) All conventional methods of classification, including wet and dry cyclones, are inefficient in sharpness of separation. Continuous return of huge tonnages of finished material to the grinding unit with the circulating load is senseless practice. In the near future the present methods will be either replaced or supplemented by precision sizing. These three fields are also the ones to which J. F. Myers has so admirably contributed in the past. Fine Grinding at Supercritical Speeds. By R. T. Hukki (Mining EnGineERInG, May 1958). Eq. 9, page 588, should be as follows: T , c, (a — 6&apos;) n D Ltph On page 584 of the article the captions for Figs. 4 and 5 have been placed under the wrong illustrations and should be interchanged.

Jan 1, 1959

By A. D. Dorenfeld

Factorial design is a mathematical method of drawing valid conclusions from a series of tests made in a predetermined pattern. It is applied to flotation ore testing using, in this case, five variables, each varied twice, requiring a total of 32 tests. Representative conclusions are checked by repeat testing. THE factorial design method of testing is particu-J- larly applicable to mill-scale plant tests because 1—the conclusions take into consideration all variabilities of the plant, 2—there are no subjective guesses as to the conclusions, 3—the number of tests necessary to check large numbers of variables are at a minimum, and 4—interactions between variables can be discerned. It is equally applicable to laboratory testing because once the variables to be investigated have been decided upon, as well as the number of levels for each variable, a well-trained assistant can run the tests, since the test work follows a predetermined mathematical pattern, leaving the conclusions to be found by the person in charge. The conclusion takes into consideration all errors of the tests. To illustrate the use of factorial design, a sample of copper porphyry ore was tested, the copper content of the batches not being uniform. Five variables were investigated, each at two levels, requiring only 32 tests. After suitable arithmetical analysis, conclusions for copper recovery and concentrate grade are arrived at. Four cases of these conclusions are checked by repeat testing. Testing Methods Reproducibility of flotation tests varies widely, depending on the experimenter, the process, the equipment, and probably a host of unknown and perhaps interacting variables. If the reproducibility of tests is good, let us say to 0.2 pct recovery of a certain metal, and the minimum difference between tests upon the variation of the interested variables is, say 1 pct, then simple averaging and inspection is ample in interpreting results. However, often such a happy situation is not the case. This is particularly true in mill-scale operations where the ores are variable, and a host of minor operating fluctuations common to most milling, and familiar to all mill operators, is the normal state of affairs. Yet, mill scale tests are run, the results averaged, and conclusions drawn therefrom. The trustworthiness of such results should be measured by comparison to the experimental error—the measure of the degree of reproducibility of the results. This comparison is a test of significance. Thus, if daily results of recoveries average 80.0 pct, and a spread of 78.0 pct to 82.0 pct, is recorded, then, clearly, on changing mill operations, an 81 pct recovery cannot be construed as significant by using averages, since it is well within the previous test limits. By using rigorous statistical tests of significance, conclusions arrived at by inspection or simple averaging may be shown to be based on inadequate evidence or even wrong. In any event, a plant super-

Jan 1, 1952

By M. E. Wadsworth, I. B. Cutler

Coprecipitation of anionic and cationic polyelectrolytes has been applied to floccula-tion of several mineral systems. Results obtained in a study of the flocculation of kaolinite and hematite suspensions of polycationic and polyanionic electrolytes are presented. Greatly increased settling rates were observed following precipitation of positive and negative polyelectrolytes on the surface of finely divided minerals in aqueous suspension. The ratios of polycationic to polyanionic electrolytes required to produce maximum sedimentation have been shown to correspond closely with the equivalence points obtained by light scattering studies of systems containing the positive and negative polyelectrolytes by themselves. DISPERSION of colloids has been of interest to surface chemists for many years. Flocculation studies have been employed to evaluate the limits of colloid stability. In water suspensions, colloids become unstable and flocculate when the ( potential, the potential of the kinetic solvation sheath, falls to a value at which van der Waals forces may act to draw the particles together. This may happen through an adjustment of the pH of the suspension or by addition of a salt. Optimum conditions for rapid formation of large flocs are found in the region of minimum ( potential. In addition to these well known considerations, colloids in water suspensions are known to be flocculated by certain large organic macromolecules. These polyelectrolytes in some cases minimize the i potential and in other instances increase the attractive forces by supplying sites for hydrogen bonding.&apos; Use of organic polyelectrolytes in flocculation of mineral colloids has increased rapidly in recent years. These materials are now recognized as indispensable aids to fluid-solid separations in many extractive metallurgical operations. They have been found to increase both the settling rate in thickeners and the filtration rate of filters. In the elucidation of the characteristics of these organic reagents as floc-culants in this laboratory, it was discovered that interaction of certain polyelectrolytes had a surprisingly good effect on flocculation of mineral colloids. Experimental Procedure: The anionic flocculants used in this study were Lytron 886 and 887, produced by Monsanto Chemical Co. Both reagents are copolymers of vinyl acetate and maleic anhydride. Lime is added to Lytron 886, whereas Lytron 887 is all organic. These reagents possess two carboxyl groups for each acetate group and are therefore anionic in character. The cationic reagents used were the Peter Cooper Co. 1-X and 2-X glues. The structures are not definitely known, but the re- agents are amines and are cationic in character. The glues were prepared in the chrome-alum form. Both types of reagents may be classified as polyelectrolytes in that they are high molecular weight polymers and ionize in water. All data presented are in terms of sedimentation rate, measured in 250-ml glass-stoppered graduated cylinders. With suitable precautions such measurements are fairly reproducible and provide a sensitive means for selecting important parameters such as pH, reagent concentration, and floc size and density.2 Agitation was standardized with five complete end over end cycles of the suspensions in the stoppered graduated cylinders.

Jan 1, 1957

By R. H. Oliver

The purposes, types, preparation, and testing of flocculants are discussed. A flocculation compendium is included, indicating choice of flocculant for a given set of conditions. An economic evaluation of the process is presented. Solid-liquid separation is a major expense in most mineral processing flowsheets today. Proper use of flocculation can often lower the overall cost of a solid-liquid separation by reducing the size of sedimentation or filtration equipment. PURPOSE OF FLOCCULATION When is use of a flocculant justified in terms of overall process economy? 1) when it results in process improvement, such as producing a clear liquor for electrolysis, precipitation, ion exchange, or solvent extraction; 2) when it results in satisfying requirements of pollution abatement ordinances; 3) when it results in increased recovery of values that would otherwise be lost; or 4) when it results in considerable savings in capital expenditure due to use of smaller equipment. When can the use of a flocculant adversely effect overall process economy? 1) when it results in increased filter cake moisture, 2) when it results in decreased filtration rates, 3) when it results in considerable bulking of sedimentation underflow, 4) when it contaminates either the supernatant or sludge, or 5) when the cost of using the flocculant is greater than the savings due to its use. First step in consideration of flocculants for a given application is establishment of desired goals in terms of increased clarity of effluent, increased recovery of solids, or decreased size of equipment. The most economical solution to a solid-liquid separation problem is that combination of equipment and flocculant costs which meets the established goals at the lowest total annual expenditure. This economic decision can be made only after technical data has been obtained as outlined in next four sections. TYPES OF FLOCCULATION Flocculation is a process wherein individual particles are united into more or less tightly bound agglomerates or flocs, thereby increasing effective particle size of solids suspended in a liquid. Degree of flocculation of a suspension of finely divided solids in a liquid is controlled by a combination of probability of collision between particles and probability of adhesion after the collision has occurred.&apos; Probability of collision can be increased commercially through use of a paddle-type flocculator, combination flocculator, and clarifier or flocculating-type feedwell. Probability of adhesion usually can be increased by addition of a reagent known as a flocculant. Reagents act as flocculants through one or a combination of three possible mechanisms. The first is electrolytic neutralization of inter molecular repulsive force due to Zeta potential. This neutralization enables Van der Waal&apos;s cohesive force to hold the particles together after they collide.&apos; The second is the precipitation, within a definite pH range, of voluminous metallic hydroxide flocs which entrap fine particles. This process is known as coagulation.14 Third is the bridging of two or more particles by either natural or synthetic long-chain, high-molecular-weight organic polymers. This bridging is accomplished by adsorption of two particles at different sites on the same molecule or by bonding of two molecules, each adsorbed to a different particle. Considerable effort has been expended in explaining the action of these polyelec-trolytes and detailed accounts are available in the literature. CHOICE OF FLOCCULANTS It is impossible, at the present time, to specify the optimum type or amount of flocculant for a given application from a knowledge of the material to be treated. The standard procedure is to choose reagents or combination of reagents most likely to be effective on the slurry; then test them.&apos; The accompanying Flocculation Compendium (Table I) contains details of all currently available reagents sold as flocculants and it should be used when selecting the group of reagents for tests. This selection can be based on information listed in columns 4 and 5. Column 4 presents industries in which each floc-

Jan 1, 1961

By R. Schuhmann, J. Th. Overbeek, P. L. De Bruyn

THE technique of concentrating valuable minerals from lean ores by flotation depends upon the creation of a finite contact angle at the three-phase contact, mineral-water-air. If the mineral is completely wetted by the water phase, contact angle zero, there is no tendency for air bubbles to attach themselves to the mineral. However, when the contact angle is finite, the surface free energy of the system, water-air bubble-mineral particle, can be diminished by contact between the bubble and the particle, and if not too heavy the mineral will be levitated in the froth. With a few exceptions, all clean minerals are completely wetted by pure water. Thus the art of flotation consists in adding substances to the water to make a finite contact angle with the mineral to be floated, but to leave the other minerals with a zero contact angle. The contact angle concept and experimental measurements of contact angles have played important roles in flotation research for several decades.&apos;-" Nevertheless, there remain unanswered some basic questions as to the scientific significance of the contact angle and the nature of the processes by which flotation reagents affect contact angles. The contact angle is a complex quantity because the properties of three different phases, or rather of three different interfaces, control its magnitude. Considering the interfaces close to the region of ternary contact to be plane, the relation among the contact angle and the three binary interfacial tensions is easily derived. The condition for equilibrium among the three surface tensions, Fig. 1, or the requirement of minimum total surface free energy leads to Young&apos;s equation, Eq. I: ysa — ysl = yLA cos 0 [1] According to this equation, the contact angle has one well-defined value. Actually it is found in many experiments that the value of the contact angle depends on whether the air is replacing liquid over the solid (receding angle) or the liquid is replacing air (advancing angle). The receding angle is always the smaller of the two.4 Two explanations have been offered for this experimental fact. According to some investigators,5-8 roughness of the surface causes apparent contact angles that are different for the receding and the advancing cases although the actual local contact angle may be completely determined by Eq. 1. The other explanation involves the hypothesis that the solid-air interface after the liquid has just receded is different from the same interface when no liquid has previously covered it.1,4 Adsorption of constituents of the air or liquid might play a role here. In this discussion the difference between advancing and receding contact angle will be neglected and plane surfaces where Eq. 1 describes the situation will be considered. But there is still a fundamental obstacle to the application of Young&apos;s equation. The surface tension of the liquid (rla) can easily be determined, but the two surface tensions of the solid (rsa and ySL) cannot be measured directly. Eq. 1, however, is not without value. By contact angle measurements it is possible to establish how ysl — ysl varies with the addition of solutes to the liquid phase. Also, Eq. 1 affords a convenient starting point for calculating net forces and energy changes involved in the process of bubble-particle attachment.1,2 . If for the moment surface tension of the liquid (yLa) is considered a constant, an increase in ysa — ysL, will tend to decrease the contact angle. A decrease in ySA — ysl, corresponds to an increase of the contact angle. In cases where ySA — ySL > yLa the contact angle is zero; it will only reach finite values when ysa — ysa has been decreased below YLA. Thus on the basis of Young&apos;s equation and contact angle measurements alone, it can be learned how flotation reagents affect the difference Ysa — ysl, but no conclusions can be drawn as to the effects of reagents on the individual surface tensions ysa, and ysL, not even as to signs or directions of the surface tension changes resulting from reagent additions. A quantitative relationship between the surface tension or interfacial tension and the adsorption occurring at a surface or an interface is given by the Gibbs equation, which for constant temperature and pressure reads dy = — 2 T, du, [2] where dy is the infinitesimal change in surface tension accompanying a change in chemical potential

Jan 1, 1955

By R. Schuhmann, J. Th. Overbeek, P. L. De Bruyn

W. E. Ewers (Commonwealth Scientific and Industrial Research organization, Melbourne, Australia)— Any attempt to elucidate further the meaning of the contact angle, particularly if it deals with the magnitudes of those mysterious quantities rsa and YRL, is certain to receive attention from those interested in the fundamentals of flotation. It is imperative, therefore, that attention should be drawn to a serious defect in the treatment by de Bruyn, Overbeek, and Schuhmann. Their Eq. 5 dy = — ld/u — Th2odiH2n states correctly the relationship between the change in surface free energy dy, the surface excesses Tx and Th2o of the components x and H2O and the changes du, and dph2o in the chemical potentials of these components. This equation applies for constant pressure and con- stant temperature and for the interface of the solid in contact with the solution of X. The relation remains valid when 1.1 is defined so that TH2O becomes zero and the equation becomes dr = — r:dur [17] From this equation de Bruyn et al. correctly concluded that increased adsorption of the collector at the liquid solid interface would reduce the interfacial energy ysl. Their second conclusion, however, that an increase in the contact angle, corresponding to a greater decrease in rsa than in rsL can be attributed to greater adsorption of X at the solid-air interface than at the solid-liquid interface is quite unjustified. The portion of the solid surface from which the collector solution has been displaced by air must be considered in terms of the equilibria between the gaseous phase and the surface. If the pressure and temperature

Jan 1, 1955

By V. I. Purcell, S. C. Sun. R. E. Snow

A study including effects of 7) pH value, 2) fatty acid collector, 3) fuel oil, 4) interfering ion, 5) particle size, and 6) operational variables. Test results indicate feasibility of fatty acid flotation under certain restrictive conditions. FATTY acids have been used as standard collectors for floating the matrix1-8 but not the leached zone phosphate ores. After a series of investigations, Davenport9 and Tarbutton10 concluded that the Florida leached zone ores tested were unamenable to fatty acid flotation and must be upgraded by floating the siliceous gangue with Armac-T as collector. The cationic flotation was estimated11 to cost almost ten times as much as a fatty acid process that would ideally give the same flotation grades and recoveries of phosphates. Mineralization: Extensive deposits of leached zone aluminiferous phosphate occur in Florida."= The ore sample used for this work was obtained from the Virginia-Carolina Chemical Corp., Clear Springs area, Florida. The ore is a potential source of uranium and aluminum as well as phosphorus. Since the uranium and the recoverable aluminum are contained in some of the phosphate minerals, an upgrading of phosphate results in a corresponding enrichment of both uranium and aluminum. The phosphate minerals in the ore are primarily pseudo-wavellite, millisite, and carbonate fluorapatite, with a minor amount of wavellite. Quartz and clay are the common gangue minerals, quartz being predominant. Others present in small amounts are ilmenite, hydrated iron oxides, staurolite, monazite, and zircon. Experimental Procedures: After being dried to less than 1 pct surface moisture the material was dry-ground in a Hardinge conical ball mill to —6 mesh. The ground product was classified into 6/20, 20/200, and —200 mesh size fractions. On the basis of Table I, the coarse and fine size fractions were considered as phosphate concentrates, because of their high phosphorus contents. In contrast, the 20/200 mesh fraction, which was low in phosphorus, was used as the flotation feed. All the flotation tests, unless otherwise stated, were performed in a laboratory Fagergren machine under the following conditions: 1) 400 g of solid feed for each test, 2) 2000 rpm impeller speed, 3) 8.9 pH and 4-min conditioning, and 4) 3-min flotation. The procedure used for the quantitative bubble pick-up tests has been described previously.18

Jan 1, 1958

By M. Sagheer

This paper gives a report of the laboratory work investigating the fundamental flotation characteristics of chromite and its commonly associated mineral, serpentine. Floatability curves, showing floatability of chromite at different concentrations of sodium oleate and pH, were obtained by pick-up method. One of the difficulties encountered in flotation of chromite in presence of serpentine has been attributed to dissolved metal ions. The influence of metal ions, dissolved in pulp filtrate, on floatability of pure chromite and serpentine has been studied and it has been shown that Mg, Ca and Fe ions depress chromite. A considerable amount of work applying bubble pick-up method was carried out to counteract the depressing action of metal ions. Various methods and reagents were applied in this connection. Validity of each method and reagent has been discussed. Probably due to the abundance of high-grade chrome ores, very little has been published concerning the flotation of chromite. As improvements in mineral beneficiation are realized, more attention is being directed to low-grade ores previously considered unusable. Currently, gravity concentration is the principal process applied to upgrade chrome ores, but this cannot be used for finely disseminated ore because of the substantial losses in the fines,1,2,3 To avoid these losses, flotation can be applied to material which is too fine for satisfactory gravity concent ration, and it is probable that a combination

Jan 1, 1967

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

In a previous article1 the function of various fatty acids as collectors for iron ores was reported for the two alternate processes; (a) the flotation of iron-oxide minerals, and (b) the flotation of calcium-activated siliceous gangue. The selectivity of separation was shown to be markedly dependent not only on such common parameters as pH and the levels of collector and activator addition, but also on the structure and the degree of unsaturation of the acids used. The iron ores employed consisted primarily of hematite, goethite, and quartz and as the effectiveness of any one acid on the separation depends on the individual response of each constituent mineral toward the given collector, a comprehensive knowledge of the floatability of the individual minerals is desirable for the interpretation of the results obtained. The mechanism by which a collector ion becomes attached to a mineral-solution interface has been a subject of much investigation. From quantitative studies of the properties of the electrical double layer on mercury2 and on several solid surfaces3,4,5,6 , the adsorption of some flotation reagents at the mineral-solution interface is beginning to be understood. For example, it has been demonstrated from streaming potential measurements made on quartz5 and corundum6 that the interfacial electrical condition is strongly dependent on the pH. Thisoccurs presumably through the following electrolytic reaction at the interface, and in the presence of a flotation collector, controls the floatability of the solids. A concurrent investigation of the flotation behavior of goethite and of quartz carried out at the School of Mines and Metallurgy7 has shown a positive correlation with the electrokinetic behavior of these minerals. The purpose of the present investigation was to collect general information on the floatability of hematite, and to investigate the function of various 18-carbon aliphatic acids on the floatability of hematite, goethite, and of activated quartz. Electrophoretic mobility measurements, simplified flotation tests using a modified Hallimond tube, and contact angle and frothability measurements were employed. The results are of direct application in the interpretation of the batch flotation results on natural ores as

Jan 1, 1961

By D. A. Rice, M. C. Fuerstenau

Flotation data indicate that sulfonate and amine adsorb physically on manganese oxide; oleate also adsorbs physically if the zero point of charge is sufficiently high. Chemisorption of oleate occurs on pyrolusite between pH 8 and 9, and possible mechan- isms of collector adsorption involving metal ion hydrolysis are discussed. The flotation characteristics of most metal oxides1-15 have been examined with standard collectors such as fatty acids and sulfonates. From these studies it would appear as though electrical phenomena are important when shorte r-chained collectors are involved. For example, Iwasaki, Cooke, and Choi9 have shown that dodecyl amine and dodecyl sulfate float iron oxide above and below the zero point of charge, respectively. In another system, D.W. Fuerstenau and Modi10 have shown that dodecyl sulfate, dodecyl

Jan 1, 1969

By Strathmore R. B. Cooke, Iwao Iwasaki, C. S. Chang

The effects of aeration on an aqueous suspension of pyrrhotite were studied and their results correlated with flotation tests using xanthates as collectors. The effects of copper activation and of pH variation were determined and possible mechanisms postulated. PYRRHOTITE has long been considered a gangue mineral to be eliminated as tailing in the treatment of various sulphide ores. However, in recent years the world-wide lack of sulphur resources has called attention to this mineral as a potential source of both sulphur and iron. Its importance as an economic mineral, however, has not been particularly emphasized. For this reason very little is known about its response to flotation, except that it can be depressed easily in alkaline circuit, by long aeration,1,2 addition of oxidizing agents,3 or by starch.&apos; The object of this work was to study the floatabil-ity of pyrrhotite. This includes the effect of oxidation by aeration, of copper activation, and of change in pH. Preparation of the Pyrrhotite Sample: It was desirable that the highest grade of pyrrhotite obtainable be used for this experiment, since the presence of other minerals could affect the surface properties.5 However, no pyrrhotite was available as crystals, and massive deposits of hydrothermal origin commonly contain considerable amounts of chalcopyrite. Pyrrhotite concentrate was, therefore, prepared from a sulphide deposit occurring near Aitkin, Minn. The deposit is of pyrometamorphic nature consistirlg mainly of pyrrhotite and pyrite with graphite, silicates, and carbonates as gangue. The ore, already crushed through 3 mesh when received, was screened at 65 mesh and the undersize discarded. The oversize was crushed through rolls, and then stage-ground dry in an Abbe porcelain mill, the —65 mesh portion being screened out after every 15 min of grinding until all the material passed through this size. The ground product was then concentrated with a drum-type dry magnetic separator. The rougher concentrate was cleaned twice and then demagnetized. The final product was split in a Jones splitter and stored in air-tight bottles. Microscopic examination of the concentrate showed that it was relatively clean and free of pyrite, locked particles, and gangue. By means of the krypton gas adsorption method," the specific surface was determined to be 3000 cm2 per g. The chemical and screen analyses of the final concentrate are given in Tables I and II respectively. It is a well-recognized fact that the oxidation of some sulphide ores during stockpiling, grinding, and conditioning affects their flotation behavior. The problem of oxidation may become serious in the case of pyrrhotite, since this is known to be more easily oxidized than many other sulphides. To ascertain the extent of oxidation, an experiment was carried out by aerating an aqueous suspension of pyrrhotite with air, oxygen, and nitrogen as follows. A 300-g sample of pyrrhotite in 2700 ml of water was agitated and simultaneously aerated in a Fager-gren-type laboratory flotation machine. A Precision wet test meter was connected to the air inlet valve, the flow rate of the gas being kept constant at 0.3 cu ft per min throughout the experiment. Samples of approximately 30 ml each were taken from the cell at 0, 4, 10, 20, 35, 60, and 90 min. After the pH was taken, each sample was filtered and the filtrate was analyzed for total iron and sulphur. The iron was determined colorimetrically by the thioglycolate method using a green filter.&apos; The filtrate was oxidized with bromine to convert all of the soluble sulphur compounds into sulphate and this was determined with a Parr turbidimeter." When aeration tests were made in alkaline circuit, calcium hydroxide or sodium hydroxide was added at regular intervals to maintain a constant pH. A similar procedure was followed in an experiment to determine the abstraction of copper. ion by pyrrhotite. In this case various quantities of cupric chloride were added. The filtrate from each sample taken was analyzed for copper, total iron, and sulphur. The carbamate method with a green filter was used for the copper analysis,&apos; since this method could tolerate a considerable amount of iron in the solution. A pneumatic cell, made from a 350-ml fritted glass Buechner funnel, was used for this experiment. The detail of the assemblage has been described elsewhere." In the present work a stainless steel baffle was inserted in the cell. This baffle overcame the tendency for the coarse pyrrhotite particles to be swirled around the wall of the cell and thus fail to collect in the froth. A 50-g sample of pyrrhotite was added to the cell which contained 260 ml of water. When pretreat-ment of the sample was desired, reagents, such as activator and pH regulator, were then added and the pulp was conditioned for a specified conditioning time. Prior to the addition of the collector approximately 15 ml of the solution were removed for pH measurement and for iron and sulphur analyses. Copper when used as activator was also determined. Collector and frother were then added and the pulp was conditioned for an additional 2 min. Air was admitted to the cell and the froth removed. The separation required from 4 to 6 min, depending on the characteristics of the froth. The float and non-float products were filtered, dried, weighed, and assayed for iron.

Jan 1, 1955

By R. A. Wyman

KYANITE schists in the Sudbury area have been generally described by Haw,&apos; who has also given particuLar information on preliminary treatment of three large samples from Dryden township, Ontario.&apos; Kyanite, garnet, biotite, quartz, and feldspar comprise the main ingredients of this ore. Kyanite is the valuable mineral. About 5½ tons of —35 +200 mesh product, made by dry rod milling, screening, and separation on the Exolon magnetic separator, were available for treatment by flotation. Magnetic separation had effectively reduced garnet and biotite so that flotation feed consisted essentially of kyanite, quartz, and feldspar, the kyanite content averaging 32 pct and representing about 86 pct of the kyanite in the original ore. Earlier work on kyanite at the Mines Branch, Ottawa, s-c employed an acid circuit, but to avoid production difficulties in the use of acids and at the same time take advantage of the fact that industrial minerals respond somewhat to soap flotation, it was decided first to investigate the higher pH levels. Exploratory tests using —65 +200 mesh feed ended with production of 96.6 pct kyanite concentrate rep-

Jan 1, 1959

By Adrian C. Dorenfeld, Theodore Balberyszski, Strathmore R. B. Cooke

This paper reports results of studies of sulfidiza-tion of base-metal oxides and silicates with gaseous sulfur, hydrogen sulfide gas and pyrite and of their subsequent flotation with xanthate collectors. It has been established that a 100% conversion of the oxides of lead, zinc, copper and manganese into their respective sulfides is possible and is a function of reaction time and temperature. Sulfidization of chrysocolla (with malachite and azurite inclusions) resulted in a product containing covellite, chalcocite and digenite in proportions depending on conditions of sulfidization. Reacting cassiterite with gaseous sulfur, H2S or pyrite at various reaction temperatures resulted in its partial conversion to a tin sulfide. Flotation of the products of sulfidization indicated that their behavior is similar to that of natural sulfide minerals, the recovery being a function of the con-version of the oxides to their respective sulfides. The gradual depletion of high-grade sulfide mineral deposits has turned the attention of the mineral industry to the recovery of metals from the oxides and silicates. Anionic (fatty acid) and cationic collectors have in some cases made flotation of nonsulfides possible, although, in general, flotation of oxide ores is not as yet an economic process. Consequently, many attempts have been made to sulfidize the surface of the relevant oxide mineral and thus make it amenable to to In recent years, Bautista and sollenberger4 have investigated the conversion of a number of oxides to sulfides by reacting them with molten sulfur. This paper presents the results of .investigations carried out at the School of Mineral and Metallurgical Engineering at the University of Minnesota to study the flotation characteristics of some sulfidized minerals. It is divided into two parts: The first describes the sulfidization of a number of oxides and a silicate, and the second presents the flotation behavior of the products of sulfidization. SULFIDIZATION OF METAL OXIDES AND SILICATES Experimental Procedure: SULFIDIZATION WITH HYDROGEN SULFIDE GAS — Sulfidization with H2S gas was carried out in a horizontal tube furnace (Fig. 1). The temperature of the furnace was regulated by means of a Variac and determined with a chromel-alumel thermocouple. The constant temperature zone in the furnace was about 6 in. in length, and it was within this zone that the experiments were carried out. The gases employed were supplied from standard compressed gas cylinders, and passed through a glass-wool filtering tower before entering the Pyrex reaction tube. Gaseous products were bubbled through a water tower (used as a flow indicator) and excess H2S was burned off at the end of the outlet tube. Samples of reagents or minerals were placed in a 3-in. Alundum boat and inserted within the constant temperature zone of the furnace. Nitrogen gas (6 ppm O2 content) was passed through the tubes at room temperature to remove air in the tube, and continued to pass while the furnace was heating up. When the required temperature was reached, H2S was allowed to enter the tube and the supply of nitrogen was shut off. At the end of the reaction period, the reverse procedure was carried out, the samples being cooled to room temperature in a nitrogen atmosphere. The flow rate of H2S gas was maintained fairly constant in all experiments, although the rate was not determined. SULFIDIZATION WITH SULFUR POWDER - To study the sulfidization of minerals with sulfur (liquid or gaseous), samples of the mineral, together with a measured quantity of sulfur powder, were placed in a stainless steel (SS316) reaction vessel (Fig. 2). The reaction vessel was heated in a vertical resistance coil furnace, the temperature being regulated through a Variac. At the end of the experiment the vessel was cooled to room temperature before the sample was removed. Cooled products were leached with CS2 to remove any free sulfur. SULFIDIZATION OF TIN CONCENTRATES - Fifty gram samples of sulfidized tin ore were prepared by mixing -48 mesh Bolivian tin concentrate with -48 mesh Ottawa sand in various proportions and sulfi-dizing these samples with H2S in the horizontal tube

Jan 1, 1969

By C. R. Ramachandra, C. C. Patel

Flotation of chalcopyrite from a low grade ore was studied by using different xanthates and dixanthogens as collectors and by conditioning the flotation pulp with oxidizing gaseous systems. The improvement in the recovery of chalcopyrite with the oxidizing systems was mainly due to the auto-activation of chalcopyrite, leading to the enhanced adsorption of xanthates and dixanthogens even when the latter were formed by oxidation of the xanthates. With air or oxygen as oxidizing system, the amount of ethyl xanthate required at pH 8.5 was much less than normally employed in the industry, the recovery being 97-99% of chalcopyrite with a concentrate of 24% copper. With isopropyl and isoamyl xanthates, the recovery was improved but the grade was lowered. By conditioning the pulp with air-CO2 or oxygen-CO2 mixture, the recovery was lowered with ethyl xanthate only, mainly due to the depressant action of bicarbonate formed. Flotation with dixanthogens gave cleaner concentrates of chalcopyrite. because iron pyrite did not float with these collectors. Of the dixanthogens tried, diisopropyl dixanthogen gave the highest recovery, 99% chalcopyrite with a concentrate containing 28.5 to 30% copper, when the pulp at pH 8.5 was condi-tioned with air. Isopropyl dixanthogen was found to be the best collector of all the xanthates and dixanthogens studied. It was found that the adsorbed ethyl, isopropyl and isoamyl xanthates at chalcopyrite surface, activated by cupric sulfate, and at iron pyrite surface were oxidized to the corresponding dixanthogens to different degrees in presence of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture.&apos; The dixanthogen formed at chalcopyrite surface was, however, firmly adsorbed while that at iron pyrite surface was easily desorbed. Since dixanthogens formed have greater hydrophobicity than the corresponding xanthates, as revealed by contact angles,1,2 the former were expected to improve the flotation results. It was, therefore, thought desirable to find out the effect of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture on the selective flotation of chalcopyrite when individual xanthates were used as collectors. The effect of direct use of dixanthogens as collectors on the selective flotation of chalcopyrite was also investigated. For these studies, low grade chalcopyrite ore, received from the Indian Copper Corp., Ltd. (I.C.C.), Ghatsila, Bihar, was employed. At the milling plant of the I.C.C., chalcopyrite is floated at a pH of 8.5, employing ethyl xanthate as collector. EXPERIMENTAL Materials Employed: XANTHATES - Ethyl, isopropyl and isoamyl xanthates of potassium were prepared in a pure state by the methods described earlier.4 A 0.03% aqueous solution of a xanthate was freshly prepared in air-free distilled water for flotation studies. DIXANTHOGENS - Diethyl, diisopropyl and diisoamyl dixanthogens were prepared by the oxidation of xanthates by iodine and purified by extraction with petroleum ether. A 0.03% alcoholic solution of dixanthogen was used as a collector. FROTHER - A 0.2% solution of pine oil in ethyl alcohol was employed. SODIUM CARBONATE - A 0.4 N solution was used. ORE - The composition of the -100 +200-mesh size (Tyler standard) powdered ore, as determined by standard methods of analysis, was: copper 1.57%, iron 10.29%, sulfur 3.15% and acid insolubles 66.10%. Iron present as chalcopyrite (CuFeS2) was 1.38%, while that present as iron pyrite (FeS,) was 1.37%. Both chalcopyrite and iron pyrite were found to be liberated from the gangue minerals and from each other, as observed under a microscope. Flotation of Chalcopyrite with Xanthates in Presence of Different Gaseous Systems: Fifteen grams of -100 + 200-mesh size ore were shaken well with 100 ml of distilled water. The pH of the pulp was adjusted to 8.5 by the addition of sodium carbonate solution. Then 0.1 ml of xanthate solution (0.005 lb xanthate per ton of the ore) was mixed with the pulp, and the latter together with washings transferred to a

Jan 1, 1963

By J. E. Papin

MORENCI ores contain as an average about 0.015 pct molybdenite, MoS2. Incidental to the concentrating operations applied for the recovery of copper minerals, approximately two-thirds of the molybdenite is floated and appears in the final copper concentrate. The economic importance of molybdenum and the success achieved in its recovery in milling operations elsewhere encouraged research directed toward its recovery in marketable form. Several procedures are utilized to effect separation of molybdenite from copper and iron sulphides, but only two have wide application. In the better known of these two methods the molybdenite in the copper concentrate is depressed in a flotation operation using soluble starch as the depressant. The tailing from this treatment thus becomes a low-grade molybdenite concentrate which after thickening, filtering, and low temperature roasting is repulped with water and subjected to additional flotation steps for recovery of the molybdenite. A thio-phos-phate collecting agent is employed in flotation of the copper minerals. Presumably the stability and lasting effects of that collector type necessitated the practice of depressing molybdenite followed by subsequent roasting to insure elimination of copper and iron sulphides from final molybdenite concentrate. The second important method of recovery is applied at various southwestern plants. In those plants xanthates are in use as copper collecting agents. It has been found that xanthates are relatively unstable and their collecting power for sulphides is destroyed by very simple procedures. The method generally applied is to subject the xanthate-acti-vated concentrate, in the form of a pulp, to prolonged heating using steam as the heating medium. A concentrate thus treated may then be subjected to a flotation operation for recovery of the molybdenite, using mineral oil collecting agents, and the copper and iron sulphides will be depressed. Neither of the above processes was applicable to Morenci concentrates. Morenci uses a thio-phos-phate collecting agent in its flotation operation. However, it was established that soluble starch was not a depressant for Morenci molybdenite; therefore the first step in that process was not applicable. AS would be expected the copper and iron sulphides in Morenci concentrates could not be depressed by the heating methods used on xanthate activated concentrates. Low-temperature roasting did eliminate the effects of the thio-phosphate but had to be rejected as uneconomic in as much as the tonnage of concentrates involved was too great. Preliminary research having established the above facts, it became obvious that new methods would have to be developed for the recovery of molybdenite from Morenci concentrates. A broad laboratory investigation was initiated along two general lines, both directed toward depression of copper and iron sulphides. One approach was through the use of water-soluble oxidizing agents with the objective of destroying the thio-phosphate in the concentrate and thus eliminating its collector effect. The other approach involved the use of soluble sulphides, which have long been recognized as sulphide depressants. These efforts resulted in laboratory processes which showed promise. Research was then extended to a pilot plant having a daily capacity of several tons of concentrate. This operation soon demonstrated that soluble sulphide—basically sodium polysulphide—when applied to a pulp made slightly acid with sulphuric acid, proved to be the most effective sulphide depressant. The rate of concentrate treatment in this plant and the low molybdenite content of the concentrate prevented carrying the process to a conclusion, namely, production of molybdenite concentrate of marketable grade. Laboratory treatment of pilot plant concentrate indicated that a marketable product could be made from it, given a tonnage suited to available cleaner flotation machine capacity. Accordingly, the molybdenite plant was enlarged to permit treatment of the total copper concentrate from the extension plant, maximum of approximately 800 tons per day. In the enlarged plant a major difficulty was encountered. No criteria could be established as a basis for control of the process. In the early stages of flotation, because of the small amount of molybdenite involved, there was no visual evidence of the relative floatability of molybdenite with respect to

Jan 1, 1956

By P. Raffinot, M. Rey, V. Formanek, G. Sitia

Six years of laboratory study, followed by three years of mill operation treating more than 10,000 tons of ore, have established the flotation of oxidized zinc ores with fatty amines as an efficient process. Recently the method has been applied with success to partially oxidized sulphide ores. In this paper the authors give an account of their research and the milling results obtained in Sardinia since 1951. CONCENTRATION of oxidized copper and lead ores by flotation has been practiced for 30 years, but flotation of oxidized zinc ores has remained unsolved until a few years ago. This problem is, however, of great importance in mining districts where large tonnages of oxidized zinc ores exist in underground mines or in dumps. Some sulphide ores also possess a degree of oxidation high enough to justify recovery of oxide zinc minerals. Various attempts have been made to float the zinc carbonate, smithsonite. It is well known&apos; that smithsonite can be collected with fatty acids, but as the predominant gangue constituents, limestone and dolomite. are floatable with the same reagents, fatty acids are difficult to use. Gaudin has shown,&apos; on mixture of pure minerals, that smithsonite can be floated with higher xanthates, or better, with higher mercaptans. Another procedure, briefly mentioned by Davis," consists in sulphidizing the ore, activating the zinc carbonate with copper sulphate, and floating with ordinary xanthates. This process was attempted in the writers&apos; laboratory in 1946, but results were erratic. It has since been taken up by Italian engineers and is in operation on certain types of ores. Two other research projects should be mentioned. In 1942 Erlenmeyer and others&apos; floated synthetic mixtures of smithsonite and gangue minerals with hydroxyquinoline, an analytical reagent for zinc. On the other hand, Bunge, Fine, and Legsdin published in 19465 Some interesting results obtained on Missouri oxidized zinc ores. With sodium oleate as a collector and a combination of sodium hydroxide, sodium silicate, and citric acid as a limestone depressor, they secured, after four cleaning operations, a good grade of concentrate. With many ores this reagent combination lacks adequate selectivity. In 1939 M. Rey initiated a program of research on this important problem. Work was pursued first at the School of Mines in Liége, Belgium"&apos; and later in Paris in the laboratories of Société Miniére et Mbtallurgique de Penarroya and Société Minerais et Métaux. Reagents used in the beginning were higher xanthates and mercaptans, later fatty amines. Sulphidization with sodium sulphide was found necessary. These conditions make possible the flotation of zinc carbonate and zinc silicate. Two milling plants were put in operation in Sardinia in 1950 and several others are now contemplated." &apos; It should be mentioned that unknown to the present writers McKenna, Lessels, and Peterson were

Jan 1, 1955

By P. L. De Bruyn

The adsorption density of dodecylammonium ions at the quartz-solution interface has been Theadsorptiondensitydetermined as a function of collector concentration and pH. A ten thoushasbeenandfold range of amine salt concentration was covered at neutral pH. Experimental results show that over a thousandfold concentration range at neutral pH, the adsorption density (I) is proportional to the square root of collector concentration. Except at high concentrations, I increases with increases with increasing pH, but in general this effect is surprisingly small. . , . . A critical pH curve has been established for the flotation of quartz with dodecylammonium acetate. The conditions along the flotation curve are correlated with the adsorption measurements. THE behavior of collectors at the mineral-solution interfaces is usually explained in terms of an ionic adsorption process. Through the distribution of collector ions between the solid surface and the- co-existing solution phase the mineral is believed to acquire a water-repellent surface coating. Quantitative adsorption studies have been made on simple flotation systems1-4 only within the last few years. Such investigations were made possible by the adoption of the radiotracer method of analysis. As a consequence of these studies a new parameter has been added to aid the understanding of the flotation process. The research investigation to be discussed in this paper was undertaken to obtain a better understanding of the behavior of a cationic-type collector. This objective was approached through the determination of the distribution of dodecylammonium acetate between the quartz-solution interface and the solution as a function of the collector salt concentration and pH. To bring this investigation to focus on the more practical aspect of flotation research, an attempt was also made to correlate the adsorption results with actual flotation tests. Quartz: A —100 mesh ground crystalline quartz was infrasized; the products of the third and fourth cones were mixed together and reserved for experimental purposes. This stock material was cleaned by leaching in boiling concentrated HC1. After leaching the quartz was rinsed with distilled water until the filtrate showed no trace of chloride ian. It was then washed several times and dried. The qwrtz had a specific surface of 1400 cma per g as deterhined by the krypton gas adsorption method. Collector: The distribution of dodecylammonium acetate between the quartz surface and the solution phase was determined by the radiotracer method of analysis with carbon 14 as the tracer element. The radioactive amine salt with C" synthesized into the hydrocarbon chain5 was supplied by Armour and Co. The tracer element was located adjacent to the polar group. The radioactive salt as received had a specific activity of about 0.14 mc per g. When desired, dilution of this activity was effected by addition of non-radioactive dodecylammonium acetate also supplied by Armour and Co. ........ All other inorganic reagents used in this research were of reagent grade. Conductivity water was used for making up all solutions. Adsorption Tests: Two different experimental methods were used. In the first, to be designated as the agitation method, a weighed amount of quartz and a measured volume of amine salt solution were agitated in a 100-ml or 50-ml glass-stoppered pyrex graduated cylinder. The cylinder was filled with solution up to the stopper, since erratic results were obtained when an air space was left over the suspension. Time of agitation varied from 1 to 2 hr. Preliminary tests at different agitation times showed that the amount adsorbed remained constant for all agitation times exceeding 1/2 hr. After this conditioning period, the solids were separated from the solution by filtration through a Buechner fritted-disk funnel. The solution was re-circulated 10 times or more to allow the fused silica disk to come to equilibrium with it. Determinations of the amount of amine adsorbed on the frit itself indicated that this amount was less than 10 pct by weight of the amine acetate abstracted by 10 g of quartz. The funnel with quartz covered by a thin layer of solution was then centrifuged for approximately 5 min, at which time the moisture content of the solids was reduced to about 5 pct by weight. The wet quartz was blown into a tared beaker, re-weighed and allowed to dry at room temperature. A final weighing was then made to determine the moisture content. The second experimental method, similar to the procedure adopted by Gaudinand Bloecher,&apos; will be referred to as the column method. Two liters of solution were passed through a bed of quartz contained in a Buechner funnel attached to a pyrex separatory funnel by means of a ball and socket joint. Preliminary tests showed that increasing the volume of solution above 2 liters does not give a measurable increase in adsorption. From 4 to 4 1/2 hr were required for 2 liters of solution to pass through the column. The moisture content of the quartz was again reduced to a minimum by centrifuging. A slightly modified column apparatus was used for experimenting with alkaline amine solutions. The same basic unit was used, but the underflow from the Buechner funnel was again fed into a Separafory

Jan 1, 1956

By J. N. Butler, R. J. Morris

A series of organic collectors has been developed which successfully float synthetic secondary uranium minerals, such as autunite, carnotite, and torbernite. Recoveries up to 97 pct have been obtained. With cleaning of rougher concentrates, ratios of concentration as high as 15:1 are possible on head assays of 0.10 to 0.50 pct U 3 O 8. Additions of certain base metal salts aid in recovery and materially increase the rate of flotation and selectivity of the collectors for the synthetic minerals. Flotation tests on natural uranium minerals have not yet equaled results obtained on the synthetics, although these collectors do give better results than conventional collectors for nonsulfide minerals. RAPID increase in mining and milling uranium ores during the past seven years and the discovery of large tonnages of uranium-bearing material too low in grade to stand the cost of leaching have developed considerable interest in possible methods for beneficiating these marginal ores. Such ores are generally classed as those containing less than 0.10 pct U3O8. Development of a beneficiation method that would give both high recovery and high ratio of concentration would also make possible the beneficiation of higher grade ores and would provide higher grades of feed to existing leaching plants. Although considerable research and testing on physical or physico-chemical beneficiation of domestic uranium ores has been done in the past, to the best knowledge of the authors only a single application of physical beneficiation is now in operation on a commercial scale. Flotation has been attempted on the basis of mineral association rather than by direct concentration of the contained uranium minerals. Flotation studies conducted on this project on certain ores gave recoveries of 60 to 70 pct of the uranium by the use of conventional flotation agents such as oleic acid, but ratios of concentration were in the order of 1.5:1 to 2.5:1. Subsequent cleaning to upgrade the concentrates rapidly increased losses in the cleaner tailings, which were not later refloated in the rougher flotation. Selection of Collector: In selecting desirable organic collectors for these oxide systems, a thorough study was made of the relationship between structural characteristics and collector activity by an examination of the behavior of a number of different functional arrangements present in a variety of organic molecules, (see Fig. 1). It is generally accepted that a collector must have a relatively long-chain hydrocarbon radical and a chemically reactive elec-trophilic center. Such a requirement could be met by a large number of compounds, but steric effects at the uranium mineral surface, solubility, compat-ability with aqueous systems, air avidity, bond strength at the mineral surface, interfacial concentration of the collector, and diffusion rate of the formed complex from the air-water interface eliminated many as unsuitable for this purpose. Preliminary investigations were completed of possible collector designs consisting of aliphatic and aromatic hydrocarbon radicals attached to stable combinations of nitrogen, oxygen, sulfur, and halogen electrophilic centers, Fig. 1.Tests were made to determine melting point spans, solubility of mineral-collector combinations in various solvents, reactions with soluble uranium salts, contact angles, and floatability by semi-quantitative beaker methods.&apos; Contact angle determinations were not too useful because of the fine particle sizes of available mineral samples. From the foregoing study it appeared that nitrogen-sulfur electrophilic centers coupled with straight-chain saturated alkyl hydrocarbon radicals would hold the most promise for floating secondary uranium minerals. As the dithiocarbamate compounds can be made to combine these attributes they were selected for further study. The dithiocarbamates form a parallel structure to the xanthates, with the fortuitous exchange of

Jan 1, 1957