Search Documents

By William B. Stephenson

The paper describes efficient and effective methods of automatically controlling sand pump installations. Particular reference is made to liquid-level controls actuating variable speed pump-driving units. Included is a discussion of various combinations of equipment, with schematic drawings. HERETOFORE many sand pump installations in the mining and metallurgical field involved guesswork as to the size of the pump and then simply direct-connecting it to a constant speed motor which was often large enough to do twice the work required. Considerable thought was given to mill location in an effort to avoid as much pumping as possible. Hillside design of mills with gravity flow of tailing was much to be desired. Power costs were low, construction was relatively cheap, and labor in many places meant nothing in comparison to the values derived from the ores being treated. Today most of these conditions are reversed. Ores are becoming more complex with lower values per ton of ore treated. Power is more expensive, construction is higher, and labor definitely takes a very important place in the profit and loss statement. Each piece of equipment incorporated in the design of a new mill is selected to provide highest efficiency based on careful metallurgical analysis of the ore to be treated both in good and poor markets. Each is designed with its flexibility in mind and with provision for adjustments in its performance to meet actual conditions in the operation of the mill. Sand pumps have taken an important position in mill design. Operators the world over are relying more and more on the performance of a sand pump in the efficient operating of their plant. Many ordinary jobs are being done today with pumps driven through adjustable pitch diameter V belt drives in order that changes in pump speed can be made easily to meet actual operating conditions. Operating costs have long since proved the folly of operating a pump at too high a speed for the volume and head to be overcome. Metallurgical results have shown the benefits derived from a smooth flow as contrasted with a surging flow from an uncontrolled pump. Construction costs have risen to such an extent that surge bins and other means formerly employed must be omitted from today's flowsheet. Several installations of considerable size and involving the use of automatic pump speed control have proved the adaptability of such controls since they were placed in service in recent years. In the following presentation each system discussed and illustrated schematically has been installed in one or more instances and has proved itself reliable and definitely economical from the standpoint of capital investment, operation and maintenance. Variable Volume-Head Speed Control: Probably the most popular, effective and efficient automatic control setup for sand pumping installations is the one shown schematically in fig. 1. Commonly used

Jan 1, 1951

By E. Martinez

Samples of chrysocolla, a hydrated copper silicate, from several sources were submitted to differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). Pure samples of chrysocolla are difficult to obtain, but comparison of the thermograms revealed that certain reactions occurred in all the samples and it is believed that these are due to chrysocolla. Endothermic reactions occurred at 50' to 175°C, 350° to 650°C, and 980° to 1100°C each of which caused a weight loss. The first was due to loss of absorbed water, the second to a dehydroxyla-tion, and the last to conversion of cupric oxide to cuprous oxide. In addition, two exotherms were detected at 685° to 695°C and 935° to 950°C. Infrared spectra of heated samples indicated the exotherms were due to changes in the Si-O bonds leading to the formation of cristobalite. The results suggest that thermoanalysis may be a more sensitive and reliable method for detection and identification of chrysocolla than X-ray diffraction, infrared spectroscopy, or optical techniques. The recovery of copper from oxide ores has been the subject of much research in recent years. One of the principal oxide copper minerals is chrysocolla, a hydrated copper silicate. The information available indicates that it has widely varying properties depending on the origin of the sample. Chrysocolla usually is found associated with cuprite, malachite, azurite, and native copper. Most of these other minerals are recoverable by present day flotation practice, but chrysocolla is lost in the tailings.1,2 Reagent combinations that are successful in laboratory chrysocolla flotation with ore from one source have been found to be ineffective with another. 3 The copper segregation process offers an altemative to beneficiation of oxide copper ores.4'5 The process is of particular interest in processing ores that cannot be treated by conventional hydrometal-lurgical methods because of high acid consumption.6 Fuller understanding of the properties of chryso- colla may prove of value in the search for improvements in both the flotation and segregation processes. Chrysocolla samples from Arizona, Nevada, Chile, Australia, and South Africa were studied by differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). In addition, a sample from Miami, Ariz., was submitted to X-ray diffraction, infrared spectrophotometry, and surface area determination. EQUlPMENT Differential thermal analysis (DTA) detects, amplifies, and records exothermic-endothermic reactions occurring in a sample as its temperature is raised at a constant rate, usually 10°C per minute. A unit made by the Robert L. Stone Co. was used in this study. The tests can be run if desired with a gas streaming through the sample. A thermobalance records weight losses or gains in a sample as its temperature is raised at a uniform rate. The gravimetric analyses (TGA) were run on a Chevenard thermobalance converted electronically for graphic recording. A heating rate of 5°C per minute up to 1100°C was used in these tests. A Perkin Elmer Infrared Spectrophotometer, Model 221, was used with sodium chloride optics obtaining IR spectra in the 2 to 15 region. A pressed pellet technique with potassium bromide was used in preparing the samples. DESCRIPTION OF SAMPLES Chrysocolla is found in the oxidation zones of copper deposits commonly associated with malachite, azurite, and limonite. It is described as varying in composition and in color from bluish green to brown or black. Inclusions of other minerals are usual so chemical analysis of chrysocolla samples may be suspect. However, it is considered by many to be a solid solution of CuO, SiO2 and H2O with a general formula of CuO.SiO2.2H2O or 2CuO.2SiO2. 3H20. Chukhrov7 has described a sample from the Urals as CU3,5(OH)2 (A1Si3) O10.nH2O and stated that chrysocolla is a montmorillonite-type mineral. On the basis of X-ray analyses, DTA, and infrared spectra of chrysocolla from the Inspiration Mine, Ariz., Sun8 concluded that there was no substantial evidence to classify it as a montmorillonite mineral. Chrysocolla samples from many sources throughout the world were used in this investigation. The following is a brief description of each:

Jan 1, 1963

By R. L. Vaughn, H. R. Cooper

Effective procedures for automating X-ray assaying by use of a computer are described. These methods have advanced the reliability and accuracy of X-ray data, resulting in improved assay measurements for operations and process control. Assaying by X-ray fluorescence has been increasingly used in the mining industry, both for analyzing solid H.R. COOPER and R.L. VAUGHN, Members AIME, are Systems Manager-Process Control and Senior Systems Analyst, respectively, Scientific & Engineering Computer Center, Ken-necott Copper Corp., Salt Lake City, Utah. TP 68B236. Manuscript, Sept. 16, 1968. Discussion of this paper, submitted in duplicate prior to Dec. 15, 1969, will appear in SME Transactions, March 1970, and AIME Transactions, 1970, Vol. 247. samples for smelter, mill, and mine quality control and for continuous stream analysis of mill pulps. The important advantages of X-ray methods in laboratory assaying are reduced analysis cost, and for on-stream analysis, immediate assay information on concentrator streams. Of the numerous reports concerning X-ray analysis in mining operations, one is referenced which can be noted particularly in conjunction with this report.' Continuous X-ray analyzers have been installed at Kennecott and other mining companies for operations guidance and trend information for a number of years. However, complete operational acceptance of X-ray data has been slow, due* to the difficulty in maintaining

Jan 1, 1970

By W. C. Hellyer, R. A. Campbell

In single-stage autogenous grinding, the buildup of a critical size fraction in the media can be corrected by removing this material through pebble ports, crushing it below the critical size range, and recycling the crusher product back to the mill. The rate at which the critical sire fraction is crushed affects the size distribution of the grinding media and this in turn affects the sire distribution of the grate discharge. This provides a means for controlling the grind produced by a single-stage autogenous grinding unit. The pilot-plant investigations on a very hard copper ore were carried out at the facilities of the Institute of Mineral Research, Michigan Technological University. In an autogenous grinding circuit in which feed at approximately 9 in. top size is reduced to a size suitable for subsequent processing, the build-up of a "critical size" fraction in the media causes problems. A "critical size" fraction has been defined as, "media too small to effect reduction by impact grinding of ore coarser than a quarter of an inch and too large to be broken by the largest size of media in the charge."' The buildup of a critical size fraction reduces the capacity of the mill, increases the grinding power requirements per ton of finished product, and generally produces a finer grind than is desired. It is general practice to overcome this problem by the addition of large-diameter steel balls to the grinding charge. This has certain disadvantages, such as 1) an increase in mill liner wear, 2) wear on the steel balls, and 3) some loss in flexibility in grinding circuit opera-tions resulting from difficulties in removing the steel balls by means other than by grinding out. A number of investigators' have suggested the use of a small crusher as an alternative to the use of large steel balls for control of a critical size fraction. With this technique the critical size fraction is removed from the mill continuously through suitable-sized pebble ports, crushed below the critical size range, and returned to the mill. An external crusher would cause no increase in mill liner wear, the wear on the crusher would be less expensive than the wear on the steel balls, and the flexibility of the grinding circuit operations would be enhanced since the crusher can be cut into and out of the circuit at will. This paper describes an autogenous grinding pilot plant investigation on a very hard copper ore, which led to the selection of an autogenous grinding flowsheet incorporating a crusher in the circuit as the preferred method for grinding this ore. Further development of the technique demonstrated that the crusher could be employed to control the product produced by an autogenous grinding unit. Description of Pilot Plant These investigations were carried out at the pilot-plant facilities of the Institute of Mineral Research, Michigan Technological University, Houghton, Mich. The autogenous grinding unit was a 6-ft-diam by 2-ft-long Hardinge Cascade mill with %-in. slotted steel grates. Pebble ports cut into the grates allowed passage of pebbles having a top size of approximately 21/2 in. The grate and pebble-port discharge material passed over a 3/16-in, trommel screen with the trommel under -size being pumped to the classification device. DSM screens were employed for classification in the early investigations, but were replaced later by a small Dorr rake classifier. The trommel oversize material returned to the mill via scissor conveyors. When the crusher was incorporated into the circuit, the trommel oversize material passed to a double-deck vibrating screen fitted with 2-in. and 3/4-in. square mesh screen cloths. The $ 2-in. and the —2 +3/4-in. size fractions were combined and crushed to a nominal 1 in. in a 4 x 6-in. jaw crusher. The crusher discharge and the —3/4-in. screen undersize returned to the mill. In the final investigation, a flap gate fitted to the top deck of the screen directed the — 21/2 +2-in. size fraction to the crusher or back to the grinding mill. The flap gate operated manually on a 15-min cycle, with this material being crushed for so many minutes out of each cycle. This was found to be a more reliable method of controlling the amount of — 21/2 +2-in. material crushed than attempting to make a split of a small weight of material on a continuous basis. Operational Techniques: Each sample was sized into three fractions and the Cascade mill feed was reconstituted from these size fractions in the proportions existing in the original sample. Sufficient feed for 15 min operation was weighed out and fed by hand over a 15-min period. The gross mill power draft was recorded every 15 min, corrected for tare power and drive efficiency, and reported as net kilowatt-hours per ton. A recording kilo-watt-hour meter provided a continuous visual record of the power drawn by the mill. Pulp densities of the trommel undersize and the classifier overflow (or DSM undersize) were taken every 15 min. The classifier overflow was sampled automatically. Timed samples of the classifier sands were taken every 15 or 30 min, weighed, and returned to the circuit. After correcting for moisture content, the weights were converted into percent circulating load. Timed samples of the +2 in., the —2 + 3/4-in., —3/4-in. screen fractions, and the crusher discharge, were

Jan 1, 1971

By R. J. Brison, O. F. Tangel

The development of a new method of mineral separation was sponsored by the International Salt Company, which requested Battelle Institute to investigate means for improving the quality and appearance of rock salt from the Company's Detroit mine. Although developed specifically for removing impurities from rock salt, the general method may be applicable to other separation problems. The principal impurities in rock salt from the Detroit mine are dolomite and anhydrite which represent 2 to 5 pct of the weight of the mined salt. In the size range from 1/4 to M in. (the range of primary interest in this project) the impurities are only partially liberated from the halite in normal production. Further size reduction to improve the liberation of impurities is not practicable in view of the market requirements for the coarse grades of rock salt. Laboratory separations in heavy liquids showed that, to improve the quality and appearance of the rock salt substantially, it would be necessary to remove not only free gangue particles but also a large proportion of the locked-in particles. Because rock salt is an inexpensive commodity, a low-cost process was required. Gravity methods were, of course, considered. The heavy-liquid separations indicated that a split at an effective specific gravity of 2.2 to 2.3 would be required. (The specific gravity of pure halite is 2.16.) Heavy-media separation was investigated but had the disadvantages that it was necessary both to operate with saturated brine and to dry the cleaned salt, and that the cleaned salt was darkened by the magnetite medium. Air tabling was tried but did not give the desired separation. It soon became apparent that established methods would not provide a satisfactory solution and work was undertaken on the development of a new process to solve the problem. PROCESS DEVELOPMENT Preliminary Experiments: At the start of the investigation, an analysis of the problem indicated that the diathermacy of rock salt—that is, its ability to transmit radiant heat—might form the basis for an efficient separation process. Under this theory, the impurities might be selectively heated by radiant heat. The particles could then be fed over a belt coated with a heat-sensitive substance so that the warm impure particles would adhere preferentially to the coating. After the initial experiments, made by heating the rock salt with an infrared lamp and separating the product on small sheets of resin-coated rubber, proved encouraging, a small continuous separation unit was set up. This comprised 1) a simple heating unit consisting of a vibrating feeder covered with aluminum foil and an infrared lamp mounted above the feeder and 2) a separation belt 6 in. wide and 36 in. long. A sketch of the device is shown in Fig. 1. Results with this apparatus confirmed the fact that a good separation was possible. It was apparent, however, that a considerable amount of experimental work would be needed to develop the scheme to a practical and economical process. The Process: Basically, the process consists of two main steps: 1) selective heating by radiation and 2) separation of the heated particles on a heat-sensitive surface. Because neither of these steps had previously been utilized commercially in mineral processing, it was necessary to do basic research on both aspects. Factors studied in the investigation included type of heat source, design of heating unit, design of separation belt, selection of heat-sensitive coating, removal of heated particles from the belt, contact between particles and coating, and maintenance of the heat-sensitive surface. Part of the experimental work was carried out on a small-scale unit consisting of the 36x6 in. belt and auxiliary apparatus, and part on a larger unit. For simplicity, discussion of work on both of these units is grouped together. SELECTIVE HEATING Radiant-Heat Source: The essential requirements for a radiant-heat source were 1) that the radiant heat be in a wave length range which is effectively absorbed by the impurities but not absorbed appreciably by the rock salt and 2) that it be dependable, practical, and economical. Selection of a heat source of suitable wave length range was one of the first considerations. It is well known that pure halite is highly transparent to radiant energy in wave lengths from 0.3 to 13 microns. However, the available data on infrared transmission by dolomite and anhydrite, particularly in the range below two microns, were not complete enough to serve as a reliable basis for selection of a heat source. Although it may have been possible to obtain sufficient data on infrared transmission and absorption to enable one to select the best heat source, a more direct procedure was used. This consisted simply of exposing the crude rock salt to each of several types of radiant-heat source on the small continuous separation device. The heat sources investigated, approximate source temperature used, and calculated wave length of maximum radiation are tabulated in Table I. Of the two types of tungsten-filament lamps investigated, both the short wave length photoflood lamps and the longer wave length infrared lamps were satisfactory from the standpoint of selectivity

Jan 1, 1961

By S. Usui, I. Iwasaki

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

Jan 1, 1971

By I. C. Edwards, G. E. Agar

INTRODUCTION Many research efforts over the past years have been devoted to the study of energy consumption in comminution and the characteristics of the broken material. The old "laws" have been superseded by more general methods of analysis1,2,3 but these methods have not received much attention outside the laboratory. The work to be described here was done in order to obtain scale-up information on wet rod milling of sylvinite. EXPERIMENTAL A 30 x 51 in. Denver rod mill with a scoop feeder was used for these tests. Ore was fed from a variable speed screw conveyor into the feed box along with saturated brine. The mill was filled to approximately 45% of its volume with 2400 lbs of a mixed rod charge. The minimum possible feed rate was 2.5 tph and the practical maximum for this mill was 8 tph. Individual tests were run by setting the solid feed rate, then adjusting the brine flow rate to give the desired pulp density. After steady state, as evidenced by constant discharge pulp density, had been reached, samples of the discharge were taken at 15-minute intervals. These samples were filtered, washed with alcohol, dried, then screened on Tyler screens using a Ro-tap shaker. The sylvinite feed was regular plant feed that had been scalped at 20 mesh and was tapped from the coarse ore bins. It had been previously passed through a roll crusher set at 3 inches and after scalping was essentially minus 1 in. plus 20 mesh. RESULTS The screen analysis of samples from each series of tests was plotted on double logarithmic scale paper to determine the reliability of the series. In some cases, a faulty sample or screening procedure dictated that the results of one test be discarded. The criterion is a simple one — if the slope of one plot was noticeably different from the others in the series, it was discarded. The only cases that caused difficulty were those in which the brine was not thoroughly displaced from the solids. When this occurred agglomerates were easily identified on the screens and the whole of the sample was retained on 3 or 4 screens. There were no border line cases. The number N given in Table 1 is the number of size analyses that were pooled to form the data shown in Fig. 1. While examining the reliability of the sizing and sampling by plotting the data, it became apparent that there was no significant difference between grinds made at 50 and 60% solids. For the analysis given here, these tests were regarded

Jan 1, 1967

By W. J. Carlson, I. lwasaki

Manganese and silica in nonmagnetic tailings from the direct reduction-magnetic separation step were so closely associated that no physical concentration methods were effective. Dilute sulfuric acid dissolved manganese and iron rapidly; manganese in the form of silicates was virtually insoluble in sulfuric acid solutions, thereby governing the maximum extraction attained by leaching. Autoclave leaching at elevated temperatures and at high oxygen pressures reduced the iron extraction to less than one tenth without affecting the manganese extraction. There was an indication that autoclaving solubilized the manganese silicates appreciably. Elimination of iron from leach solutions was carried out by aerating the solution in the pH range of 6.5-7; a platinum-calomel electrode pair provided a reliable indicator for the end point of iron removal. The purified leach solutions were free of interfering impurities for the recovery of manganese as metal, dioxide, or carbonate. Manganiferous iron ores of the Cuyuna Range, Minn., are too low in iron content to be directly useful as iron ores, and too low in manganese content to be termed manganese ores. Common methods of physical beneficiation such as gravity concentration and flotation have not been successful in upgrading this material. The R-N process, when applied to manganiferous iron ores,1'2 produces sponge iron by direct reduction of iron oxides, while leaving manganese as manganous oxide. From fine grinding and magnetic separation of the reduced ore, a tailing is produced which is relatively rich in manganese and low in iron content. The nonmagnetic tailing can readily be leached with acids and then further processed to recover manganese in various forms. In this way, two processes, neither of which would be an economic solution to the Cuyuna ore problem by itself, may be combined to produce two marketable products, thereby rendering the over-all process economically more attractive. This article reports on the results of a leaching study on the nonmagnetic tailings, both in a batch agitation vessel and in an autoclave; on the problems involved in the purification of leach liquors by aeration; and on the recovery techniques used to produce manganese metal by electrolysis, manganese dioxide either by electrolysis or by autoclaving, and

Jan 1, 1968

By J. R. Johanson

With the continuing development of automatic operations, it is important that reliable discharge rates through hoppers and bins be obtained for bulk solids. In most applications a feeder is used to control the feed rate. The selection of this feeder size is usually based on the rated ability of the feeder to move material without regard to the capability of the material to flow through the opening in the hopper onto the feeder. The dis charge rate through a given hopper opening may vary from zero to a theoretical maximum depending on the flow properties of the material. This paper presents an equation for the maximum discharge rate for a bulk material as a function of the hopper shape, the opening size and the flow properties of the material. Discharge-rate tests on various materials indicate good agreement with calculated values for opening sizes from a fraction of an inch to several feet. As part of a program to develop improved design criteria for bulk storage and handling facilities, a specific study of the methods involved in calculating the rate of discharge of bulk solids from hoppers and bins has been made. This study is closely related to earlier work done on the prevention of flow problems through the choice of proper design geometry of hoppers and bins1. The present paper is an extension of this earlier work, and therefore many of the concepts previously derived and explained will be used without explanation. Two types of flow will be considered: plane-strain flow in a wedge-shaped hopper and axial-symmetry flow in a conical hopper. The literature contains many studies on discharge rate for regular-shaped dry particles2*3 such as glass beads, lead shot and coarse dry sand. However, the work is empirical in nature, and it is difficult to extrapolate the given information to apply to moist, sticky materials typical of many industries. Some studies4 have shown a decrease in discharge rate with an increased moisture content of the material. However, no attempt was made to correlate the discharge rate with moisture. The method of calculating discharge rate presented in this paper was derived analytically from the basic knowledge of the flow of frictional, cohesive, granular solids, The equation includes the effect of material cohesion (stickiness), hopper configuration, opening size and bulk density of the material. THEORETICAL WORK The basic assumptions made in deriving the discharge rate equation are as follows: 1) The material is a frictional, cohesive, granular solid. 2) The flow is considered continuous and without significant density change. 3) The material at the hopper opening forms a continuously failing arch that is in dynamic equilibrium. 4) The strength of the material in the arch, which must be overcome for the arch to fail, is caused by the pressure in the flowing mass near the hopper opening. 5) This pressure at the hopper opening can be calculated5. If we consider the dynamic equilibrium of an arch of uniform thickness such as that shown in Fig. 1,

Jan 1, 1965

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 Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951

By Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951

By R. B. Booth, W. F. Linke

The continuous interest of the American Cyanamid Company in producing superior polymeric flocculants and dispersants for the mining industry has resulted in a broad, general study of the physical chemistry of mineral-polymer systems. The aim of this program has been to establish a sound theoretical basis on which practical problems could be analyzed, and which would lead to the development of new and improved flocculating and dispersing agents. This study has required the cooperative effort of physical, analytical, and polymer chemists, as well as engineers familiar with mining problems. This report summarizes the results of some of our investigations over the past two years. Further studies along these and other lines are in progress and will be reported in the near future. MEASURING FLOCCULATION IN THE LABORATORY The foremost problem in studying flocculation in the laboratory has been to establish reproducible methods of comparing and evaluating the effectiveness of various reagents. Previous workers have measured settling -rates 1,8 filtration rates2 , and sediment volumes3 . Direct observations of the floc size, or the clarity of the supernatant liquid, can also be used. AIthough these parameters are all interrelated, they do not always measure the same physical properties of the suspensions, and hence are not equally useful or reliable. Whatever the method of comparison selected. the testing procedure must be carefully defined and standardized. The results must be satisfactorily reproducible, and for practical purposes must accurately reflect the results obtained in mill use. In order to compare the data obtained in these various types of test, the physical nature of the flocculated system must be examined in some detail. In an unflocculated system, only the larger or denser particles (the "sands") settle at a useful rate. The sediment can be observed to build up from the bottom of the cylinder, while the fines remain in suspension for a long period of time. If the system is not disturbed, even the microscopic fines can be seen to classify themselves as they fall over a period of days. When a high molecular weight polymer is added to the system, the colloidal particles are tied together randomly in flocs, and the sands and fines settle together. The line of demarkation between solid and liquid phases moves downward from the top of the cylinder. The settling rate of the flocs depends on their size and on their effective density. For a given mineral sample, the larger the flocs, the faster the settling. Since the large flocs cannot pack as closely as small ones, they produce a sediment with more pore spaces, and hence of larger volume, which is easier to pump and to agitate. Although it is nearly impossible to measure individual floc sizes, it is relatively easy to determine settling rates and sediment volumes. Of these two, the settling rate is simpIer to measure, and hence generally more practical. Reproducible sediment volumes can only be obtained by allowing the solids to pack for a rather long period of time (e.g., hours to days), and although the volumes thus observed are directly related to the degree of flocculation (average floc size), it is often impractical to wait so long for the results.

Jan 1, 1961

By J. F. Shirley, H. K. Burke

In January, 1964, three years of laboratory and pilot plant test work reached its culmination. The San Manuel Div. of the Magma Copper Co. switched its molybdenite extraction circuit from the sodium hypochlorite-ferrocyanide process to the new process described in this paper. The San Manuel Div. Concentrator of the Magma Copper Co. is located near San Manuel, Ariz., approximately 45 miles northeast of Tucson. The concentrator treats approximately 35,000 tons of ore per day containing about 0.85% copper and 0.025% molybdenite. The San Manuel ore body is a disseminated copper deposit in quartz monzonite and quartz monzonite porphyry. The primary copper mineral is chalcopyrite; minor amounts of chalcocite, bornite and covellite are also present in varying degrees. Pyrite makes up somewhat more than half the sulfide content of the rock. The molybdenite is nearly always found on fracture faces or associated with quartz veinlets.' On January 9, 1964, the sodium hypochlorite-ferrocyanide process that had been used for molybdenite recovery was replaced by the following process: 1) A conditioning step with hydrogen peroxide, sulfuric acid, sodium cyanide and zinc sulfate. 2) A two stage rougher flotation step with stove oil and sodium ferrocyanide. 3) One cleaning step with sodium ferrocyanide. 4) One cleaning step with sodium hypochlorite and potassium ferricyanide. 5) Four cleaning steps with an anti-foam Exfoam 636 and potassium ferricyanide. COPPER CONCENTRATOR Since molybdenite recovery is closely allied with previous steps in the operation, the copper concentrator will be briefly described first. The ore is ground to approximately 5% plus 65-mesh and 65% minus 200-mesh in a two-step grinding circuit consisting of rod mills in open circuit and ball mills in closed circuit with rake classifiers. A straight forward rougher circuit producing only one concentrate is used. The rougher concentrate is combined with the discharge from the regrind ball mills and is then cycloned, with the overflow going directly to the cleaner circuit and the underflow going to the regrind mills. The feed to the cleaner circuit is approximately 90% to 94% minus 325 mesh. The cleaner circuit consists of two stages in which the tailing from the final cleaner stage is returned to the primary cleaner and the primary cleaner tailing is returned to the rod mill discharge at the head of the circuit. Approximately 0.013 lb of the allyl ester of amyl xanthate (S-3302) per ton of ore is added to the rod mills and approximately 0.006 lb of sodium isopropyl xanthate (2-11) per ton of ore is stage added down the rougher banks. The frother used is methyl iso-butyl carbinol (MIBC) and it is added at various points in the grinding and the rougher flotation circuits. Lime is added to the ball mills and/or to the rod mills to maintain the desired pH in the rougher flotation. The rougher flotation has been operated at various pH levels between 10.5 and 11.5. Fuel oil is added in the regrind mills at a loading of approximately 0.01 lb per ton of ore. Factors Which Directly Affect Molybdenite Recovery: It has been found at San Manuel that, to obtain maximum molybdenite recovery, an oily type flotation collector must be added in the grinding circuit. When molybdenite is crystallized in the characteristic platy form, which appears to be the case almost all of the time at San Manuel, its surface has a greater tendency to adhere to oils than water. 2 Of the several oily type collectors available on the market, the allyl ester of amyl xanthate seems to be the best molybdenite promoter. The floatability of San Manuel's molybdenite does not appear to be affected, at least up to a pH of 11.5, by lime. It has been reported that at several plant operations high lime alkaline circuits have adversely affected the molybdenite recovery.3,4 It has been found through microscopic studies that the bulk of the chalcopyrite will float before the molybdenite in both the rougher and cleaner circuits. It is necessary to add fuel oil to the regrind mills to enhance the

Jan 1, 1965

By J. Walter Snavely

PRESENT-DAY processing in many industries, calcining, sintering, briquetting, beneficiation and nodulizing, increasingly calls for the handling of large volumes of hot bulk materials. Various types of conveyors have been employed. This discussion will cover the factors governing their selection. For temperature ranges up to 400°F, or approximately 200 °C, a wide range of conveyors is available. Special constructions of rubber conveyor belts, steel conveyor belts, vibrating and shaker conveyors, apron conveyors, and drag chain conveyors, all are used successfully. As temperatures go well above 400 2F, however, choice of conveyors is narrowly limited. This paper will consider the problem of handling bulk materials only where the temperatures exceed 400°F. The arbitrary selection of 400 °F as a dividing point undoubtedly can be challenged, as special conveyor belting constructions are available which are suitable for temperatures in excess of 400°F. However, when the relatively short life of such belts and the cost of their replacement, with the attendant down time, are balanced against the reliability and long service life of the properly designed steel constructed units to be discussed, there is little question in any operator's mind that the special belts are more expensive to use. Because the conveyors under study are for the handling of bulk materials, inevitably including a high proportion of fines, obviously wire mesh belts cannot be included for consideration. Even though this type of conveyor is widely used at high temperatures, i.e., for carrying glassware through a lehr, it is unsuited for the conveying of bulk materials, and therefore will be excluded from further discussion in this paper. Preliminary to the study of the conveyor itself is the determination as to whether the material is to be cooled while it is being handled, or whether the processing requires retention of all heat and the maintenance of a given temperature range. In the majority of cases cooling is incidental to or part of the handling process, when the handling, for example, follows completion of sintering, roasting, calcining, refining, or some other process. To meet such operating conditions successfully, the conveying medium used must have: 1—a construction capable of withstanding maximum initial temperatures of the material being handled. 2—a construction providing efficient heat transfer for cooling. 3—a construction providing dependable operation and long life with minimum service requirements, and 4—a construction providing controlled and efficient conveying. Under the usual conditions of cooling during the handling, the construction selected to withstand the initial maximum temperatures does not necessarily involve using alloys, as excellent results can be achieved with normal carbon steels and cast irons, when they are properly applied and proportioned. The earliest and simplest type of conveyor for handling very hot materials is the cast steel drag chain conveyor, still widely used for handling hot cement clinker, as illustrated by Figs. I and 2. Because of the rugged and generous proportions of the chain link design, low carbon steels are entirely suitable for the links. The pins, however, must be alloy steel. The simple, rugged construction of this type of conveyor makes it readily capable of withstanding high initial temperatures, even though the chain is operating buried in the material. The drag-chain type of conveyor has advantages and limitations. Although the efficiency of the heat transfer is relatively poor, the life of the conveyor is reasonably long, and because of its crude simplicity it does not require much servicing. However, as a conveyor, it is limited in capacity, and largely limited to horizontal runs. Furthermore, because of the crude design, heavy weight, and the chain operating at the temperature of the material, greatly reducing permissible operating chain pulls, this type of conveyor is limited to relatively short centers. Another type of conveyor that has been used for very hot materials is the cast pan conveyor. Because of its very generous proportions the cast pan, which is made of either cast iron or malleable iron, can withstand initial maximum temperatures. It also provides efficient heat transfer for cooling. Further, it is on efficient conveyor construction, which can be used for inclines. Because the chain employs rolling friction instead of sliding friction, and is not in the maximum temperature zone, much longer centers are possible. It is this type of conveyor that is frequently used in the casting of various metal pigs, pig iron, and aluminum; it is obvious, therefore, that very high initial temperatures are being handled. With this kind of conveyor the return run is frequently sprayed with water to accelerate heat transfer. The build-up of residual heat in the very heavy cast pans is thus overcome. The outboard roller steel pan conveyor is an improved pan conveyor' which provides high rates of heat transfer and substitutes formed steel pans for the heavy cast pans. It is a very efficient conveying medium. The details of this particular construction are clearly shown in Fig. 3. An early application of this type of conveyor is shown in Fig. 4. In this case the conveyor units are handling roasted phosphate rock at average temperatures of 1000" to 1500°F, and frequent maximum temperatures as high as 1900°F. Several widths are used. The capacity of the unit at a speed of 50 fpm is approximately 30 tph per inch of width at peak loadings, average capacity being about 1/3 of peak loading. The assembled conveyor is shown in Fig. 5, with views of both the top and the underside to show all the construction details. In particular, the following general design principles were carried out in this construction:

Jan 1, 1954

By H. W. Hitzrot, R. H. Limons, L. V. Fegan

The testing of a new method for measuring pellet-bed temperatures in the traveling-grate or shaft furnace is presented in this paper. This method, using silver-palladium alloy wires as indicators, overcomes disadvantages existing in the methods based on the use of thermocouples which are costly and time-consuming. The authors have developed and tested a new method for measuring pellet-bed temperatures in the traveling-grate or shaft turnace through the use of silver-palladium alloy wires as indicators. This method overcomes the disadvantages of existing methods, which, based on the use of thermocouples, are generally costly and time-consuming. For example, in the case of the traveling-grate furnace, the installation of an instrumented pellet, together with its attached thermocouples, requires stopping the furnace, with resulting loss of production time. As for temperature measurement in the shaft furnace by the existing method, the movement of the shuttle-car feeder seriously interferes with both the actual insertion and the resulting distribution of continuously fed thermocouples. The authors have overcome these difficulties by using, as temperature indicators, a range of silver-palladium alloy wires that can be easily recovered, after firing at any desired location in a pellet furnace. Silver-palladium wires having specific melting points from 2000°F to 2500°F are enclosed in ceramic capsules. Each capsule is placed in a ni-chrome wire basket filled with a sample of green pellets, the baskets being the means for recovering the capsule after firing. After the encapsulated wires have been exposed to the firing cycle, the alloy wires are examined to determine the maximum temperatures attained by the pellet sample. This method gives accurate, rapid determinations of maximum temperatures attained during pellet induration in either the traveling-grate or shaft furnace The new method has the following advantages over existing methods: 1) it is less expensive; 2) no furnace downtime is required; 3) the testing units can, be made in quantity in advance; 4) measurements can be made in rapid succession at all desired locations; 5) measurements of maximum firing temperatures are more accurate; 6) on the basis of points 4 and 5, the new method gives a more reliable picture of the correlation of pellet quality, petrographic textures and maximum temperatures, thus providing a more effective quality-control tool. A disadvantage of the method is, of course, that the measurements are not continuous and only maximum temperatures are determined. TEST PROGRAM Preparation of Materials: CALIBRATION OF ALLOYS - In the following discussion the term "bead point" (see Fig. 1) refers to the temperature at which each alloy begins to melt and form a bead. Fig. 2 gives the bead point curve for silver-palladium alloys. Before the alloys are used in plant tests, the bead point of each alloy should be determined in the laboratory to verify the alloy compositions. Once a given alloy has been calibrated, it can be

Jan 1, 1965

By T. P. Chen, F. W. Bowdish

Studies made with a captive bubble apparatus on the sulfidization and collection by amyl xanthate of true chrysocolla specimens have defined the ranges of pH value and sulfide concentration which permit contact between the bubble and the mineral surface. Titanium compounds were the most effective of the materials found to activate the sulfidization of chrysocolla. With titanium activation, the contact angles and the ranges of pH value and sulfide wncentration giving bubble contact were all increased. Chrysocolla ores were concentrated by flotation. Chrysocolla ores occur at many localities in grade and quantity sufficient to make mining and millin feasible, but no satisfactory method of concentratio has been found. Although chrysocolla may be leached with acid, only those ores without acid-consuming gangue may be leached economically. Because of its potential importance, a study of the conditions nece sary for flotation of chrysocolla has been carried ou The literature contains a few references to flotation of chrysocolla. Two methods were developed by the U. S. Bureau of Mines.1,2 The first consisted of a fatty acid soap and a high xanthate as collectors of chrysocolla from a synthetic ore, while the second involved the use of hydrogen sulfide and xanthate. Ludt and DeWitt3 demonstrated the difference in adsorptive powers of chrysocolla and quartz for bas triphenyl methane dyes and suggested the use of butyl, hexyl or octyl-substituted malachite green as collector. Jackel4 emphasized the effects of combin tions of reagents such as Aerofloat 31, pine oil, and Reagents 404 and 425 with sodium sulfide and zinc hydrosulfite as conditioning agents. Although he reported recoveries of 89% from a synthetic ore and 98% from a natural ore containing azurite, malachite, chalcopyrite and chrysocolla, careful application of Jackel's method to chrysocolla from Tyrone, N.M., failed to give a high recovery. MATERIALS AND TECHNIQUE Samples from Inspiration, Ariz., and Tyrone and Magdalena, N. M., were used for experimentation and verified as true chrysocolla by leaching tests, specific gravity tests and X-ray diffraction. Chrysocolla does not dissolve at pH 4, although malachite and azurite do. Chrysocolla is about half as dense as the copper carbonates. X-ray diffraction analyses by the powder camera method confirmed the samples as true chrysocolla. A captive bubble apparatus, which cast an enlarged image of the air bubble and the mineral surface upon a screen, was used to check on the character of the surfaces. The specimens were prepared by grinding a flat surface on a glass plate using fine abrasive; then they were washed and kept in distilled water until they were to be treated with reagents. Before each reagent treatment, the specimen was carefully checked for cleanliness in the captive bubble apparatus. It was assumed that the surface was clean if, after fine grinding and washing of the specimen, the bubble would not stick. Specimens were handled with glass forceps, and precautions were taken to avoid contamination of the mineral surfaces. Contact angle measurements were carefully made several times on each treated specimen to obtain reliable average values. EFFECT OF pH VALUE AND SODIUM SULFIDE CONCENTRATION In each experiment, a specimen with a freshly ground surface was immersed for 10 min in a solution of sodium sulfide, washed and immersed for 15 min in a solution containing 30 mg per 1 of potassium amyl xanthate. The specimen was then washed again in distilled water and tested for contact angle in the captive bubble apparatus while submerged in distilled water. In this series of experiments, the pH of the sulfidizing solution was varied from 3 to 7, and the concentration of sodium sulfide, containing 60% Na2S, was varied from 50 to 650 mg per 1. Many combinations of pH value and sulfide concentration resulted in no contact between the bubble and the surface, but over a limited range of conditions, contact angles varying from 24ºto 52ºwere obtained. The data in Fig. 1 show sulfidization conditions that lead to bubble contact and those that do not. The region of contact is surprisingly small, which may indicate why flotation of chrysocolla involving sulfidization has proven so difficult in practice. Several features of the system are illustrated in Fig. 1. In the region between pH values of 4 and 6 with sodium sulfide concentrations below about 350

Jan 1, 1963

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 E. J. Roberts

The theory is developed that the tons ground through a given mesh per day in a wet ball mill is proportional to the percent plus that mesh in contact with the balls and the net power applied to the balls at this point. A grindability test is described. DURING the course of a study of the fundamentals of classification in 1937, the need for a more basic understanding of the action of a ball mill became acute. Unless one knows how classification affects grinding, one cannot hope to effectively improve on classification. The methods of evaluating grinding efficiency that depend on surface developed were studied but soon discarded for two reasons: 1. There was no apparent method which could be generally used to give a reliable figure for the actual new surface developed as a result of grinding. Subsequent papers have not changed this conclusion. 2. The practical evaluation of grinding in the main ore dressing applications was in terms of the percentage retained on a screen which passes 90 to 99 pct of the material and not in terms of surface area. The Probability Theory With the background of our experience in the field of closed-circuit grinding, together with the papers of Lennox,1 Gow,2 Gaudin,8 Fahrenwald,4 Coghill, and others, the approach of the theoretical physicist was then tried. The thought was somewhat as follows: When one grinds in a ball mill, a given expenditure of power leads either to a certain number of point to point blows per hp-hr or to a certain distance of line contact per hp-hr, depending on whether the action of the balls is considered to be cascading or rolling. It is also assumed that the balls actually come together on each blow or during the roll. Then a volume of slurry will be covered per minute which is some function of the size of the particle being considered (see fig. 1). All particles coarser than this size will be reduced through this size. This volume of slurry contains a certain weight of ore, depending on the percent solids and the density of the solids. If we fix the percent solids and the density of the solids and let w be this certain weight of ore in the volume covered, then, in mathematical terms, what we have just postulated is, w —— 8 hp (a) dt If W is the total weight of ore present in the mill, then we can write. W w/8 hp (b) W dt and if C is the cumulative percent plus the size chosen at the start of the time interval dt, w w c/dt W 8 hp x c (c) wc But wc/100 is the weight plus the size chosen which at 100 wc the close of time dt is finer than that size, and W is the decrease in the percent plus of the whole mass of ore or —dC. Then, —W dC/dt 8 hp x C. (d) In other words, the mesh tons ground through a given size per unit of time is proportional to the hp and the percent plus the mesh. A crude analogy would be to picture a 1-ft-wide steam roller going down the road at 1 ft per sec. If we place one egg on the road per square foot, one egg will be smashed per second. If we place a dozen eggs per square foot, a dozen eggs will be crushed per second. Similarly, if all the particles in w are plus the mesh, i.e., C=100, we should have a maximum rate of reduction. If only 10 pct of them are plus the mesh (C=10), we would have only one tenth the maximum rate; if only 1 pct are plus the mesh, the balls have a hard time finding anything to work on. This is where the term "probability theory" comes from. The chances of the balls crushing a particle through a given mesh depends directly on the concentration of particles coarser than this mesh in the general pulp in the mill. Giving W the units of tons and dividing equation (d) through by W, we obtain -dC hp ----- = k---— C [1] dt ton where k is a constant for any one size of particle, density of solid and moisture content of pulp. Eq 1 is the rate equation for a first order reaction and says that the rate of decrease of the percent plus a given mesh with time is directly proportional to the hp per ton applied to the body of ore and to the percent plus the mesh in the ore mass as a whole. Since it is a differential equation, it only

Jan 1, 1951