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By H. E. Lee, D. Ingvoldstad

AT Bunker Hill the original charge storage and preparation system was installed in 1917 to accommodate lead-silver gravity mill products. Only minor tonnages of wet fines such as vanner and flotation concentrates were received. Charge flux and diluent requirements were provided by an ample supply of siderite middlings and coarse lime-rock. While oversize material was crushed through %-in., the use of roll crushers in series resulted in a more or less granular, free-running charge of adequate porosity. Under such conditions, receipts could be proportioned directly from receiving bins, and suitable blending was obtained by passing the resultant layered composite through a small Stedman disintegrator. This elementary system served well for production and smelting of an extremely lead-rich sinter, in the low column blast furnace. Here sinter physical quality is less critical, and appreciable quantities of raw flux and oxidized ore may be charged directly to the blast furnace without seriously impairing furnace capacity. Prior to 1938 the relative proportions of zinc receipts were low and extraneous flux requirements at a minimum. However, subsequent war-inflated zinc demands brought marked increase in fine concentrate and slimy zinc leach residue receipts and an abrupt introduction to smelting a more refractory zinciferous charge. The higher temperature demands of the more refractory zinciferous charge required a higher smelting column and the higher column, in turn, rendered the blast furnace less receptive to improperly prepared feed. Inadequate presinter processing facilities prevented the production of acceptable sinter, and blast furnace production declined to 50 pct of former capacity. In view of a prevailing contention among lead metallurgists as to the deleterious action of high zinc slag on blast furnace capacity, it was natural to overstress this factor and to concentrate on possible changes in furnace design. Another interpretation was that primary correction should be sought through expanded sintering facilities. This divergence of opinion, together with the extreme material shortages of the 1940's, delayed corrective

Jan 1, 1957

By C. S. Simons, G. Major-Marothy, M. A. K. Grice, D. A. Dahlstrom

The vacuum filter operating variables that influence cake moisture are discussed. The influence of temperature control, particularly through application of steam to the cake, is emphasized. Results of pilot plant studies on filtration of fine hematite concentrates are presented and discussed, and are shown to support the theoretically-derived conclusions. Results on fine magnetite concentrates are also used to support the argument. The relative merits of disc and drum filters from the standpoint of cake moisture are discussed. Moisture content has always been recognized as one of the most important properties of concentrates used as pellet plant feed. Most iron ore concentrates are produced by wet methods, and are finally de-watered on filters. Obviously, a real economic advantage accrues to the ability of control the moisture content of the filter cake within the range required for optimum pellet production. Another consideration, also, has been receiving increasing attention recently. Transoceanic shippers of concentrate cargoes are critically assessing the hazards of excess moisture. The nature and magnitude of these risks have been described, ' and it appears that their elimination may require that residual moisture be somewhat lower than the limit for proper pelletization. Coarse, very free-filtering materials, such as the products of spirals, are usually best handled on top-feed types of machines. The moisture content of the cake may be held at a specified value by proper design and operation, taking into account cake thickness, air-flow rate, and drainage time. Fine-grained concentrates, with which we will be concerned here, must be filtered on Drum or Agidisc machines. With many of these materials, it is possible, by proper design, to meet cake moisture requirements with a conventional filter station. The first installations to produce high-grade pellets from magnetic taconites fell in this class. In fact, the design criteria were developed in connection with operations of Reserve Mining Co. at Silver Bay, inn.,2 and these have been verified repeatedly in other mills. The trend today, however, is toward the production of more difficultly dewatered concentrates. This is due on the one hand to the increasing attention being given to non-magnetic ores, which tend to be far slimier, and therefore much more retentive of moisture, than the magnetites. On the other hand, the pressure to improve the grade of the magnetic ores is leading to finer grinds, which also are more difficult to dewater. In the latter case, the decision to improve grade in an existing operation by finer grinding for better liberation may lead to two unattractive alternatives: 1) substantially higher bentonite consumption, plus the risk of poorer quality pellets, from high cake moisture, or 2) installation of a thermal dryer. There is, however, a third alternative. In recent papers by two of the authors3,4 it was shown that for many materials a significant decrease in cake moisture can be obtained by applying live steam to the cake face during the drying part of the filter cycle. Further, this limited drying appears to have distinct economic advantages compared to thermal drying. It is the purpose of this paper to explore the third alternative. To do this, the results of an extensive pilot study of steam filtration of a hematite flotation concentrate will be critically examined. This study was sufficiently broad so that a number of alternate suggestions (besides steaming the cake) for reducing moisture content were tested, and comparative data for both Drum and Disc filters were obtained. The work was particularly interesting since the concentrate has a high Blaine surface, and is therefore particularly difficult to dewater to the levels required. FILTER CAKE MOISTURE The significant difference between concentrate filter cake as discharged from the filter, a cargo of that same filter cake during or after shipment, and a green ball formed from that filter cake is in the relative volume of voids, or porosity. In every case, the material is a collection of small solid particles held together by the cohesive forces of a liquid. The physical properties of the aggregate are determined by the

Jan 1, 1967

By R. Adair, W. R. Hudspeth, W. T. McDaniel

D. C. Ralston (U. S. Bureau of Mines, Washington, D. C.)—Flake mica can be beneficiated by a wider variety of methods than almost any known mineral. However, most of these methods are not recorded. It is therefore a pleasure to find a written record of another application of a new method of beneficiation. The fact that the Humphries spiral gives a concentrate containing the mica flake mixed with the slime and small grit has forced a further separation by screening, in which the finest of the flake will pass through. Fine flake has its uses and if mixed with slimes or clay is recoverable reliably only by flotation. This loss must be charged against the spiral. On the other hand, the new muscovite mica paper recently brought out by General Electric Co., can be made only with grit free flake larger than 50 mesh. It is of interest to learn from the authors whether the washing of the flake on a 50 mesh screen can take the grit content below 1 pct, to meet this new demand. Must another degritting process be used on the flake, like electrostatic treatment of the dried material? H. D. Snedden (The Humphreys Investment Co., Denver)—When recovering mica in the spiral, gangue particles coarser than 16 mesh can be rejected more readily when operating at a high pulp density (35 pct solids) or a low volume rate (—15 gpm). A launder screen as used by Lehigh Coal Co. has proved very satisfactory to partially dewater the mica concentrate and to reject any gangue slimes it may contain. Under certain conditions mica cleaning in the spiral can be improved by reagentizing the spiral feed, the object being to insure wetting of the gangue minerals and to agglomerate or prevent wetting of the mica particles. R. B. Adair (authors' reply)—The grit content of the mica recovered on Humphreys Spirals varies with the fineness of the launder screen used to clean the mica product. Normal practice has been the use of 80 mesh screen, which will produce a mica product of acceptable grit content for most purposes. By using a 50 mesh screen, and increasing the screen area per ton of feed, it is felt by the authors that a mica of 99 pct plus purity could be obtained. This statement would hold only in the case of mica feed which was well de-laminated, and free from "clay balls," (semikaclinized feldspar), which tends to follow the mica through the process.

Jan 1, 1952

By P. Somasundaran, D. W. Fuerstenau

The length of the collector is found to influence the flotation of the mineral even at incipient conditions, which are below the concentration at which interaction at the solid-liquid interface begins to take place to form hemi-micelles. To study this dependence, concentration for incipient flotation of quartz was determined as a function of pH with collectors of various chain lengths. The observed effect of chain length on flotation is ascribed to that of collector adsorbed on the bubble surface. In previous studies, it was shown that at low concentrations the alkyl collector ions adsorb at the solid-liquid interface as individuals.''2 At higher concentrations, the collector ions adsorbed at the solid-liquid interface associate with each other to form two-dimensional aggregates called hemi-micelles. Above the hemi-micelle concentration, the length of the hydrocarbon chain is extremely important since the hydrocarbons are in effect removed from water during the association, making the energetic conditions more favorable for adsorption at the interface. Because of this enhanced adsorption, one observes a very rapid increase in flotation associated with the hemi-micelle formation at the solid-liquid interface. However, a dependence of flotation on the chain length at concentrations below that required for hemi-micelle association was also observed,' and this cannot be explained by the above mechanism which postulated hydrocarbon chain interactions only at the solid-liquid interface. This prompted an investigation into other possible reactions of the hydrocarbon chains and an examination of the conditions at the bubble surface involved in the flotation system and how these observations might explain the reactions at the solid-gas interface which cause the particle-bubble attachment required for flotation. To obtain more information on chain length effects, flotation, under incipient conditions, was tested by vacuum flotation techniques. The collector-concen-tration-pH relationships for flotation of quartz with alkyl ammonium acetate collectors was delineated by observing the pH at which quartz particles begin to float to the liquid surface. By investigating flotation as a function of pH, it was also possible to study the effect of neutral molecules on incipient flotation conditions, since the aminium ions hydrolyze to amine molecules at higher pH values. EXPERIMENTAL WORK Brazilian quartz specimens were crushed and sized, and the 270 x 400 mesh fraction was used for flotation studies. The samples were leached with concentrated hydrochloric acid until no coloration of the acid occurred. The leached material was washed free of chloride ions and stored in distilled water. The vacuum flotation technique developed by Schuhmann and prakash3 was used to determine the critical pH-concentration curves. This method, which can be used to delineate conditions for incipient flotation, is fairly simple and rapid. About 0.5 gm of 270 x 400 mesh quartz was placed in a 100 ml graduated cylinder which was then filled to the 100 ml mark with the collector solution made from high-purity alkyl ammonium acetate salts. The water used for the test was conductivity water saturated with air that had been passed through a cleansing train consisting of Drierite, Ascarite, a water wash bottle, and a trap. After the pH was adjusted, the cylinder was then conditioned for thirty minutes. In the tests where an acid pH was desired, sufficient acid was added before the collector solution to avoid any effect due to slow desorption of collector from the quartz surface. After conditioning, vacuum was applied to the system and the flotation or nonflotation of the quartz was noted. The pH at which the quartz particles began to float to the liquid-gas interface was taken as the critical PH. Critical pH curves were thus determined for different concentrations of the various collectors. Hallimond tube flotation data were taken from the authors' previous publication1 for correlation with that from the vacuum flotation. RESULTS AND DISCUSSION The results of vacuum flotation studies for determining critical pH-concentration curves, i.e., curves which delineate conditions for incipient flotation, are shown in Fig. I. In generaI, all the curves exhibit an upper and lower pH limit between which flotation will

Jan 1, 1969

By O. A. E. Jackson

Pebble milling has been practiced in the reduction works of South Africa gold mines for well over 50 years. Originally flint pebbles were imported from Denmark to grind stamp-mill amalgamation-process tailing, which contained a good deal of extractable gold, but local operators soon found that large pieces of ore could be used for the same purpose. The ore is a hard, tough conglomerate in which quartz pebbles are cemented together by a matrix of redeposited silica interspersed with pyrite crystals. The gold, rarely visible, occurs as fine particles mostly segregated at the interface of the pebble and matrix, although a small fraction occurs within the pyrite crystals. There is seldom any gold in the pebbles themselves. Following the usual South African practice in pebble milling, Union Corp. grinds the ore wet in two or three milling stages incorporating classification. The sized broken ore used as grinding media is separated from the main ore stream in the crushing section that prepares the ore for milling. Where the ore channel, or reef, is narrow there is a shortage of large pebbles. In this case primary grinding may be done in ball mills or, more recently, in rod mills, which cost less per ton to operate. The trend, however, is to prepare finer feed for the milling section. This makes it possible to use smaller primary pebbles and eliminates the need for steel. REDUCTION WORKS OF UNION CORP. LTD. Union Corp. Ltd. exercises financial and technical control over a group of seven gold mines in the Transvaal and Orange Free State. In the Transvaal, with one exception. the mines lie 20 to 40 miles east of Johannesburg, in flat or gently rolling countryside. Winkelhaak, the first of several new mines that will be developed by Union Corp., is located in similar terrain in an entirely new gold mining district about 80 miles east of Johannesburg. Table I gives details of milling units for six of the Union Corp. mines, together with the tonnage milled in 1957. Winkelhaak Mines Ltd. is not included, as it did not begin milling until 1958. This reduction plant has no crushing section; ore is ground directly from the mine (autogenously) in 12x16-ft mills. Because these operations are still in development, they are not described in this article. It will be noted that certain reduction works have mills of more than one size in the same milling stage. This came about when plant extensions in- corporated larger units. In the case of Geduld Propty. Mines Ltd., which began milling operations 50 years ago, the primary stage is stamp milling. The reduction works follow a uniform pattern and are usually joined to the main ore shaft. Ore from other shafts is brought by standard-gage railway and dumped into a common transfer bin. The trend is to increase surface storage capacity to enable the crushing and milling sections to operate at a steady rate, independent of fluctuating ore deliveries from mine. Milling and cyanide extraction divisions of new mines are always designed to allow for extensions as mine production increases. The conveying, washing, and screening system of the crushing section is usually laid out in final form, with additional space for more crushing equipment. The crushing sections operate on one shift during early years of mine production; a second shift is introduced when the mining rate warrants it. Ample surge capacity is provided. Crushing and milling is done only on weekdays, as the law does not allow these operations to take place on Sunday in any plants constructed since 1911. The cyanide extraction sections, however, operate continuously seven days a week, drawing on mill pulp gradually built up in the thickeners during the week. Construction and equipment of milling plants follow standard practice. Dilution water is drawn from a large, high-level tank to obtain constant pressure, but gland service water for the pulp pumps is reticulated from high-pressure, two-stage pumps. The mills are equipped with the most up-to-date machinery and are designed to save labor. They compare favorably with milling plants in countries where native labor does not exist, and automatic controls are being installed wherever feasible. Hydrocyclone classifiers have replaced mechanical classifiers in modern milling plants, chiefly because of the saving in capital outlay, maintenance, and building space. The hydroclones are fed from steady head boxes rather than directly from pumps, and dilution water is introduced into these boxes. Tests have shown that in steadiness of operation and separating efficiency cyclones are comparable to mechanical classifiers. but protective stationarv screens are needed to keep the spigots clear. Rubber-lined pumps are used for pulp of about 3 mesh or finer and metal-lined pumps for coarser material. None of the Union Corp. milling plants practices gravity concentration of coarse gold by amalgamation or the use of corduroy blankets. Studies have proved that no economic case can be made for these methods, which complicate the milling process and demand extra precautions against theft.

Jan 1, 1960

By J. W. Snavely

THE high tension belt conveyor is introducing a new and tremendously expanded era of low cost bulk material handling. High tension belt conveyors are generally those installations involving very long centers, high lifts, or drops, in which the belts are stressed up to their maximum tension values, and further, where the belt construction provides tension capacity far beyond what is possible with conventional belt constructions. With these high tension installations, the magnitude of the forces involved demands careful refinement of accepted design practice in order to achieve optimum balance of all factors. No attempt will be made to evaluate the relative merits of belt conveyor haulage with other means of transportation. For present purposes, it is assumed this has already been done in favor of belt conveyor. Neither will any attempt be made to evaluate the various conveyor belt constructions now available or to balance the advantages of various types of mechanical equipment. It is also assumed that the basic haulage information on which the conveyor design is based is accurate and complete. A sustained maximum, uniform load on the belt at all times must be achieved through proper feed control and the use of adequate surge storage to level the peaks and valleys of any varying demand for the material being handled. General Belt Capacity Considerations The belt conveyor capacity tables published by various belting and conveyor equipment manufacturers vary to a considerable degree, and the ratings given are quite conservative. Of necessity, these published ratings are based on the handling of average materials under average conditions. In applying a high tension belt, all possible capacity from the belt must be obtained in order to hold its width to a minimum and thereby limit the initial cost. Two factors are involved, loading to maximum cross section area and traveling at a maximum practical speed. Belt Loading: Proper treatment of the loading of the belt will result in maximum cross section to the load, and published capacity ratings can be exceeded, sometimes by appreciable margins. On the 10-mile conveyor haul used in the construction of Shasta Dam, California, although the rated capacity of the belt line was 1100 tons per hr, at times the system handled peak loads of 1400 tons per hr, almost 25 pct better than the rated capacity. One of the large coal companies has been able to exceed rated capacity by as much as 50 pct. Loading conditions which must be controlled are: 1. Large lumps must be scalped off and rejected or the load must be primary crushed before being placed on the belt. 2. The material weight per cubic foot must be accurate, must be known for all the materials being handled, and must be known for the complete range of conditions of the individual material being handled. Long centers and high lifts magnify small differences into serious proportions. 3. Uniform feeding to the belt is most important. Various types of feeders are available, which can be used to place a constant predetermined volume of material on the belt, or, where an appreciable range of material weight exists, through electrical control actuated by current demand, to place a predetermined uniform tonnage on the belt. One long slope belt in a coal mine in Pennsylvania is being fed at three separate stations with the controls so arranged that whenever the maximum load is going onto the belt from the first station, the other two stations automatically cut out. Whenever the load from the first station drops back, the other two stations again automatically cut in. 4. Careful design of the chutes and skirts is necessary to get the load centered on the belt with a minimum of free margin along each edge. Some free margin at the edge of the belt is necessary to prevent spillage, but if the load can be kept accurately centered, this free margin area can be reduced, and more material can be carried on the belt. What can be accomplished in this respect will vary widely, depending on the nature of the material being hauled. The chute and skirt design must also protect the belt. 5. The design of chutes and skirts should also get the load traveling in the same direction and close to belt speed, so that the load comes to rest on the belt as quickly as possible. The design of the chutes and skirts is worthy of careful study, and after a system is put into operation it should be experimented with to get the best results. Belt Speed: High belt speeds should be used in high tension work. Obviously, high belt speeds enable haulage on a narrower belt, reducing initial cost. The major portion of belt wear takes place at the loading point and around the terminal pulleys. The

Jan 1, 1952

By B. C. Mariacher

EXTRACTION of uranium from ores is being ac-complished by processes which. for the most part, subject the entire ore to acid or carbonate leaching. Ore deposits with a U 3 O 8 content below 0.10 pct U 3 O 8 are seldom considered suitable for treatment by leaching. A preliminary concentration that would enrich the uranium content of an ore by a simple, low cost process based on physical properties of the ore might result in some low grade deposits becoming commercial ores. In addition, the process might be employed in existing operations to reduce transportation and leaching costs and to increase capacity of existing leaching plants. A study to attempt the development of a preliminary concentration process for primary uranium ores was undertaken by the Colorado School of Mines Research Foundation under sponsorship of the U.S. Atomic Energy Commission. The objective of this work was to produce concentrates containing 0.25 pct U3O8 from the low grade ores tested. Ores Tested: The main effort was devoted to the low grade primary uranium ores from northwestern Saskatchewan. Samples were obtained from the Beaverlodge operation of the Eldorado Mining & Refining Ltd. Additional primary ores, obtained from deposits in Gilpin County, Colo., contained from 0.07 to 0.10 pct U3O8. Summary of Concentration Tests: The Beaverlodge ore was tested to determine amenability of the ore to concentration by magnetic, electrostatic, gravity, and scrubbing processes. None of these produced satisfactory results. Both gravity and magnetic processes produced fairly good concentrates when closely sized fractions of the ore were treated, but on the basis of treating the total ore, recovery was poor. Preparation of sized fractions and the low capacity of equipment for suitable concentration made these methods impractical. As flotation offered the advantage of treating the total ore without intermediate sizing, the main effort was in this direction. A flotation process was developed that fulfilled the concentration objectives as set by the AEC. Pilot plant testing was used to verify results obtained from laboratory batch testing. Mineralogy: A petrographic examination of the Beaverlodge ore included a study of polished sur- faces and identification of the radioactive mineral by autoradiograph and X-ray diffraction. Approximate quantitative mineral identification was as follows: quartz, 60 pct; orthoclase feldspar, 20 pct; chlorite, 10 pct; carbonates, 5 pct; and miscellaneous minerals, 5 pct. Included in this last group were plagioclase feldspar, pyrite, mica, chalcopyrite, pyroxene, sericite, magnetite, galena, and uraninite. The most general occurrence of uraninite was in the form of crusts and thin coatings on limonite-stained grains of pyrite, quartz, and pyrite-quartz intergrowth. At least 90 pct of the uraninite was still attached to other minerals in a 100 by 200-mesh size fraction. The uraninite crusts were as small as 10 to 20 µ diam, and 5 to 10 µ thick. The Flotation Process Petrographic examinations of the Beaverlodge ore had indicated the impracticability of attempting to concentrate the uranium by floating individual grains of uraninite. Liberation of the uraninite required grinding to sizes below those suitable for flotation. However, there was preferential association of the uraninite with some minerals while others were free of uraninite attachment. The approach to the development of a flotation process was, therefore, based upon an attempt to concentrate the uraninite by floating carrier minerals. The following paragraphs discuss the various stages of the process with regard to the factors tested and the conditions under which best results were obtained. Grinding: The most effective size range for flotation was —150 mesh + 13 µ. The —13 µ material in the final concentrate had a higher U3O8 content than the total ore, but not as high as the average concentrate; however, rejection of slimes before flotation was prohibitive because of the loss in uranium carried in the —13 µ fraction. Grinding techniques which contributed to a minimum production of fines, such as stage grinding, were then employed. Quartz and Silicate Depression: These minerals represented approximately 80 pct of the ore and were free to a large degree of uraninite attachment. Significant improvement in the grade of the concentrate was obtained by depression of these minerals with hydrofluoric acid or sodium fluoride. Promoter: Selective stage flotation of uraninite carrier minerals was simplified by development of a single promoter mixture. The mixture consisted of an emulsion of a fatty acid, fuel oil, and a petroleum sulfonate and was selected after a comprehensive series of tests. It contained three parts by weight of an oleic and linoleic acid such as Emersol 300,

Jan 1, 1957

By C. W. Hoffman, W. R. Hinken

Probability sizing, a fairly recent development in the field of particle separation, is now under investigation to determine its value to the mining industry. The method employed is probability sizing through a vibrating network of decreasing screen openings. While probability sizing is not new, the possibility of using it for coarse, highly abrasive ores at high tonnage rates has only recently appeared to merit practical consideration. The advantages of probability sizing vs. conventional screening methods are: (1) reduced space requirements for a high capacity unit, (2) fewer moving parts, (3) more versatile product sizing selection, (4) reduced maintenance requirements, and (5) lower power consumption. Screening is a necessary part of virtually every crushing circuit. At some stage in the process, particle separation becomes a controlling factor in the operation. The basic type of screen in use today is the vibrating unit, which usually employs a mechanical vibrator of one type or another to obtain sizing and material transportation across the screen surface. The vibrating screen has its limitations and inherent problems. Operators must live with: (1) high maintenance — worn rods, screen cloth, spacers, cross members, chutes, etc., (2) low capacity in relation to floor space requirements, (3) extreme noise levels, (4) high power consumption, and (5) structural problems in the screens and supporting steelwork. All plant operators have become tolerant of these drawbacks through the years, and little innovation has been developed in screening. The vibrating screens in use today incorporate improved metals, more efficient power sources, better lubricants, and increased manufacturing skill, but the basic machine itself remains unchanged. Screening is a relatively small part of most plants' operating expenditures and gets relatively little attention from research and development. The HiProb sizer is a new approach to the problem of particle separation on a high tonnage basis. Developed by Dr. Fredrik Mogensen (president of Fred-rik Mogensen, A. B. of Djursholn, Sweden), this unit has found its primary acceptance in the sizing of wet or dry materials in the fine ranges. Climax has been working cooperatively with Mine and Smelter Supply Company, the United States licensee, on the adaptation of the sizer to obtain consistent recovery at a high tonnage rate on a — 3/8-in. separation. This was a new application for the sizer and opened up a whole new field of problems. Dr. Mogensen's machine utilizes the control of normal probability of falling particles and is designed to eliminate some of the undesirable features of vibrating screens mentioned earlier. The machine is light, compact, consumes less power than conventional screens and has a low noise level. PRINCIPLES Probability sizing is a process designed to allow solid particles to segregate with a high ratio of the materials reporting into their respective size fractions with a minimum amount of misplaced particles. Sizing is defined in this paper as segregation of dry solid particles into fractions of comparable sizes. Probability is usually estimated as the ratio of the number of ways an event might occur vs. the total number of allowable ways. The probability process described depends upon the fact that particles falling through a network will segregate across the sizing surface even though all openings are larger than the particles. The degree of segregation is dependent upon the particle size and the trajectory applied by the sizing network to each falling particle. Specifically, the process utilizes a maze made up of screen cloth to control the path of a particle in free fall as illustrated in Fig. 1. The usefulness of this probability sizing process is dependent upon the degree of efficiency of an effective maze. Primarily, a maze must be extremely flexible to meet the multi-faceted nature of solid particles and, at the same time, the degree of control must be reproducible so that a desired separation is predictable. Dr. Fredrik Mogensen has formulated a maze that is not only flexible and easily controlled, but, mote importantly, easily adaptable to a working machine.

Jan 1, 1970

By L. W. Pommier, F. B. Brien

In order to determine the ability of cyclones with heavy liquids to preconcentrate ores, a study of the variables which govern the behavior of the solids in the apparatus has been made. To assess the test results, the cyclone products have been examined by sink and float analysis at various specific gravities and the results examined by the Tromp type curves. This analysis has allowed the development of the SG50 unit — specific gravity of the particles, 50% of which goes to overflow and 50% to underflow — rather than the d50 (size) unit used in classification. The results of tests using the closed circuit laboratory cyclone assembly with dry feed samples show the validity of the equation: Results indicate that 83% of the feed may be rejected containing 0.1% Sn. Recoveries in excess of 80% are obtained at a grade of 1.7% Sn from feed samples assaying 0.4% Sn. As water-wetted particles may comprise the plant feed, separate tests were performed to examine the effect of water in the feed material. Results indicate that the entrained water divides into two fractions, one of which is defined as free water and is rejected to the cyclone overflow, the other fraction forms an adhering envelope of water on the ore particles, altering their effective size and density. The presence of the envelope of water on the particles increases the ease with which the heavy liquid can be recovered but decreases the concentration efficiency. It is proposed that the detrimental effect of water in the cyclone may be minimized by using a higher feed pressure cx by the use of surfactants. A higher feed pressure, by increasing shearing forces, would reduce the thickness of the water envelope on the particles. The surfactants, by adsorbing on the particles and producing hydrophobic surfaces, would displace the water envelope and allow the heavy liquids to wet the modified surfaces. In both cases a larger proportion of the entrained water would report to the cyclone overflow as free water, thus minimizing the detrimental effect of water on the concentration efficiency. Additional testing is required on the behavior of water in the system; to determine the optimum particle size for the preconcentration step; to determine final particle size for optimum liberation and subsequent treatment of the products in present plants and to determine the efficiency with which the heavy liquids may be recovered from the products. For many years tin mining in Bolivia has been very important to the economy of the country. In recent years the tin content of the ore reserves has been steadily declining. An improvement in concentration efficiency is highly desirable in order to increase productivity from the present low-grade mine ores and to allow retreatment of the many millions of tons of tailings from past operations, some of which equal the present ores in tin content. A number of investigators have been or at present are working on various phases or approaches to the problem. Gravity concentration is used in Boliva today as it has been in the past to recover tin minerals from the mine ores. An objective of the work reported in this paper was to examine the possibility of treating present low-grade ores and sink-float tailings products, at a relatively coarse grind, in order to produce a higher grade feed material to the present gravity concentration plants. The possibility of using heavy liquids in cyclones to treat large tonnages with a minimum of outlay presented itself and is of particular interest since heavy liquids are now available with physical properties which lend themselves to simple recovery methods. Tests were carried out using a closed circuit laboratory cyclone assembly. It was desirable to determine the value of using the SG50 concept and Tromp curves for analyzing the cyclone concentration test results. The effects of entrained water on the behavior of ore particles in the cyclone were examined. THEORY AND TECHNICAL CONSIDERATIONS Three types of heavy liquids are available: aqueous solutions of very soluble salts, salts or mixtures of

Jan 1, 1964

By D. S. Flett, A. W. Fletcher

Equilibrium studies on the extraction of nickel and cobalt with kerosine solutions of naphthenic acid have shown that an exchange extraction reaction occurs at pH 5.5. The nickel/cobalt separation factor is constant at 1.8 for constant total metal molarity and varying nickel/cobalt ratios. The separation factor decreases with increasing total metal molarity in the organic phase beyond 0.2 M and also decreases with increasing temperature. From the equilibrium data, it has been possible to derive a mathematical model for the separation of nickel from cobalt by exchange extraction in multistage systems. Experimental data from a continuously operated multistage mixer/settler apparatus has shown a reasonable correspondence with computer-calculated data. The effective separation of nickel and cobalt in sulfate solution remains a problem in hydrometallurgy and the hope that this would be solved by solvent extraction has not yet been fulfilled. With chloride solutions, advantage may be taken of the ability of cobalt to form anionic complexes with the chloride ion. It can then be readily separated from nickel, which does not form stable chloro complexes, by extraction with a suitable long chain amine. However, in hydro metallurgical operations, sulfate solutions are generally obtained in which no extractable anionic species are present. Thus, the possibility of using cationic extractants must be considered, and in this paper attention is directed to the use of carboxylic acids. The method of separation studied has been termed exchange extraction, which involves replacement of a metal in the organic phase with a more acidic metal in the aqueous phase. Thus, (BR2).+ (Az+)aqt=(AR,).+ (B2+)aq (I) where metal A is more acidic than metal B, R represents the acidic radical derived from the acid RH, and the subscripts e and aq refer to the organic and aqueous phases, respectively. Ashbrook and Ritcey' have used this method for the separation of cobalt from nickel using the sodium salt of di-2 ethyl hexyl phosphoric acid, which preferentially extracts cobalt. Some nickel is coextracted, and this is removed by exchange with cobalt ions in the feed solution by suitable countercurrent operation in a pulsed column. Much work has been carried out by a number of workers in Russia on the general use of exchange extraction for the separation of metal ions using car-boxylic acids. Gindin et aL a have demonstrated that this technique could be applied to the separation of nickel from cobalt using a C--C. carboxylic acid and have applied the technique to the production of high purity cobalt solutions for electrolysis. Further worka was concerned with the development of a process for the separation of nickel from cobalt in a pulsed column. This system permitted the separation of iron and copper from nickel and cobalt in one system. The procedure involved center feeding with acid backwashing at the top and alkali addition lower down the column. Thus the system operated under a pH gradient and the metals were distributed in the column in the order of their basicities. A similar application was studied by Gel'perin et al,4,5 for the removal of copper and iron impurities from a nickel anolyte by means of a C10-C,12 fatty acid fraction. Ginden et al,' and Fletcher and Wilson' have studied the effect of pH on the extraction of a number of metals with carboxylic acids. These studies showed that metals such as iron, copper, lead, zinc, nickel cobalt, and manganese are extracted at pH values close to the pH of hydroxide precipitation. Nickel is extracted at a slightly lower pH than cobalt and thus the nickel/cobalt separation factor has a value not much greater than 1. More basic work on complex identification has been reported by Fletcher and Flett: by Tanaka; and Jay-cock and Jones." These studies have suggested that at low loadings in the organic phase, the nickel and cobalt carboxylates appear to be dimeric and solvated by free carboxylic acid molecules. As the concentration of metal in the organic phase increases, the complex changes and larger polymeric species are formed. In order to permit assessment of the potential of carboxylic acids as extraction reagents for separation of

Jan 1, 1971

By G. A. Stamboltzis, N. Arbiter, C. C. Harris

Fracture under low-velocity free-fall and double impact and under slow compression have been investigated. The pattern of breakage and the size distribution of resulting fragments of sand-cement and glass spheres have been determined. Photoelasticity methods were used to simulate the stress distributions in free-fall impact in order to explain the observed patterns of breakage. Oblique fracture planes, occurring only in free-fall impact, develop along lines which coincide with the trajectories of maximum compression as determined by mathematical analysis. Breakage efficiencies for different modes of fracture were compared for both types of spheres. For the same specimen and loading system, static loading and low-velocity dynamic loading induce geometrically similar stress fields resulting in reasonably similar fracture patterns and shapes of fragments. This work had its origin in a desire to understand three fundamental processes involved in autogenous comminution: namely; free-fall impact, double impact, and slow compression. To simplify the investigation, it is preferable to reduce the entire multi-stage fracture process to its most elementary form; that is, to study products resulting from a single-stage operation, such as the breakage of a specimen under single fracture conditions. The goal of this research is to study the energy utilization in the single fracture of brittle specimens and to relate the pattern of breakage and the resultant fragment size distribution with the nature of the material, the specimen size, the manner of load application, and the rate of loading. Published works on single fracture have mainly been concerned with tests on large (> 1-in.) irregular mineral pieces, especially coal and coke. These were mainly friability tests employing a single blow on closely sized pieces. More recently, small mineral specimens in the sieve range and glass spheres have been investigated.5-7 Kick8 has reported free-fall and double impact tests with large large cast-iron, cement, and clay spheres. The establishment of a minimum breaking height independent of size in the free-fall impact tests, and the direct proportionality between the minimum work required for fracture and the volume of the specimen in double impact tests were used by Kick as an experimental proof of his classical law of "Proportional Resistances." In the present work, spherical shapes were chosen because of their simple body geometry and consequent impact and stress field symmetry. Because the study involves several physical principles in connection with brittle-fracture, it may be of interest in fields where the strength of materials is of importance. APPARATUS The equipment for the free-fall impact testing is illustrated in Figs. 1 and 2. This consisted of a massive (17 x 17 x 3-in.) hard steel plate onto which specimens were dropped and a specimen release mechanism which could be set at any desired height up to 10 ft above the plate so that the impact velocity could be varied up to about 25 ft per sec. Double impact breakage is characterized by two points of loading situated at opposite poles. For this mode of breakage the apparatus was modified (Figs. 3, 4 and 5) so that a falling mass provided an impact of predetermined magnitude on a specimen resting on the plate. A spring-loaded device operated immediately after fracture to arrest the falling mass and to record its residual kinetic energy. Precautions were taken to avoid secondary breakage. Slow compression tests were performed on a conventional hydraulic testing machine, the specimen being held between two hardened parallel bearing blocks. In the case of glass spheres, carbide bits were used. Photoelasticity studies were performed on two-dimensional models held in a loading frame equipped with a force gauge (Fig. 6). The models were viewed in a polariscope and isochromatic fringe patterns were recorded photographically. SPECIMENS Sand-Cement Spheres: A series of sand-cement spheres were prepared by molding in spherical glass flasks.

Jan 1, 1970

By R. B. Walker, R. Palmer, L. C. Bryner

The purpose of this investigation was to further study some of the factors affecting the air oxidation of sulfide minerals. Data obtained from laboratory studies on the effect of temperature, concentration of iron salts, pH, and other catalysts on the biological and nonbiological oxidation of sulfide minerals are presented. The oxidation of copper sulfide minerals and subsequent leaching of the soluble copper has long been a profitable operation. The chief oxidizing agent is oxygen from the air. This reaction is influenced by a number of different factors such as temperature, light, hydrogen ion concentration, oxidizing agents, and various catalysts. Sullivan1 found that copper sulfide minerals are slowly oxidized by the air and that dilute sulfuric acid and iron (111) increased their solubility. At that time it was not known that bacteria were involved in the oxidation of the sulfide minerals. It has been shown that pyrite, molybdenite, free sulfur, and the copper sulfide minerals found in the waste rock dumps in Bingham Canyon, Utah, are largely oxidized by chemosynthetic autotrophic microorganisms found in the leaching streams.2-6 Similar organisms have been isolated in other mining areas7,8 and from the effluents from coal rnines.9 The following are some of the questions that have arisen regarding the mechanisms of the reactions involved in the leaching process. How dependent is the leaching operation upon the biological reactions? How does the temperature effect the biological and non-biological reactions? How does the concentration of iron (II) and (111) effect the rate of oxidation of copper (II) sulfide? Does the iron act as a catalyst? Does activated carbon act as a catalyst on the oxidation of sulfide minerals? What effect does the hydrogen ion concentration (pH) have on the above reactions? The purpose of this paper is to report the results of laboratory studies on these problems. EXPERIMENTAL The apparatus used in this investigation was an air lift percolator described in our earlier work.' The percolator consisted essentially of an air lift which enabled the leaching solution and microorganisms to pass freely through a suspension of mineral (substrate) in Ottawa sand. The air lift served the additional purpose of maintaining an ample concentration of oxygen and carbon dioxide in the nutrient (leaching) solution. (See Fig. 1.) The percolators were so constructed that they could be immersed in liquid constant temperature baths for temperature studies. The thermostats were controlled to +0.5°C. The reactions were all run in the dark. The percolators were charged with one to three g of mineral, dispersed in 100 g of Ottawa sand and 100 ml of nutrient or leaching solution. The sand was purified to remove traces of iron by digesting with concentrated hydrochloric acid and washng with distilled water until free from chloride. After drying, the sand was ready for use. The compressed air supply was regulated to 5 psi and bubbled through an aerator containing distilled water, which was maintained at the same temperature as the bath used. The air was filtered through cotton plugs to maintain sterile conditions. The tops of the

Jan 1, 1968

By S. C. Sun, T. Chao

The reaction of gaseous sulfur trioxide with different aluminum-bearing minerals for the conversion of their alumina content into water soluble sulfates has been investigated. Theoretical considerations indicate that this reaction is thermodynamically feasible and that preferential sulfatization can be accomplished. The five minerals tested were gibbsite, boehmite, diaspore, kaolinite, and halloysite. In general, over 90% sulfatization could be achieved in 1 hr when using sodium sulfate as an additive at a temperature between 500° and 600°C. It was found that sulfates of the light alkali and alkaline earth metals exhibited some effectiveness in promoting the sulfa-tizing reaction. Selective sulfatization of alumina in a mixture of calcined halloysite and ferric oxide was accomplished. The gas-solid type of sulfatizing reactions have become increasingly important in recent years for the treatment of the oxides of such nonferrous metals as cobalt,' nickel,' manganese,3 zinc, and copper.4-6 Similar reactions were attempted in the present investigation for the recovery of aluminum from the low grade aluminous ores. Unlike the Bayer process in which silica is a serious contaminant, or the acid processes whereby iron usually causes difficulties, the reactions between gaseous sulfur trioxide and various aluminous minerals could hopefully be selective for aluminum. THEORETICAL CONSIDERATIONS Reactions between aluminum-bearing minerals and gaseous sulfur trioxide have been shown respectively in Table I and Fig. 1 to be thermodynamically favorable for temperatures below 1000°K approximately. Calculations of the thermodynamic functions were where ?G, ?H, ?Cp and ?S are changes in free energy, enthalpy, constant-pressure specific heat, and entropy of the system under consideration, respectively; T is the absolute temperature, R the universal gas constant, and K the equilibrium constant. Thermochemical data so far available for these calculations were based on sources given by T. Chao.7 In Table I, the free energy values indicate that reactions between the gaseous sulfur trioxide and

Jan 1, 1968

By N. M. Levine, W. M. Fassell

In this laboratory study the problem of aerative conditioning to separate chalcopyrite and pyrite from cobaltite was simply effected with a sulfy-drate collector and pH by proper choice of mixing variables. It was shown that for a given impeller type and geometry, selectivity is governed by a relation between the Froude number and a modified power number multiplied by impeller tank-diameter ratio wherein the power factor is expressed as power intensity or horsepower per unit volume of contained pulp. In applying this approach to other systems, other criteria would be used in place of selectivity. The purpose of this paper is to show how solutions to problems of aerative conditioning of pulps may be enhanced by development of mixing parameters in a given system through application of a fundamental approach to agitation. The term aerative conditioning refers to a combination of processes which include mechanical gas dispersion, gas solution in the liquid phase, transport of dissolved gases and other reactants to the reaction zone, physico-chemico interaction within the liquid phase and/or at the solid surface, and, finally transport of products away from the reaction zone. Important elements of these processes are: 1) Air-Liquid Contact Area: This area limits the rate at which a given gas may dissolve in a given liquid. Thus, an aerative conditioning system should be designed for optimum gas dispersion. 2) Diffusional Factors: Such factors as concentration gradient, diffusion film thickness, and diffusivity may govern the rate at which dissolved constituents are transferred to the zone of interaction at a particle surface and the rate at which reaction products are transported therefrom. Thus, such systems would be designed for maximum turbulence to effect good mixing and to minimize diffusion film thickness. 3) Solids-Turbulent Liquid Contact Time: Sufficient time should be provided for interactions to take place at solid-liquid interfaces. The region of maximum turbulence for mechanical systems being in the impeller zone, it is important that the number of passes through this zone be optimized. This means that a proper balance should be struck be- tween flow and turbulent power to achieve optimum conditioning time. Too high a flow rate and insufficient turbulence may be improper if the reactions are diffusion limited. On the other hand, too much turbulent power and too little time in the reaction zone due to insufficient flow power may also be limiting. 4) Geometry and Impeller Design: The design of the conditioning vessel and the type of impeller employed are important factors in determining the relative distribution between flow and turbulent power and the degree of gas dispersion. Maximization of power input would be to no avail if the distribution between flow and turbulent power were improper or gas dispersion insufficient. 5. Physical Chemistry: The reactions between dissolved gases and solid surfaces may be limiting if gas solubility, activation energies and/or free energies are not favorable. J. H. Rushton, E. W. Costich, and H. J. Everett' have summarized the mechanical elements of the combination of processes listed in a fundamental mixing equation. This equation was derived by dimensional analysis and relates the physical variables for a single impeller, centered in a cylindrical, vertical axis, flat-bottomed tank. It is, where T is tank diameter in feet; H is liquid depth in feet; C is height of impeller off tank bottom in feet; S is pitch of impeller; L is length of impeller blades; W is width of impeller blades; R is number of baffles; D is impeller diameter in feet; Np is the power number, pg/dN3D5; P is power in foot-pounds per second; g is the gravitational constant, feet per second; J is width of baffles; B is number of im-

Jan 1, 1961

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By Robert C. Emmett, Donald A. Dahlstrom

FOR many years gravitational classification has been employed as a basic tool in beneficiation of minerals and coal. While improvements have been made to increase efficiency and fields of application, the basic flow pattern has remained relatively unchanged in most gravitational classifiers. Feed slurry is introduced into a settling pool wherein the coarser solids settle against either an upward or horizontal fluid current, or a combination of both, and are withdrawn in a concentrated underflow stream. The finer solids below the classification point possess too low a settling velocity and thus are primarily swept out through a weir overflow by the bulk of the fluid. Because of its simplicity, efficiency, and low power and maintenance costs, gravitational classification finds a wide acceptance for separations in the range of 20 to 325 mesh. Two factors resulting from the basic flow pattern employed in gravitational classification present difficult problems for certain installations. If classification at 60 to 200 mesh of separation is desired, capacity in terms of gallons of overflow per minute per square foot of settling pool area must be low to allow required solids to settle to the underflow. If a thousand or more gal per min of feed slurry must be handled, either a very large classifier or a number of smaller units must be employed, taking up floor space and head room. Examples of this are the large classifiers, wetting tanks, or settling tanks used in the coal industry for recovery of process water from fine solids. Units as large as 50 or 60 ft in diam and 60 ft high, requiring substantial amounts of steel, are not uncommon. A second factor is the lack of self-regulation in gravitational classification. As overflow rate increases, mesh of separation increases correspond-ingly. This normally heightens difficulties of bene-ficiation. If a fine mesh of separation is required at varying feed rates, the problem is more than doubled in complexity. One of the primary advantages of gravitational classification is low power requirement. Gravity feeds or very low pressure heads are employed. However, fluid energy is still present in the inlet stream and provision must be made to dissipate it properly to obtain uniform upward or horizontal currents. Rather than waste this energy, it would be desirable to utilize it to secure a greater degree of self-regulation and simultaneously increase classification capacity per square foot of area. A free vortex could be realized by a tangential entry of feed, resulting in generation of centrifugal force to complement force of gravity in its action. With greater force it should be possible to gain finer separation at any overflow rate per square foot of settling pool area as compared with conventional gravitational classification. Furthermore, as feed rate increases, centrifugal force will also increase and at least partially offset faster currents sweeping solid particles to the overflow. A more stable classification point should result. The liquid-solid cyclone is the most common application of a free vortex in the beneficiation field.' Many previous papers have emphasized uniformity of flow pattern within the cyclone which is largely responsible for the sharp classification that can be obtained.'-4 If a free vortex can be maintained in a non-pressurized vessel with a free surface it may be possible to realize a third advantage of sharper classification. This open-top cyclone would then achieve a minimum energy loss comparable to that with conventional classifiers, and a mesh of separation coarser than 150 mesh, extending the range of application of the free vortex. This is not presently possible with ordinary cyclones. To test the applicability of the free vortex to gravitational classification, a 30-in. diameter open-top cyclone was constructed at Northwestern University. The cylindrical section was 30 in. high and the conical section included an angle of 60". The feed nozzle was a standard 4-in. pipe entering tan-gentially, with centerline 17 1/2 in. below the top of the cyclone. The conical section was flanged near

Jan 1, 1954

By A. Y. Bethune, W. G. Woolf

THE hydrometallurgy of special high-grade zinc as practiced by the Sullivan Mining Co. at its electrolytic zinc plant, Kellogg, Idaho, involves an important filtration step immediately following the leaching process. By means of the filtration the heavy zinc sulphate solution is separated from the residual products which remain after the zinc calcine has been dissolved in the sulphuric acid electrolyte. Because this plant uses the so-called high-acid, high-density process' for the production of First, the strength of the electrolyte (270g H,SO, per liter) results in a saturated zinc sulphate solution, having a specific gravity of 1.510 to 1.540, which must be kept warm during filtration because of its property of "seeding out" small crystals if allowed to drop much below 60°C. Second, the action of the "high" acid on zinc calcine under the temperature conditions of the leach (80" to 102 "C), although favorable to good zinc extraction, causes a considerable quantity of iron to be dissolved (8 to 18. g per liter) along with variable quantities of alumina and silica, depending on the grade and type of original zinc concentrates roasted. These three, iron, alumina, and silica, are almost completely precipitated during the neutralization of the leach (only a few. milligrams per liter of each remain in solution), so that the resulting pulp, instead of being a granular, sand-like product having a particle-size distribution dependent on the fineness of the zinc calcines leached, is in reality a slimy, chemical precipitate whose filtration characteristics constantly change depending on the amounts of iron silica, and other impurities, which are dissolved and reprecipi-tated. Third, the combination of supersaturated solution of high specific gravity plus a dense, semi-gelatinous residue creates a difficult washing problem requiring a positive displacement wash to liberate the zinc sulphate entrapped in the pulp. In a closed-cycle hydrometallurgical operation, such as practiced in this plant, the extent of washing is determined by the volum,e limitations imposed on the intermediate wash waters by the amount of "fresh" (or process) water which may be added. The volume of fresh water used for makeup purposes is limited to the amount which is lost during the closed cycle by evaporation in the leach, sulphate content of the calcines leached, moisture content of the residue, and spillage. The Burt filter as modified and improved by the Sullivan Mining Co. has successfully met and overcome these difficulties under a variety of zinc plant operating conditions since 1928. It might have many interesting applications to metallurgical fields other than that of electrolytic zinc, and its possible usefulness to hydrometallurgists in general warrants its description and discussion. The Burt filter is so named from its inventor who originated it in Mexico for pulp filtration in the cyanide process for gold and silver ores. While retaining the basic principle of Burt's earlier revolving pressure-type filter with internal filtration media, a number of modifications and improvements have been made in Sullivan Mining Co.'s installation. The Burt filter may be classified as a batch-type pressure filter in contradistinction to either the conventional vacuum-type filter, which depends on atmospheric pressure to force solution through a cloth medium, or to the filter-press, which employs whatever pressure is imparted by the pump delivering the liquid being filtered. The Burt consists essentially of a hollow steel cylinder about 40 ft long, 5 ft in diameter, resting horizontally, and capable of rotation about its long axis. It is supported on one end by a hollow trunnion and near the other end by a riding-ring and roller combination. The cylinder is lined with filter units each fastened against the inside of the shell and parallel to the long axis so as to form a hollow cavity into which pulp may be charged. A specific amount of pulp is admitted to the filter and a unique valving arrangement prevents the loss of pulp while air pressure forces the solution through a canvas medium to the discharge port of each filter unit. The residue is left on the surface of the canvas inside the cavity. The remainder of the filter cycle is concerned with washing the residue free of zinc sulphate, discharging it from the Burt, and preparing the filter for the next charge. A more detailed description of Burt filter construction, a typical filter cycle, and its operating characteristics when employed on material encountered in this plant will be given in that order. Description of the Filter: Fig. 1 shows a side elevation view of a filter with riveted shell construction. Since this drawing was made shells have been fabricated by welding, instead of riveting, with complete success. Shells are lagged on the outside to retain heat. Fig. 1 shows a side elevation and plan view of a Burt filter in operating position. The 1/2-in. steel shells are lined with 3/16-in. copper sheet as protection against the corrosive action of the solution (containing about 500 mg Cu per liter) on iron, and the copper is given a thin protective coating of plastic-base paint. Fig. 2 is a view from the discharge end of the filter, with head removed, before filter units are fastened to the periphery. It shows

Jan 1, 1951

By W. G. Woolf, A. Y. Bethune

THE hydrometallurgy of special high-grade zinc as practiced by the Sullivan Mining Co. at its electrolytic zinc plant, Kellogg, Idaho, involves an important filtration step immediately following the leaching process. By means of the filtration the heavy zinc sulphate solution is separated from the residual products which remain after the zinc calcine has been dissolved in the sulphuric acid electrolyte. Because this plant uses the so-called high-acid, high-density process' for the production of First, the strength of the electrolyte (270g H,SO, per liter) results in a saturated zinc sulphate solution, having a specific gravity of 1.510 to 1.540, which must be kept warm during filtration because of its property of "seeding out" small crystals if allowed to drop much below 60°C. Second, the action of the "high" acid on zinc calcine under the temperature conditions of the leach (80" to 102 "C), although favorable to good zinc extraction, causes a considerable quantity of iron to be dissolved (8 to 18. g per liter) along with variable quantities of alumina and silica, depending on the grade and type of original zinc concentrates roasted. These three, iron, alumina, and silica, are almost completely precipitated during the neutralization of the leach (only a few. milligrams per liter of each remain in solution), so that the resulting pulp, instead of being a granular, sand-like product having a particle-size distribution dependent on the fineness of the zinc calcines leached, is in reality a slimy, chemical precipitate whose filtration characteristics constantly change depending on the amounts of iron silica, and other impurities, which are dissolved and reprecipi-tated. Third, the combination of supersaturated solution of high specific gravity plus a dense, semi-gelatinous residue creates a difficult washing problem requiring a positive displacement wash to liberate the zinc sulphate entrapped in the pulp. In a closed-cycle hydrometallurgical operation, such as practiced in this plant, the extent of washing is determined by the volum,e limitations imposed on the intermediate wash waters by the amount of "fresh" (or process) water which may be added. The volume of fresh water used for makeup purposes is limited to the amount which is lost during the closed cycle by evaporation in the leach, sulphate content of the calcines leached, moisture content of the residue, and spillage. The Burt filter as modified and improved by the Sullivan Mining Co. has successfully met and overcome these difficulties under a variety of zinc plant operating conditions since 1928. It might have many interesting applications to metallurgical fields other than that of electrolytic zinc, and its possible usefulness to hydrometallurgists in general warrants its description and discussion. The Burt filter is so named from its inventor who originated it in Mexico for pulp filtration in the cyanide process for gold and silver ores. While retaining the basic principle of Burt's earlier revolving pressure-type filter with internal filtration media, a number of modifications and improvements have been made in Sullivan Mining Co.'s installation. The Burt filter may be classified as a batch-type pressure filter in contradistinction to either the conventional vacuum-type filter, which depends on atmospheric pressure to force solution through a cloth medium, or to the filter-press, which employs whatever pressure is imparted by the pump delivering the liquid being filtered. The Burt consists essentially of a hollow steel cylinder about 40 ft long, 5 ft in diameter, resting horizontally, and capable of rotation about its long axis. It is supported on one end by a hollow trunnion and near the other end by a riding-ring and roller combination. The cylinder is lined with filter units each fastened against the inside of the shell and parallel to the long axis so as to form a hollow cavity into which pulp may be charged. A specific amount of pulp is admitted to the filter and a unique valving arrangement prevents the loss of pulp while air pressure forces the solution through a canvas medium to the discharge port of each filter unit. The residue is left on the surface of the canvas inside the cavity. The remainder of the filter cycle is concerned with washing the residue free of zinc sulphate, discharging it from the Burt, and preparing the filter for the next charge. A more detailed description of Burt filter construction, a typical filter cycle, and its operating characteristics when employed on material encountered in this plant will be given in that order. Description of the Filter: Fig. 1 shows a side elevation view of a filter with riveted shell construction. Since this drawing was made shells have been fabricated by welding, instead of riveting, with complete success. Shells are lagged on the outside to retain heat. Fig. 1 shows a side elevation and plan view of a Burt filter in operating position. The 1/2-in. steel shells are lined with 3/16-in. copper sheet as protection against the corrosive action of the solution (containing about 500 mg Cu per liter) on iron, and the copper is given a thin protective coating of plastic-base paint. Fig. 2 is a view from the discharge end of the filter, with head removed, before filter units are fastened to the periphery. It shows

Jan 1, 1951

By James E. Lawver, James L. Wright

This paper discusses the design and calibration of a Faraday pail for measuring electric charge. The paper also shows that at least for two minerals, quartz and calcite, the phenomenon that Johnson termed "reversibility" can be explained in terms of the contact electrification between the minerals and the metal rotor of an electrostatic machine. Each year the free world produces about 13 million tons of mineral concentrates by electrical separation. However, despite the rather general acceptance of electrical concentration as an industrial process, research in this area has lagged far behind the efforts devoted to competitive processes such as froth flotation. The purpose of this paper is to describe the design and calibration of a Faraday ice pail for measuring charge density on mineral grains and to show some of its applications in the area of electrostatic and electrodynamic research. THE DESIGN AND CALIBRATION OF FARADAY PAILS FOR ABSOLUTE CHARGE MEASUREMENT The electric charge on mineral grains at a given time during an electrical concentration process is de- pendent on many factors and cannot, in general, be accurately calculated a priori. It is thus important to devise a method of measuring absolute charge. One method of measuring electric charge is to note the potential rise of an electrometer when charged particles drop into a Faraday pail. Two types of Faraday pail have been constructed and calibrated. The first type, Faraday Pail I (Fig. 1), is a vertical pail originally designed to measure the charge due to contact electrification between mineral grains and a metal rotor or due to conductive induction when mineral grains are passed over a rotating cylinder in the presence of an electric field. The second type, labeled Faraday Pail II (Fig. 2), is a horizontal pail designed to determine the charge distribution of mineral grains after they pass through the separating zone of the electrical concentrator shown in Fig. 4. The main advantage of the horizontal pail is its low head-room requirement and the long opening which enables the experimenter to sample the entire width of the stream of minerals being separated. The Faraday pails were calibrated in the following manner: A calibration head was constructed in the form of a hollow brass sphere that could be held in an insulated metal cylinder by vacuum (Fig. 1). The capacitance of the sphere was calculated by measuring the capacitance of the sphere and cylinder combined and that of the cylinder alone and subtracting the latter measurement from the former. The sphere was then charged to a known potential (V) and allowed to drop into the Faraday pail which was attached to a Keithley Model 610A electrometer. This instrument has an accuracy of 2% full scale and has

Jan 1, 1969