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By Nathaniel Arbiter, Norman L. Weiss

Factors now accelerating the trend to larger concentrators and larger equipment units are reviewed. After almost 40 years of stability with unit sizes less than 100 cu ft, 200 and 300-cu-ft flotation-cell units are now operating and a 500-cu-ft machine is being built. The design of these cells is considered, as is the design and operation of flotation circuits against the background of the need to achieve greater process efficiency while holding down operating and investment costs. Particular attention is paid to the scale-up problem. With the newer cells having five to ten times higher unit capacities, the advantages resulting from this in sampling, mill operation nnd control, testing and plant layout are indicated, as well as the broad direction for changes in the overall design of 50,000 to 100,000-tpd concentrators. A comparison is given of the available larger cells of U.S. manufacture with analysis of differences and similarities. After almost 60 years of application in this country, froth flotation is still by far the most important process for concentrating metallic ores as well as a number of nonmetallic ores including fluorspar, phosphate rock, and potash. Its use on iron ores and coal is increasing. In the metals groups, the largest plants and the greatest progress are found in copper and molybdenum mills, and the section of this paper that deals with applications will be found most pertinent to that field. During its long reign, the flotation process has passed through several phases of evolution and has achieved progress in economy, efficiency. and simplicity of operation. Much of today's success in the flotation of low-grade ores can be attributed to that progress. Even more important to us than our ability to cope with today's conditions are tomorrow's problems, which we can contemplate with some prescience and confidence. Our annual production of copper, lead, zinc, and molybdenum—to mention only the most important metals occurring in the U.S. as sulfides- must be held at a high level in the face of a declining grade of ore. It is probable that the tonnage of such ores mined and processed will double in this country before the turn of the century. Not only will many of the existing flotation mills be enlarged again and again, but larger mills will be built here and abroad to meet our needs. At this moment, in fact, there are at least three proposed new foreign mills or expansions that envision 100,000-tpd milling rate on copper ores. Table 1' shows the trend in copper ore grades in U.S. in recent years. Table 2' shows the projected annual growth rate of consumption 1966-1985. A study by one of the authors 25 years ago showed that of 32 copper mills in North America with more than 100-tpd capacity, only eight had capacities exceeding 10,000 tpd. The average for these eight was 25,000 tpd. These included the three giants of that day, Magna, Arthur, and Morenci; the other five averaged a little over 10,000 tpd. Today instead of the eight mills with capacities exceeding 10,000-tpd capacity we have 25, and the mill with less than 1000-tpd capacity is fast becoming extinct. Much less range is found in lead and zinc mills, which are still generally small except for the few with over 1000-tpd capacity in Missouri, Utah, and Idaho, and a few zinc ore mills in the same size category. Molybdenum and potash operations are the fastest growers in North America today, but when one considers the projected increase in consumption of copper together with the projected decrease in grade of ore (3% annually), he can foresee that copper ore milling will soon lead the list. Technological improvements to meet the challenge of lower-grade ores have taken the wholly predictable path toward bigger equipment. In mining, shovels, draglines, scrapers, and trucks have led the way; in ore processing, we have seen a rapid increase in size of grinding mills, flotation cells, and pumps. Twenty-five years ago, for example, the 10-ft-diam ball mills treating 2000 tpd at Morenci were considered the giants, but today some grinding mills in similar service have five times that capacity. The specific matters to be dealt with in this paper are 1) the underlying principles that affect flotation-machine design, and the conditions required to produce

Jan 1, 1971

By Rollin A. Slater

This paper presents a composite of ideas on materials handling systems by describing the design of an over-all system for a hypothetical steel pellet plant located on the Great Lakes. The bulk materials handling industry has served for years as the skeleton and the central system of nerves and nourishment for the body of basic industries. During the past two decades, this body has undergone substantial growth and mutation to meet the surging demands of our national economy for more and better products. The steel industry has kept pace with spiralling demand by innovation which has resulted in new processes and improved designs which generally yield greater production from fewer units. Materials handling technology has likewise improved its techniques and currently has a new, stronger generation of equipment to perform its necessary supporting function. Drawing upon excerpts from several recent plant designs, this paper describes a basic approach to the selection and application of modern bulk materials handling systems to an entirely new steel plant. Specific details are not available for publication, and therefore the plant described herein is a composite of ideas. This imaginary plant will be located on the United States side of the Great Lakes and its owner will be called the Hypothetical Steel Corp. for purposes of discussion. TYPES OF EQUIPMENT The basic bulk materials handling systems and components required in a plant so situated are: lake boat unloaders; railroad train unloading systems; storage pile stackers and reclaimers; large bins and regulating feeders; and the conveyor systems — with their attendant controls— which connect the foregoing. PROGRAM OF DEVELOPMENT The initial work on the plant must be that of the process engineers who develop and refine the basic data, i.e., the logistics of all raw materials, waste materials and products within the plant. During this phase of study, the materials handling engineer may advance his planning only in basic areas such as movements at the dock and movements at the railroad. This initial planning work is concerned only with defining logical plant locations in the light of channels, currents, existing traffic, governmental regulations, locations of existing railroads and available rights-of-way. These systems alone have their own individual criteria which must be respected and accommodated in the final layout. All other elements, such as the arrangement of the blast furnaces, power house, coke plant, by-product plant, scrap preparation plant, BOF shop, and facilities for continuous casting, degassing, calcining, etc., are dictated by their own specific needs and their relation to each other. In the arrangement of these latter facilities, the materials handling system is secondary and must be designed to accommodate. BASIC LOGISTICS The result of the process engineer's work can be expressed in terms similar to those shown in Table I. Plants are customarily rated in terms of annual tons of ingots. Therefore, the basic size of this plant can be determined by referring to the BOF Shop wherein it is sized at 5 million tpy. From the base, the required hot metal can be traced to the blast furnace, which in turn indicates the amount of required flux, iron bearing materials and coke —which in turn establishes the demand for coal. Likewise, the total demand for flux can be determined by the combined demand of the lime plant and blast furnace. The generation of fines and waste products, which have significant value but are not usable as such, create the need for a sintering plant. From projections such as the foregoing, the complete flow of materials can be established and accommodated. The basic premise upon which this typical materials balance is predicted is that the plant will be a producer of high tonnages of non-SpecialtV- products. All basic raw materials (flux bearing, iron bearing, coal) will be prime quality. For example, the basic iron charge will be pellets, and the iron ore fraction will be sized lump, all of which

Jan 1, 1970

By Shiou-Chuan Sun

THE presence of an excessive amount of tough froth in the flotation of minerals, particularly coals, may create trouble in dewatering, filtering, and handling. Froth is also a nuisance in many chemical industries.' This paper presents a study on the destruction of extremely tough froths with intense high-frequency sound. The data indicate that sound waves can be employed for continuous atandsoundwavescan instantaneous defrothing. A powerful high-frequency siren was used in obtaining the data. Also tested was an ultrasonorator of the crystal type with a frequency range of 400, 700, 1000, and 1500 kc per sec and a maximum power output from its amplifier of 198 w. The results, not presented, indicate that as now designed this machine is not suitable for defrothing. Although the sound generators of the magnetostriction type2,3 and of the electromagnetic type'.' were not available, it is beelectromagneticlieved they are capable of producing the required sound intensity for defrothing. The use of ultrasonics for defrothing was suggested by Ross and McBain1 in 1944. Ramsey8 reported in 1948 that E. H. Rose mentioned a supersonic device that broke down flotation froth but with low capacity. The writer has not been able to find any published literature containing practical experiments. Theoretical Considerations The mechanism of defrothing by sound is attributed to the periodically collapsing force of the propagated sound waves and the induced resonant vibration of the bubbles. The collapse of froth is further facilitated by the sonic wind and the heat of the siren. Sound waves can exert a radiation pressure'," against any obstacle upon which they impinge. When a froth surface is subjected to the periodic puncturing of sound waves, the bubbles are broken. According to Rayleigh9 and Bergmann,12 the radiation pressure of sound, P, in dynes per sq cm is given as: P = 1/2 (r+1)i/v where r is the ratio of the specific heats of the medium through which sound is traveling and is equal to 1 on the basis of Boyle's law; i is the sound intensity in ergs per sec per sq cm, and v is the sound velocity in cm per sec. In this case, the accuracy of the formula is only approximate, because a perfect reflection can hardly result from a column of froth. In addition to the radiation pressure, the propagated sound waves cause the bubbles of the froth to have a resonant vibration.'" he vibratory motion of the bubbles causes collision and coalescence, thereby weakening if not breaking the bubble walls. Sonic wind and heat were also generated." The sonic wind can exert pressure on the froth surface, and the heat can evaporate the moisture content of the bubble walls as well as expand the enclosed air. Apparatus The defrothing apparatus, shown in Figs. 1 and 2, consists of a powerful high-frequency siren, a glass or stainless steel beaker of 2-liter capacity with 12.4 cm diam and 17.1 cm height, and a metal reflector. The beaker was placed 2 in. above the top point of the siren. The metal reflector was adjusted to reflect and focus the generated sound waves into the central part of the beaker. Fig. 2 shows the crystal probe microphone used to measure the acoustic intensity and the mandler bacteriological filter employed to introduce compressed air into the beaker for frothing. The apparatus was enclosed in a soundproof cabinet equipped with a glass window. The siren, shown in Fig. 3, consists of a rotor that interrupts the flow of air through the orifices in a stator. The rotor, a 6-in. diam disk with 100 equally spaced slots, is driven by a 2/3 hp, Dumore W2 motor at 133 rps. The frequency of the siren can be varied from 3 to 34 kc. The maximum chamber pressure is about 2 atm, yielding acoustic outputs of approximately 2 kw at an efficiency of about 20 pct. The siren itself is relatively small and can be operated in any orientation. A detailed description of the siren has been given by Allen and Rudnick.11 Collapse of Froth To study the sequence of the collapse of froth, the glass beaker was partially filled with 920 cc water, 100 g of —150 mesh bituminous coal, 0.3 cc petroleum light oil, 0.2 cc pine oil and 1.54 cc Pyrene foam compound. This mineral pulp was agitated for 5 min and then aerated through a mandler filter until the empty space of the beaker, approximately 9 cm high, was filled completely with min-

Jan 1, 1952

By M. C. Fuerstenau, A. M. Gaudin

A homogeneous suspension is viewed by X-rays. The radiation density seen, affords a measure of the extra absorption due to the solids contained. This radiation density, at a predetermined depth, varies with time. The chart of this relationship gives the desired determination. s it is frequently necessary to know the distribu-rt tion of particle sizes in various products used in industrial applications, a number of means have been devised for these determinations. Some of them include direct microscopic measurement, screening of product, extraction of samples from a settling slurry, air and liquid elutriation, adsorption of gases and liquids, and light absorption. The absorption of light as a means for particle size analysis has been known and employed for some time. X-ray absorption has not received the attention that light has enjoyed primarily because of the complex and costly instrumentation involved with X-rays. In 1954, Brown and Skrebowskil suggested the use of X-rays for this type of determination but did not supplement their suggestion with experiment. In 1958, the authors conducted size analysis experiments on suspensions of quartz with the Trans-viewer: an X-ray unit specifically adapted for sedimentation research, and Rossa conducted experiments in a similar manner in 1959 with a stable gamma emitter, AmZ4l, on suspensions of uranium oxide. To illustrate the applicability of X-ray absorption for analyzing particle size distributions, an essentially mono-energetic and well collimated X-ray beam was used as a source. The beam was made mono-energetic by placing a zirconium filter before the counting window together with feeding the pulses from the proportional counter into a reverter, a device capable of discriminating against pulses of various heights. The sketch in Fig. 1 shows the collimation arrangement in more detail. As shown, the X-ray beam was collimated by inserting a 3/8-in. ID pipe into one of the ports of the X-ray tube and further by two small steel plates with slits 3/32-in. wide and 1/2-in. long placed in front of the window of the Geiger-Mueller (G-M) tube. One of the plates was located 1/2 in. and the other 2 1/2 in. from the window of the counter; both plates were contained in an aluminum holder mounted on the box containing the G-M tube. This type and degree of collimation should eliminate any intensity due to scattered radiation; the transmitted intensity would then be solely a functior. of the absorption by the suspension at that level. Experimentally, homogeneous suspensions of finely ground quartz were prepared, Daxad (a sulfo-nated compound made from wood) was added as dis-persant, and the transmitted X-ray intensity measured as a function of time at a fixed depth in the suspension. In more detail, the experimental procedure involved the following: 1) measuring the transmitted X-ray intensity through the Daxad solution at a predetermined position or settling depth in the absence of solids, 2) placing a known amount of dry solids into the vessel and agitating thoroughly, 3) measuring the transmitted X-ray intensity at time zero (i.e., immediately upon cessation of stirring), and 4) measuring the transmitted X-ray intensity as a function of time (a 10-sec counting interval being chosen and counting being started 5 sec before the desired time). Since a linear relationship exists between percentage solid in a slurry and the logarithm of transmitted intensity for a homogeneous material, the cumu-

Jan 1, 1961

By William Lodding

Muscovite was the crystalline material used in the experiments discussed in this paper. It was both wet and dry ground for various periods of time. The effects (physical and chemical changes) produced in grinding are analyzed by use of thermal analysis. That grinding may change a crystalline material both in a physical and in a chemical sense has been shown by a number of authors in recent years. Mackenziel and Bartha2 reported the effects of careful grinding on muscovite, quartz, gypsum, dolomite, etc. Increase in specific surface produced by grinding implies increase in total energy of a system. Newly produced surfaces are chemically more reactive; dislocations may be introduced during grinding and, on the other hand, existing strain may be released by the same process, as shown by Holt, et a1.3 All these changes are easily detected by thermal analysis, but it is not easy to separate the multiplicity of effects registered and to arrive at quantitative interpretations. This paper is an attempt to analyze, by the use of thermal analysis, some effects produced in grinding. In differential thermal analysis (DTA) one measures the sum of all thermal changes occurring in a substance during heating or cooling; e.g., chemical changes, such as decompositions, synthesis, oxidation, reduction, etc., and physical changes such as phase transformations, changes in specific surface and thermal conductivity, strain release, etc. By thermogravimetric analysis (TGA), on the other hand, only chemical changes involving weight loss or gain are recorded. Similarly, gas evolution analysis (GEA) will register only reactions involving gas evolution. By performing two or all three of these methods of thermal analysis simultaneously and on the same specimen, one can separate the different effects registered and, by difference, evaluate their magnitude. The results of TGA are usually plotted as percent weight against furnace temp. When one differentiates weight against furnace temp, a rate of weight-change curve is obtained, and the resulting differential thermogravimetric (DTG) plots are similar in appearance to DTA thermograms for the same reaction. While the integrated area under a DTA peak represents enthalpy and can be expressed in calories per gram the area under the corresponding DTG peak represents the weight change which took place during a reaction. The similarity of DTA and DTG thermograms obtained from one material can be seen in Figs. 1 and 2. EXPERIMENTAL A Fisher mortar grinder and a porcelain mill with burundum balls were used for both dry and wet grinding. Muscovite was first carefully hand-filed and screened on 20 and 100 mesh. The middle fraction was considered essentially unaltered muscovite.

Jan 1, 1968

By N. F. Schulz

The purpose of this investigation was to evaluate the effects of variables in equipment and procedure on tube test results. This information was then to be used to formulate specifications, preferably within the practical ranges of existing machines, for test procedures capable of yielding precise and meaningful data concerning the magnetic separation of any magnetite ore, regardless of its origin. A survey of current practice was conducted among laboratories using the Davis tube test. Many tube tests were performed to experimentally evaluate the variables most likely to affect test results. A simple method was then sought for adequately interpreting tube test results. The Davis magnetic tube concentrator is a laboratory machine designed to separate a small sample of pulverized magnetic iron ore into magnetic and non-magnetic fractions. It was developed in 1921 by mechanization of a test previously performed manually No significant changes have been made in the design of the apparatus since that time. The Davis magnetic tube test continues to serve as a practical basis for judging the amenability of an ore to magnetic separation and for controlling magnetic separation plant operations. The apparatus consists of an inclined cylindrical glass tube supported adjacent to closely spaced pointed poles of a powerful C-magnet. (Fig. 1). To conduct a tube test, a sample of suitably prepared ore is poured into the water-filled tube. With wash water flowing through, the tube is oscillated to gradually reject the non-magnetic and weakly magnetic ore particles until the desired degree of separation is attained. Both the magnetic concentrate and the tailings may be recovered for examination. MAGNETIC SEPARATION BY THE DAVIS TUBE Mechanical separation of a specified mineral component from an ore usually requires comminution to fine sizes for sufficient liberation. Some distinguish- ing property of the mineral is then exploited to sort the desired mineral particles from the gangue particles. The magnetic separation of iron ores is based on the fact that magnetite, Fe3O4, is strongly attracted by a magnetic field while the gangue minerals are not. Maghemite, ?-Fe2O3, which may occur in certain chemically processed ores, responds similarly to magnetite. The magnetic process is practical only when the loss of iron in the form of relatively nonmagnetic hematite, Fe2O3, goethite, Fe2O3.nH2O, or other iron mineral can be tolerated. In the Davis magnetic tube concentrator, magnetic attraction holds magnetically susceptible particles in a magnetic field. Forces due to gravity, inertia, and fluid-solid friction tend to remove the less susceptible particles from the field. The probability that a given middling particle will be rejected from the tube in a given period of time is an inverse function of its magnetite content, with some modification due to particle size and mineral specific gravities. The net result is that ore particles tend to be rejected in their order of increasing magnetite content, while the retained portion or magnetic concentrate increases in grade continuously during the process. MAGNETIC SEPARATION CHARACTERISTIC The magnetic separation characteristic or MSC of a magnetite ore is the correlation between concentrate

Jan 1, 1964

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 J. N. Ong, J. P. Ratledge, J. H. Boyce

SINCE German operators opened the Tsumeb mine in the early 1900's, continuous operation has been interrupted only by enforced shutdowns during two world wars and the depression of the 1930's. Original metallurgical practice was blast furnace smelting. This was applicable only to low zinc ore, leaving large tonnages of high zinc ore to be stockpiled on the surface. A gravity concentrator using tables, buddles, and jigs operated for several years, but differential flotation of the ore did not get beyond the experimental stage. When the property was put up for sale by the South African Custodian of Enemy Property there was an estimated 434,000 tons of ore on surface dumps. The dump material, assaying 4.48 pct Cu, 15.8 pct Pb, and 10.9 pct Zn, had become highly oxidized during a stay of many years on the surface. But these dump ores had to be used in evaluating possible metallurgy because the shaft became flooded after the 1940 shutdown and underground ore was unavailable. Many companies examined Tsumeb but quickly lost enthusiasm because of the formidable metallurgical problems presented by the dump ore. However, on the basis of extensive test work, the present company decided that differential flotation would be successful and purchased the property after competitive bidding. Geology and Mineral Association is Complex The Tsumeb mine lies near the northern boundary of South West Africa at an altitude of 4200 ft. The village of Tsumeb is 335 air miles northeast of Wal-vis Bay, territorial port of entry. The orebody is a hydrothermal replacement deposit associated with an irregular intrusion of aplite cutting folded pre-Cambrian dolomite beds. In plan the deposit is a lens of 250x600 ft maximum dimensions. In section the orebody dips 55" south from the surface to 2000 ft, below which there is a reversal in dip to 75" north. Mineralization is localized in the aplite and in the adjoining brecciated dolomite. In the upper levels the ore is largely confined to a single lens, but below the 1890 level massive sulphide ore forms north and south lenses that partially enclose a pipelike body of medium to low grade disseminated ore, part of which is of too poor a grade to mine. Drilling below the 24th, or bottom level, has disclosed that a high percentage of the mineralization from the 3000 to 3300-ft horizons is comprised of copper and lead oxides. The deep oxidation may be local, since it appears due to circulating waters in a major fault cutting the orebody. Tsumeb ranks as one of the highest grade base metal deposits in the world. It is geologically and mineralogically famous for 41 copper, lead, zinc, vanadium, and germanium ore minerals, but its complex mineral assemblage is not attractive to metallurgists. In addition, there is a gradual change in oxidation intensity from the upper levels to the lowest level now operating, which is the 24th. This is well illustrated by the copper minerals. In the upper levels malachite and azurite are predominant, in the intermediate levels chalcocite, and in the lower levels tennantite. Throughout the mine there are sufficient heavy metal soluble salts to activate zinc, and large amounts of depressants are required even for so-called clean sulphide ores. When the same sulphide ore is exposed to prolonged oxidation aboveground, as was the case with the dump ores, reagent requirements become fantastic. Plant Construction First staff members arrived at Tsumeb in March 1947, and by May essential services had been restored and mine unwatering began. In 2 1/2 months company employees built a temporary jig and sorting plant for treating existing dumps to provide a saleable product during high metal prices. Meanwhile, concentrator design and construction was contracted to a South African firm, with subcon-

Jan 1, 1956

By E. H. Crabtree, V. W. Winters, T. G. Chapman

IN 1939 one of the authors* described advances in carbon-cyanidation for the period 1932 to 1939 • T. G. Chapman: A Cyanide Process Based on the Simultaneous Dissolution and Adsorption of Gold. Trans. AIME (1939) 134, 207. and included: (1) the dissolution of gold in an ore pulp by cyanide and its simultaneous adsorption by carbon, (2) the stage addition of adsorptive carbon and its movement countercurrent to the flow of the ore pulp, and (3) the recycling of adsorptive carbon. Considerable additional experimental and pilot plant work have been done by the authors since 1939 and the results of such work E. H. CRABTREE, JR., Member AIME, is Director of Milling, Eagle-Picher Mining and Smelting Co., Miami, Okla., V. W. WINTERS is Superintendent, Sahuarita Concentrator, Eagle-Picher Mining and Smelting Co., Tucson, Ariz., and T. G. CHAPMAN, Member AIME, is Dean, College of Mines, University of Arizona, Tucson. AIME Sun Francisco Meeting, February 1949. TP 2744 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received March 3, 1949. have been sufficiently encouraging to justify a second paper. Developments from 1939 to 1948: In 1941, Verne W. Winters did the first experimental work in eliminating agitation during dissolution and adsorption and thereby demonstrated that the dissolution of gold by cyanide and its simultaneous adsorption by carbon take place essentially to completion in a stagnant pulp. Shortly thereafter, and as a result of the stagnant pulp experimental work, Winters and Chapman demonstrated that soluble gold, in a cyanide ore pulp, would migrate from the ore pulp through a perforated container and be adsorbed by carbon within the container. During the same year, the Eagle-Picher Mining and Smelting Co. erected a 25-ton capacity pilot plant at the Harquahala mine about 9 miles from Salome, Ariz. Experimental work at this plant comprised chiefly the stagnant pulp modification of carbon-cyanidation although considerable work was done in the grinding and activation of carbon, and testing various methods of extracting adsorbed gold from carbon. Soon after the completion of the Harquahala

Jan 1, 1951

By E. H. Crabtree, T. G. Chapman, V. W. Winters

IN 1939 one of the authors* described advances in carbon-cyanidation for the period 1932 to 1939 • T. G. Chapman: A Cyanide Process Based on the Simultaneous Dissolution and Adsorption of Gold. Trans. AIME (1939) 134, 207. and included: (1) the dissolution of gold in an ore pulp by cyanide and its simultaneous adsorption by carbon, (2) the stage addition of adsorptive carbon and its movement countercurrent to the flow of the ore pulp, and (3) the recycling of adsorptive carbon. Considerable additional experimental and pilot plant work have been done by the authors since 1939 and the results of such work E. H. CRABTREE, JR., Member AIME, is Director of Milling, Eagle-Picher Mining and Smelting Co., Miami, Okla., V. W. WINTERS is Superintendent, Sahuarita Concentrator, Eagle-Picher Mining and Smelting Co., Tucson, Ariz., and T. G. CHAPMAN, Member AIME, is Dean, College of Mines, University of Arizona, Tucson. AIME Sun Francisco Meeting, February 1949. TP 2744 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received March 3, 1949. have been sufficiently encouraging to justify a second paper. Developments from 1939 to 1948: In 1941, Verne W. Winters did the first experimental work in eliminating agitation during dissolution and adsorption and thereby demonstrated that the dissolution of gold by cyanide and its simultaneous adsorption by carbon take place essentially to completion in a stagnant pulp. Shortly thereafter, and as a result of the stagnant pulp experimental work, Winters and Chapman demonstrated that soluble gold, in a cyanide ore pulp, would migrate from the ore pulp through a perforated container and be adsorbed by carbon within the container. During the same year, the Eagle-Picher Mining and Smelting Co. erected a 25-ton capacity pilot plant at the Harquahala mine about 9 miles from Salome, Ariz. Experimental work at this plant comprised chiefly the stagnant pulp modification of carbon-cyanidation although considerable work was done in the grinding and activation of carbon, and testing various methods of extracting adsorbed gold from carbon. Soon after the completion of the Harquahala

Jan 1, 1951

By Milton E. Wadsworth, Edward M. Eyring

BETTER understanding of solid surfaces and their associated adsorption products is of both academic and practical value. The study of detergents and their behavior in cleaning surfaces is fundamentally related to the nature of the products adsorbed and their ability to compete for certain surfaces. Similarly in flotation, the nature of the adsorbate before and after adsorption is basic to clear understanding of the process. The recently developed pellet technique1,2 employing infrared spectroscopy provides a powerful tool for the study of solid surfaces. Absorption bands, which are infrared-active, may be observed for adsorbed products. Also any changes in the surface substrate that occur as a result of adsorption may be detected by means of differential spectra. In applying the pellet technique to solids of large surface area, French and his coworkers3 have surveyed the adsorption of organic flocculants on clay minerals, oleic acid on fluorite, and ammonia on cracking catalysts. Recent work on application of differential infrared spectroscopy to the study of solid surfaces' has demonstrated that the difference between a sample and a reference can be obtained with double-beam instrumentation. In this way displacement of surface products, adsorption of new products, and differences in the substrate, if any, may be detected. Experimental Procedure: Potassium iodide pellets were used in this study, prepared from recrystallized analytical reagent grade KI. Grinding was carried out in a Fisher mechanical agate mortar. Two-gram plates containing the solid under study were pressed in a special evacuated die constructed in the laboratory at the University of Utah." Concentration depended on the solid in question. Zinc sulfide and willemite are sufficiently transpar- ent in the infrared region for concentrations as high as 20 mg of mineral in 2g of KI to be used. Zinc oxide is even more transparent, permitting concentrations as high as 60 mg in a 2-g plate. For non-differential operation, however, much lower concentrations (2 to 10mg/2g KI) had to be used. Spectra were obtained on a Per kin-Elmer 21 double-beam recording infrared spectrophotometer. A sodium chloride prism was used giving a spectral range of approximately 2 to 15 v wave length. For differential runs a resolution of 1000 was used and the scanning speed was controlled commensurate with instrument response. The response was checked continuously and each spectrum was repeated several times. For nondifferential runs a resolution of 927 was used. Zinc oxide used was analytical grade oxide fume. The zinc sulfide was precipitated from solution with analytical grade zinc sulfate and sodium sulfide. The precipitate was washed and centrifuged several times and dried. Before use the precipitate was washed in dilute HCl and then treated with saturated NH4Cl to remove any bulk oxides and sulfates. X-ray patterns of the precipitate were comparable to that of the natural mineral sphalerite. The precipitate was much more satisfactory than finely ground natural ZnS because of its very large surface

Jan 1, 1957

By Rudolph G. Wuerker

WITH the growing complexity of ore dressing processes and the diversity of equipment, efficiency control has become increasingly important in beneficiation. In the case of iron ore dressing, there have been sporadic attempts'-3 to establish the optimum separation for various ores so that grade and recovery would be readily predictable. But no rigorous statistical analysis and graphical presenta- tion of such tests could be found in the literature of this country. Tests Made with Iron Ores Expanding upon existing methods of plant supervision and efficiency control, an investigation was started in the ore dressing laboratory at the University of Illinois on the applicability of distribution analysis to iron ores. A sample of crude ore was obtained from the Ohio mine of Cleveland-Cliffs Iron Co. at Ishpeming, Mich., consisting mostly of alternating bands of hematite and siliceous gangue, both about 1/4 in. wide. The ore was heavily coated with limonite, which often filled the cracks.

Jan 1, 1959

By R. Janes, F. Everard

Pilot plant studies have been conducted on a variety of iron ores of differing composition and grain size to test their amenability to dry autogeneous grinding and dry magnetic concentration. A general description of the pilot plant and its equipment is given with particular reference to the Aerofall mill and the dry magnetic separators developed by the Ontario Research Foundation. Test results are reviewed in detail with emphasis on power consumption, mill speed, the use of ball charges, types of ores, moisture content, separator speeds, and design. In 1957 the Ontario Research Foundation installed a pilot plant (Fig. 1) for the study of dry autogenous grinding and dry magnetic concentration of iron ores. A 61/2-ft Aerofall mill was installed as the dry autogenous crushing and grinding medium. The Cavanagh fast drum and the Double Helix, two dry magnetic separators developed at Ontario Research Foundation, were also installed. Iron ore samples of 10 to 150 tons have been investigated. These ores varied widely in mineralogy and grain size. Taconite (-325 mesh), fine-grained magnetite (200 to 325 mesh), and medium-grained magnetites (65 to 150 mesh) have been treated with varying degrees of success. Lately, considerable research has also been conducted on the metamorphosed magnetite-hematite complexes of the Labrador Trough. Operational and metallurgical data have been compiled from these tests. The effects of ore type, air flow, mill speed, and ball charge on mill operation are discussed. Metallurgical data are also presented to illustrate the operation of the dry magnetic separators. PILOT PLANT DESCRIPTION Fig. 2 depicts the normal flowsheet used. Minus 6-in. ore from a vibrating feeder is fed to the 6%-ft Aerofall mill powered by a 50-hp 550-v induction motor. The main blower, a 25-hp 550-v fan rated at 7000 cfm, draws air through the mill. Finely ground ore is conveyed by the air stream to a horizontal classifier, where expansion and a change in direction cause the +I50 mesh fraction to fall onto a vibrating

Jan 1, 1962

By O. E. Palasvirta

Several drum-type, low-intensity dry magnetic separators have recently been described 1,2,3,4 all of which share at least two generic similarities. The magnet consists of a multiplicity of poles of alternating polarity and, the speed of the drum shell, relatiw to the magnet, is such that the magnetics are agitated through rotation of the magnetic particles around their individual axes. The speed of the drum relative to the magnet (differential speed) determines the frequency of the field reversals. For effective agitation this frequency generally has to be above a specific value, which varies according to the mean particle size of the magnetics. With separators of small diameter, say less than 11/2 ft., a sufficient differential speed for treatment of fine magnetite can be attained only by rotating both the magnet and the drum separately; separators of large diameter may attain sufficiently high differential speeds with stationary magnets. Separators of large diameter thus have the advantage of simplicity of design. The capacity of a separator of this type is directly proportional to the speed of the shell. As the centrifugal force opposing the magnetic force increases at a slower rate than the peripheral speed of the shell when drum diameter is increased, large drums can be operated with higher peripheral speeds. Thus they have a higher capacity than smaller drums. Separation is also a statistical phenomenon, i.e. the grade of concentrate improves as the number of agitations or rotations of the magnetic particles per separation increases. As separators of large diameter have space for a large number of magnets they are best suited for production of optimum grade concentrate. The statements above are general clues For separator design. In the following an expression IS derived showing a more exact relationship between the pertinent separator variables. As both magnetic attraction and centrifugal forces are directly proportional to the volume of the affected bodies, complications in mathematical analysis, arising from variable particle size, are avoided by considering the force per unit volume of affected material. For the magnetic force per unit volume Runolinna (3) gives the following equation: V " 1 + N(ur, - 1) uo dr (1) where ur and uo stand for the relative permeability and the permeability of vacuum, respectively, N is the demagnetization factor, B the intensity of the magnetic field, and r the distance from the face of the magnetic pole. The permeability of a specific magnetite ore can generally be considered as a constant. For ground, liberated material N can also be treated as a constant, and the first simplification can be made by writing ur-1 1=k1. 1 + N(ur - 1) uD k1. Fig. 1 shows the field intensity curves for two U-magnets of different size, but of roughly similar geometry of cross-section as plotted against the distance from the plane of the pole faces. The curves

Jan 1, 1961

By Pablo Hadzeriga

Great Salt Lake brine was subjected, in laboratory scale, to conditions simulating solar evaporation. Solid phases and the variation in composition of the liquid phase throughout the potassium salts crystallization zone are presented. These experimental results are discussed and compare with stable equilibrium data of the quinary system Cl-, SO4, Na+, K+, Mg + + and H2O. The Great Salt Lake brine in Utah is a vast reserve of several potential chemicals which has attracted much attention in the recent past. From the standpoint of an ore reserve, it may be said that the lake contains a low grade ore which has to be concentrated by evaporation of water. Because of the fairly favorable weather conditions in the Great Salt Lake area, it seems reasonable to consider solar evaporation as an economical means of concentration. From the view of phase chemistry, the main constituents of Great Salt Lake brine make it belong to the five component system: C1-, SO4=, Na+, K+, Mg++ and H2O. Because of their low concentration, the presence of other elements may be regarded as having little effect on the equilibrium of the five main components. Equilibrium data for the quinary system of our interest has been available for many years. Using these data, it is possible to calculate how the brine of the lake will concentrate when subjected to solar evaporation. In practicing solar evaporation, the weather factor and supersaturation conditions cannot be ignored, but their introduction into calculations using equilibrium diagrams is almost an impossible task. With the experimental work presented in this paper, an attempt is made to compare the stable equilibrium data of the quinary system as applied to the case of the Great Salt Lake brine and the dynamic equilibrium when that brine is subjected to continuous solar evaporation. Sea water brine belongs to this quinary system when only the major constituents are considered. The study of this system has been carried on since the last decades of the nineteenth century. The reason for these intensive studies was to give a geological explanation to the underground potash deposits. AS a result, many theories were presented trying to relate the equilibrium data with the geological formation, paragenesis and metamorphism of these deposits. Two main approaches were followed for the equilibrium studies of this quinary system. One of them, and the best known to us, was the school started in Germany in the last century by J. H. Van't Hoff. The other one, which started about the same time, was the school of N. S. Kurnakov in Russia. Both approaches were different and as a consequence generated opposite theories about the geological formation of potash deposits. The Van't Hoff school studied the quinary system of sea water under careful laboratory control, resulting in stable phase diagrams. During these studies, Van't Hoff recognized the difficulties in crystallizing some double salts.' ". . . . we unexpectedly found that some compounds whose formation in the solutions under investigation was possible at 25°C, nevertheless totally failed to put in an appearance. These, not to mention the calcium salts, were leonite, MgSO4 . K2SO4 4 H2O; kainite, KC1-MgSO4 . 3 H2O; and kieserite, MgSO4 .H2O. Even exceedingly slow crystallization with the addition of the compounds themselves did not remove the condition of supersaturation in the case of these bodies." However, those salts (leonite, kainite, and kieserite) were included in the stable phase diagram at 25°C as shown in Fig. 1 using D'Ans data.3 The figure also shows the sequence of crystallization of the Great Salt Lake brine according to the stable diagram. In later studies by the Van't Hoff school, especially by D'Ans,3 Karsten,4 Autenrieth,5,6 attention

Jan 1, 1968

By R. J. Linney

Reserve Mining Co. produces 5 million tons of iron ore pellets per year. The finished product runs about 65.50 pct Fe, with 7.75 to 8.00 pct SiO2. Less than 12 pct arrives at the blast furnaces smaller than % in. This plant was originally designed to produce 3.75 tons annually—not 5 million. To handle the increased tonnage a fifth fine crushing line was added, and some very minor changes were made. When production can be stepped up 33 1/3 pct without extensive additions or changes—within the short period of a year—basic engineering has been sound. Operators of the plant consider it the best design job they have seen or heard of, and competent authorities have said that start-up was quicker and easier than for any previous large-scale minerals beneficiation plant. One big advantage is the elbow room allowed. Generally speaking, there is an eternal conflict between operator and designer over space. The designer has his costs per cubic foot, and he is keenly aware that if he permits the number of cubic feet to rise unnecessarily, total costs will be too high. But this was a new operation, and on a large scale. There were no proven standards to go by. Realizing this, Reserve's engineers provided ample and efficient access to machinery, allowing extra space for unanticipated situations. In three years of production this planning has paid off handsomely. Gravity Flow Saves Time and Money: This plant is on a rocky hillside, the car dumper about 270 ft above Lake Superior. In locating buildings and placing machinery the designers mad= good use of gravity flow, at considerable financial savings. In the concentrator building from rod mills to filters, gravity carries the material between all stages of the operation except one—the closed circuit pumping of concentrate between cone classifiers and ball mills. Since no machine is located directly above any other, an observer can stand on the truck ramp adjacent to the offices and laboratory and see practically every phase of activity in the huge building. The first mill section started in August 1955. After a day's run several modifications were made on auxiliary chutes and launders where spillage had occurred. This procedure was followed for the first

Jan 1, 1960

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemi-sorption. FLOTATION of quartz is of practical importance as something to be avoided in soap-floating many types of ores. Clean, unactivated quartz is not floated with fatty acids and soaps, such as oleic acid and sodium oleate, in the quantities normally used for flotation. However, data in the literature indicate that almost any multivalent cation will activate quartz if given an opportunity. Thus, a common problem is to prevent activation of quartz by the various inorganic cations inevitably present in flotation pulps. Wark and his coworkers1 have demonstrated the reversibility of the chemical reactions and adsorptions involved in the activation, depression, and collection of the common sulphide minerals. The procedure in much of their work was to bring a mineral surface to equilibrium with solutions of known pH, collector concentration, and activator concentration, and then to test the floatability of the mineral by contact-angle measurement. From the data, graphs were constructed with pH and reagent concentrations as coordinates. These graphs show fields of flotation and fields of nonflotation, separated by narrow transition regions whose locations are shown by so-called contact curves. From the shapes and locations of the contact curves, which roughly separate fields of flotation from fields of nonflotation, a quantitative understanding of the interaction of the reagents with each other and with the minerals often can be deduced. The study of quartz flotation to be described in this paper follows in broad lines the approach of Wark and coworkers. That is, pH, activator concentration, and collector concentration were varied to find equilibrium conditions of flotation and non- flotation, and the results are presented graphically by means of contact curves. However, instead of testing for floatability by measuring the contact angle on a polished surface, a simple vacuum flotation technique was developed and used. Purified oleic acid was the collector and terpineol the frother. Barium activation was studied in some detail, and exploratory studies were made of activation with calcium, aluminum, ferric iron, and stannic tin. Preparation of Materials Quartz: Large lumps of high-grade vein quartz were crushed dry in a cone crusher and rolls. The —20, +28-mesh portion was screened out and used in the subsequent steps. This material was passed through a high-intensity magnetic separator to discard iron, then leached twice with hot concentrated HCl and washed repeatedly with distilled water. The cleaned sand was then wet ground with porcelain balls in a porcelain pebble mill, deslimed repeatedly by settling and decantation to discard —800-mesh material, and again washed with hot HCl followed by distilled water. The resulting stock of quartz was stored under water. Chemical analysis gave 99.8 pct SiO2. Table I gives the size analysis of the quartz used for flotation tests. Calculations from these data, using shape factors given by Gaudin and Hukki9 indicate a specific surface of about 500 cm2 per g. Blank flotation tests in distilled water, and in water with added frother, showed the prepared quartz to be completely nonfloatable and thus indicated the absence of organic contamination. Oleic Acid: The preparation of oleic acid was based on fractional vacuum distillation of methyl oleate2,3 followed by regeneration of oleic acid, and finally fractional crystallization of oleic acid from acetone solutions at low temperatures." The pure oleic acid was stored in a refrigerator. The iodine number of the oleic acid was found to be 90.0 (theoretical 89.93). Oleic acid was used in the form of a dilute water solution of sodium oleate, after preliminary flotation tests showed no effects of form of addition and order of addition of reagents when an adequate conditioning time (that is, 30 min) was provided. Other Reagents: Sodium hydroxide solutions low in carbonate were prepared by first making 1:1

Jan 1, 1951

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemi-sorption. FLOTATION of quartz is of practical importance as something to be avoided in soap-floating many types of ores. Clean, unactivated quartz is not floated with fatty acids and soaps, such as oleic acid and sodium oleate, in the quantities normally used for flotation. However, data in the literature indicate that almost any multivalent cation will activate quartz if given an opportunity. Thus, a common problem is to prevent activation of quartz by the various inorganic cations inevitably present in flotation pulps. Wark and his coworkers1 have demonstrated the reversibility of the chemical reactions and adsorptions involved in the activation, depression, and collection of the common sulphide minerals. The procedure in much of their work was to bring a mineral surface to equilibrium with solutions of known pH, collector concentration, and activator concentration, and then to test the floatability of the mineral by contact-angle measurement. From the data, graphs were constructed with pH and reagent concentrations as coordinates. These graphs show fields of flotation and fields of nonflotation, separated by narrow transition regions whose locations are shown by so-called contact curves. From the shapes and locations of the contact curves, which roughly separate fields of flotation from fields of nonflotation, a quantitative understanding of the interaction of the reagents with each other and with the minerals often can be deduced. The study of quartz flotation to be described in this paper follows in broad lines the approach of Wark and coworkers. That is, pH, activator concentration, and collector concentration were varied to find equilibrium conditions of flotation and non- flotation, and the results are presented graphically by means of contact curves. However, instead of testing for floatability by measuring the contact angle on a polished surface, a simple vacuum flotation technique was developed and used. Purified oleic acid was the collector and terpineol the frother. Barium activation was studied in some detail, and exploratory studies were made of activation with calcium, aluminum, ferric iron, and stannic tin. Preparation of Materials Quartz: Large lumps of high-grade vein quartz were crushed dry in a cone crusher and rolls. The —20, +28-mesh portion was screened out and used in the subsequent steps. This material was passed through a high-intensity magnetic separator to discard iron, then leached twice with hot concentrated HCl and washed repeatedly with distilled water. The cleaned sand was then wet ground with porcelain balls in a porcelain pebble mill, deslimed repeatedly by settling and decantation to discard —800-mesh material, and again washed with hot HCl followed by distilled water. The resulting stock of quartz was stored under water. Chemical analysis gave 99.8 pct SiO2. Table I gives the size analysis of the quartz used for flotation tests. Calculations from these data, using shape factors given by Gaudin and Hukki9 indicate a specific surface of about 500 cm2 per g. Blank flotation tests in distilled water, and in water with added frother, showed the prepared quartz to be completely nonfloatable and thus indicated the absence of organic contamination. Oleic Acid: The preparation of oleic acid was based on fractional vacuum distillation of methyl oleate2,3 followed by regeneration of oleic acid, and finally fractional crystallization of oleic acid from acetone solutions at low temperatures." The pure oleic acid was stored in a refrigerator. The iodine number of the oleic acid was found to be 90.0 (theoretical 89.93). Oleic acid was used in the form of a dilute water solution of sodium oleate, after preliminary flotation tests showed no effects of form of addition and order of addition of reagents when an adequate conditioning time (that is, 30 min) was provided. Other Reagents: Sodium hydroxide solutions low in carbonate were prepared by first making 1:1

Jan 1, 1951

By C. H. Wayman, D. W. Fuerstenau

The effect of bubble size and concentration of certain reagents on the terminal velocity, shape, path, and drag coefficients of single air bubbles in distilled water has been investigated. Bubbles of a certain size range rise considerably faster than Reynolds number-drag coefficient relationships predict, whereas a small amount of frothing agent reduces their terminal velocity. The rapid velocity at which bubbles rise in distilled water appears to result from slip at the boundary and from circulation within the bubbles. Surface-active agents retard bubble motion through prevention of both circulation and slip at the boundary. A surface tension concentration gradient helps maintain bubble sphericity. HE gas phase is one of the indispensible in-gredients in flotation operations. Flotation depends on the collision of an air bubble and a mineral particle in a pulp and their ability to remain in contact long enough for adherence to take place. Before it is possible to understand the mechanics of air bubble-mineral particle encounter, it is necessary to learn more about the nature of bubbles themselves before collision with a mineral particle has taken place. To date, little work has been done on air bubbles in flotation systems. In 1945 Fahren-wald' presented a study of the role of frothers on air bubbles and on aeration in flotation, and recently Wark and Sutherland' discussed the work of Rosenberg- in relation to flotation. The research reported in this present article was undertaken to investigate systematically the effect of bubble size and the concentration of certain reagents on the terminal velocity, shape, path, and drag coefficient of single bubbles in distilled water. Most of the work is concerned with the effect of a-terpineol on bubble motion in water, but studies were made with potassium chloride, potassium hydroxide, potassium ethyl xanthate, and potassium amyl xanthate. Experimental Method and Materials: To obtain free bubble rise and reduce wall and surface effects, the investigators used a tank of 5½x5½x30-in. internal dimensions, filled with water to within 4 % in. of the top.' The tank was made of acrylic plastic so that the walls would have good optical properties. Distilled water, which was re-distilled in a block tin still and saturated with air before each test, was used in all the experiments. All solutions were made with reagent-grade chemicals. During each experiment the water was at room temperature, and during the course of an experiment the temperature did not fluctuate more than 0.5°C. Since the experiments were performed over a period of six months, the temperature varied between 25° and 27°C. The temperature at which each series of experiments was conducted is recorded with the data. The experimental apparatus is presented dia-grammatically in Fig. l. Individual air bubbles were generated from purified compressed air through a capillary of 0.003-cm internal diameter at a rate

Jan 1, 1959

By S. Korman, S. B. Tuwiner

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

Jan 1, 1951