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By C. E. Taylor, L. V. Fegan

With on-stream electrical-resistance measurement as a basis, Bethlehem Steel developed a system for the completely automatic control of moisture in the feed to pelletizing. This system was later successfully adapted to the sintering process, with automatic control of moisture serving to maintain the air permeability required for effective burnthrough. Control of moisture content is an important factor in the agglomeration processes by which iron ores are converted into either pellets or sinter for use in blast furnace burdens. In both pelletizing and sintering the iron ore feed must be mixed with the proper amount of water so that capillary attraction and surface tension forces can be utilized to cause the ore particles to coalesce and form larger agglomerates. The success or failure of the entire pelletizing or sintering process depends on maintaining the proper moisture. Therefore, continuous automatic control of moisture is one of the conditions for the automation and optimum use of both pelletizing and sintering processes. In pelletizing, variation in moisture can cause ball size to change and is a common cause of cycling which results in a fluctuating feed to induration. Excessive moisture can cause lumps to form at the cutter bar and contributes to the formation of buildup on the green pellet screens. Green pellets with excess moisture deform and reduce bed permeability. In sintering, accurate control of moisture is also important because moisture must be within the limits that give the bed the required permeability for good burnthrough. In response to these problems, a moisture meter, based on the principle of electrical resistance measurement, was developed. This moisture meter has been a key factor in the development of systems for providing uniform feed of concentrates in most of Bethlehem's pelletizing and sintering plants. The automatic moisture control system for pelletizing was successfully plant-tested and resulted in the following: (1) Control of moisture in the green pellets to +0.2%. (2) Up to 50% reduction of bentonite consumption. (3) Control of unscreened green pellets so that 85% of the furnace product at Cornwall was within the size range of -5/8 +3/8 in. (4) Improved quality of product with a tumble test of 94 and a "Q" index of 90 as the result of feeding a consistently uniform feed to a shaft furnace. In sintering, where sinter-mix temperature change and variable feed could cause errors in moisture measurement, the system was designed to maintain the correct moisture under virtually all plant operating conditions. It was successfully plant-tested and is now in use in most of Bethlehem's sintering plants. The system achieves: (1) Control of moisture within the desired limits for optimum production. (2) Continuous measurement and control of moisture even with intermittent feeding of mix, so that attention by the operator is seldom required. The two automatic control systems provide smoother startup, minimize upsets and contribute to the consistent control of process variables in pelletizing and sintering. MOISTURE-CONTROL SYSTEM FOR PELLETIZING Fig. 1 shows the general arrangement of the system for automatic control of moisture in pelletizing.1 In this system a bin level controller, using a single probe which regulates filter speed, maintains a uniform low concentrate level in the bin.' The moisture meter, comprised of an analyzer and control components, continuously measures the moisture in the bed and adds water to adjust to the desired level. The meter is placed in an open-loop (feed-forward) control, i.e., water addition occurs downstream from the probes. Filter vacuum must be set so as to produce a filtercake moisture which is below the optimum for pelletizing. Bentonite is added after the

Jan 1, 1969

By J. W. Axelson, J. N. S. Kwong, J. T. Adams, E. L. Pire, J. F. Johnson

CRUSHING has always been a major operation in the chemical and metallurgical industries, yet little is known about the theory of crushing, and today, the design of crushers is still based almost entirely on empirical knowledge and accumulated practical experience. In view of the increasing national need for the economic working of poorer grades of ores, the lack of a fundamental understanding of this unit operation hardly presents a satisfactory situation. Basic investigations of crushing have been concerned mainly with three phases of the problem, 1—the mechanism of the fracture process itself, 2— the particle size distribution of the crushed product, and 3—the relationship between the energy input and the amount of new surface produced. Probably adequate information on these phases will be required for a comprehensive understanding of the process of crushing. A bibliography covering these phases will be given as well as a review of the work done in recent years at the University of Minnesota on the relationship between the energy input for crushing and the amount of new surface produced. Fracture of Solids This phase of the crushing problem is the most fundamental because it is concerned with the actual mechanism of fracture. The main problems in a study of the fracture process are concerned with the questions of why and how fracture occurs and why there is such a discrepancy between the actual and theoretical energies needed in a fracture process. In considering the how of fracture, Poncelet40 carried Out experiments On the crushing of glass under compression. Actual photographs were taken of the glass plates in various stages of fracture-From these experiments Poncelet was able to postulate a probable mechanism of fracture. Probably because of the quantitative nature of the problem, considerable effort has been expended to find an explanation for the low tensile strength of materials and the conversely high energy for crushing. Tensile strengths are often only 1/500 of the theoretical while the energy for crushing is usually at least 500 times the theoretical. A review of the work to explain the low strength of glass is given by Weyl,52 in which the work of Griffith," Joffe,32 and Powell and Preston " are discussed. In a recent paper Seitz46 holds that the presence of flaws and fissures, as originally expounded by Griffith, is the most probable reason for the low strength of metals. Since a mass of data has been accumulated on the fracture strength of materials, several investigators have attempted to correlate this data in the form of an equation. Glathart and Preston," Taylor,40 Ponce-let,40 Machlin and Norwich," and Frederickson and Eyring,20 have all considered this problem, and the final relationship has always been that the stress for fracture is proportional to the logarithm of time. Murgatroyd30 arrived at a very similar relationship by a different method. The study of the fracture process has used many methods. Among some of these are the thermody-namic approach used by Furth21 which was based on the work of BornxL and later used by Saibel,43 and the statistical method applied by Fisher and Hollo-mon.10 In recent years metallurgists have concerned themselves with the problem of brittle .fracture as opposed to the usual plastic failure of metals. The great amount of interest shown is reflected by the number of contributors to the recent symposium on the fracture of metals.' Repeated failures of welded ships during World War II have accelerated these Studies.31,50,34 The shifting of thought from an emphasis on the

Jan 1, 1952

By C. C. Harris

An analytical solution of the integro-differential equation of batch grinding is proposed as a first appoximation to the xYt surface. The size distribution equation (xY plane) is a three parameter equation previously suggested. Possible refinements are outlined. Comminution kinetics research is presently following several related courses. Some are equivalent, in that they represent different ways of viewing the same facet of the problem, though the entire endeavor cannot be described as being completely integrated. The purpose of this note is to propose and discuss a framework into which the various approaches might be fitted. This must clearly be founded upon an analysis of the xYt surface.' (See also Fig. 1 of Reference 26.) The investigations referred to above comprise mainly researches into the following: (1) The kinetics of the creation of the finer sizes during relatively small grinding times. This is the Fahrenwald-Arbiter-Bhrany 2-3 approach which leads to equations equivalent to a special case of the charles4 relationship. It emerges that initial production of fines follow essentially zero order* kinetics. (2) Disappearance kinetics of the coarser fractions. This appears to be a first order relationship.5-6 (3) Prediction of the xYt surface via the selection and breakage functions utilizing computer-programmed or approximate solutions of the integro-differential equation of batch grinding.5,8-17 A recent discussion by Mika et al.5 references the relevant literature up to about 1966. In addition to these main-stream researches there is some work on the minor topic of extremely long grinding time considered as a limiting case.18-19 The field is supported by investigations into the size distribution of fragments produced by breakage processes, some of this work being most valuably statistical-mechanical in nature.20-25 Better fitting equations, however, are obtained by statistical and curve-fitting methods.26 This

Jan 1, 1969

By A. W. Asman

The problem of balanced design in a belt conveyor has been under constant study. Increased use of this method of transport in recent years has stimulated investigations. Conveyor belt, power transmission and other machinery should be adequate to perform the required service. All components of the belt conveyor should work as an efficient unit moving its load continuously between terminals. As part of a long term research project on belt conveyors, Hewitt-Robins established in 1954 a 2-year grant at Pennsylvania State University. Under terms of the grant, certain fundamental phenomena were to be studied. This work evolved a new concept in belt conveyor power determination based on the principle of belt tension. The Penn State contract furnished the foundation for later more detailed studies. From these most recent investigations, more specific factors were defined. In the final analysis, all of the data collected has been worthwhile in belt conveyor design. Units now operate over long distances at speeds of over 800 fpm. Long high lifts are commonplace. Where vertical or inclined shafts were the rule, belt conveyors receive first consideration. Although belt conveyors have moved forward rapidly, the future will see a tremendous gain in their use. The ideal test belt conveyor would be several thousand feet in length and have a maximum tension of at least 70,000 lb., Also, this conveyor should be horizontai and have a variable feed rate of more than one material. Data from such a field belt conveyor combined with laboratory information would permit accurate analysis. Since this type of conveyor was not available, many field locations were visited: An alternate would have been to erect a closed circuit system in the laboratory. Either method had limitations and the first was selected as having the best possibilities. Manufacturers of conveyor belting and machinery basically have been using the same horsepower formula. As more information became available from field installations, factors in the basic formula were modified. Practically all conveyors are designed today by adding together the horsepowers required to move the empty belt, move the load horizontally, and elevate the load. In selecting size of drive motor, the first two elements are relatively unimportant if the last is large. However, it is the combination of

Jan 1, 1961

By C. D. Thompson, D. C. Violetta, C. A. Czako

The mechanics of the continuous sintering process are briefly reviewed and the application of this process for the beneficiation of principal minerals used in making cement is shown. Pilot plant tests using pyrobeneficiation techniques were conducted with mixtures of limestone and shale in various ratios. Results are illustrated and show the percentage removal of carbon, carbonates, sulfur, alkalies, and volatiles as a function of the limestone-shale ratio. Tests with varying quantities of solid fwls, added to mixtures containing limestone and shale in a 4 to 1 ratio, were made to attain optimum elimination of the volatile constituents. Experimental results show that 90 pct of these constituents were removed during the calcination sintem'ng process. The advantages of using the beneficiated material for making cement are discussed with regard to prepam'ng, handling, and clinkering operations. A sintering process for treating raw materials or kiln waste products which contain high amounts of alkaline materials is described. Tests were performed with a high alkaline waste dust mixed in a 1 to 1 ratio with raw cement meal. The data from these tests and others using different halides for aiding alkali removal are evaluated by the alkali content of the cement clinker. Test results showed that the Dwight-Lloyd process could be applied with facility in beneficiating raw materials for cement making. Substantial quantities of deleterious volutile constituents were removed by calcining and sintering techniques. The continuous sintering process has been applied for a number of years as a means of beneficiating minerals for subsequent conversion to merchantable products. The process was developed and first used for nonferrous metallurgical applicatfons. However, it has risen to prominence for ferrous metallurgical applications where last year in the U.S. and Canada nearly 40 million tons of iron ore were treated by grate firing processes. Through the years of process development many investigators have realized the versatility of the sintering process and have directed experiments toward the use of it for applications with nonmetallic minerals. These have included calcining limestone, carbonizing coal, expanding shale, and burning cement raw meal. Today commercial sintering plants produce cement clinker from mixtures of limestone and shale, lightweight aggregate from shale or clay, and calcined products from phosphate shale. The mechanics of the downdraft sintering process have been presented by many authors. In general practice the materials to be agglomated are mixed as moistened fines with a solid fuel and this mixture is spread uniformly on the grates of the sintering machine. The charge is then momentarily directed under an ignition furnace. After ignition a continuous downdraft of air promotes combustion of the fuel and through heat transfer particles become partially fused into a coherent mass. This is cooled from continued draft induced after combustion. Many investigators have shown that distinct zones are formed during the process of sintering iron ores. At the midway point of the sintering process these are prominent and are, respectively, from the top of the bed to the bottom, comprised of a cooling zone of sintered material, a combustion zone of sinter formation, a calcination zone, a drying zone, and a zone of unreacted materials. Completion of sintering occurs near the discharge end of the machine where the firing zone approaches either the grates or grate dressing material called hearth layer. Here the peak draft temperature usually can be noted in the windboxes. The zonal pattern has been illustrated with the use of graphs which show thermal gradients as they progress through the bed of materials during sintering. Fig. 1 shows, in simplified form, thermal gradients existing when clinkering cement on a sintering machine. The thermal gradients illustrated are for periods of initiation, midway and completion of the clinkering process. The occurrence of these zones has importance for the beneficiation of cement raw materials. A research program was performed at the Dwight-

Jan 1, 1961

By R. J. Brison, W. C. Bleimeister

The International Salt Company has long been interested in finding an efficient process for the removal of impurities from rock salt, and particularly from the rock salt produced at the Detroit mine. Methods which have been used previously at the Detroit mine included both hand sorting and differential crushing in a Bradford breaker with rejection of the oversize. Neither of these methods was adequately effective. The development of a new process for removal of impurities from rock salt has been described in the preceding article." Basically, the process consists of exposure of crude rock salt to radiant heating which selectively heats the impure particles. Subsequently the heated particles are removed by adhesion to a heat sensitive-coating on a conveyor belt. This method was found to be effective in improving both the quality and appearance of rock salt by the rejection of the dark, impure particles. The first commercial unit employing the new thermoadhesive method of dry separation of minerals was installed at the Detroit mine of International Salt Company in Detroit, Mich. (Fig. l). This unit was first put into operation in March 1957. The purpose of the Detroit installation was 1) to serve as a pilot plant for evaluating the process on a commercial scale and 2) to produce marketable quantities of rock salt of improved quality and ap- pearance. The design of this plant was based on the results and information obtained from a 1-tph pilot plant at Battelle Memorial Institute. DESCRIPTION OF PLANT Location: The separation plant was installed underground near the crushing and screening plant, for this was the most suitable location from the standpoint of material handling. Furthermore, ample space was availabe at this location. Because the process requires a relatively constant temperature, the uniform year-around temperature of the mine (57°F) gave an added advantage to an underground operation. Feeding: The rock salt is conveyed to a surge bin, from which it is fed to the heating unit by a conveyor. The feed-rate control consists simply of an adjustable gate at the bottom of the surge bin. The plant may be fed either No. 1 grade of rock salt (0.279 to 0.375 in.) or No. 2 grade (0.375 to 0.500 in.) with a maximum input capacity of 500 tons per 15-hour day. Heating Unit: The heating unit (Fig. 2) consists of a panel of infrared lamps suspended inside a low-speed revolving screen. The radiant-heat panel, which is curved for most effective utilization of the heat, holds 240 infrared lamps of standard industrial type, each rated at 500 watts. The position of the panel is adjustable so that it can be placed centrally over and to within 6 in. of the salt bed on the screen. The entire heat panel is suspended by rollers in ¦ R. J. Brison and O. F. Tangel: Development of a Thermo-adhesive Method for Dry Separation of Minerals.

Jan 1, 1961

By R. M. Hays

At the present time, oxidized taconite and semitaconite are not processed on the Mesabi Range except in vely limited tonnages. However, it is estimated that by 1990 shipments from the Range will approach 75 million tons of iron ore per year. To achieve this and future production, either additional magnetic taconite reserves must be proven or oxidized taconite and semitaconite must be processed. One beneficiation process which uses high-intensity magnetic separation and flotation for concentrating oxidized taconite and semitaconite is described. Detailed metallurgical results are given and an economic evaluation of the process is presented with comparison to other processes. The results of the present study indicate that the high-intensity magnetic separation and flotation process appears nearer commercialization for the processing of Mesabi Range semitaconite. The disadvantages of the process are nearer solution and are being solved by well-defined programs. During the twentieth century, the Mesabi Range in northeastern Minnesota has been the main source of iron ore for American steel mills. The Mesabi Range, which is about 120 miles long and up to 3 miles wide, is the pre-glacial outcrop of a group of sedimentary iron-bearing rocks known as the Biwabik formation. The formation, which is 200 to 650 ft thick and dips 50 to 10° to the southeast, is divided into four stratigraphic units. From top to bottom, these are the Upper Slaty, Upper Cherty, Lower Slaty, and Lower Cherty. Since its deposition in Precambrian time, most of the original Biwabik formation has been subject to alteration. Thus, although the iron content of the original formation was quite uniform, major variations in mineral composition now occur both laterally along the length of the formation and vertically within its units. Details of the mineralogical variations in the Biwabik formation have been described by Wolff,' Gruner, 2 and White.3 The unaltered portion of the Biwabik formation includes the magnetite taconites which are being commercially processed by Reserve Mining, Erie Mining, Eveleth Taconite, National Steel, Butler Taconite, and U.S. Steel. These six operations have a capacity of 32 million tons of iron ore pellets per year and have produced approximately 170milion tons of pellets since 1956. Capacity for another 30 million tons of pellets per year is being considered by the industry for the next ten years. This increase will be obtained by constructing new plants and by expanding present operations. Thus, by ! 978, it is estimated that approximately 60 million tons of iron ore vellets will be produced each vear from the magnetite taconites of the Biwabik formation. The portions of the Biwabik formation subjected to intense alteration resulted in direct shipping, wash, and heavy-media ores. Since 1892, 1.8 billion tons of direct shipping ore have been shipped from the Mesabi Range. This ore has been the major source of iron ore for American steel mills. In 1967, about 11 million tons of direct shipping ore and 13 million tons of wash, jig, heavy-media, and other gravity concentrates were shipped from the Mesabi Range.4 These products generally have poorer physical and chemical properties than the iron pellets produced from magnetic taconite. It is predicted that the shipment of these materials will decrease from 24 million tons in 1967 to 6 to 8 million tons in 1975.' This decrease will

Jan 1, 1970

By N. J. Sather

A detailed description is given of Bunker Hill's concentration process employed at the company's lead-zinc property in the Coeur d'Alene district, Idaho. The plant is equipped to process 3000 tpd of Pb-Zn ore. The history of the Bunker Hill mine dates back to Aug. 26, 1885, when Noah S. Kellogg found the outcrop of the Bunker Hill orebody on the hillside of Milo Gulch above the present town of Wardner, Idaho. A small concentrator was erected in 1886, and before the railroad came, the mine output was transported by mule teams to the head of navigation on Coeur d'Alene Lake. After ten years of production, an interest was purchased in the Tacoma smelter at Tacoma, Wash., to which the entire output was shipped. In 1916 and 1917, a modern lead smelting plant and refinery was erected, and a modern electrolytic zinc reduction plant was built in 1927-1928 by the Sullivan Mining Co. which was 50 pct owned (and now totally owned) by the Bunker Hill Co. The first concentrator, a 100-tpd capacity plant, was built in 1886 near the Reed Tunnel in Wardner. The treatment consisted simply of crushing ore to 2 in. and then to 1/2 in. through rolls in closed circuit with a bucket elevator and trommel screen. The undersize was further sized in trommels to 10 mm and 3 mm. Oversize from the latter was treated in Hartz jigs. The 3 mm undersize was classified in crude hydraulic classifiers furnishing sands to sand jigs, the overflow being dewatered in a long V tank which supplied sands to a Cornish buddle and to Frue vanners. The concentrates were of excellent grade, old records showing assays of 69 pct lead and 29 oz. of silver per ton. Tailings were high inasmuch as the provision for treating the slimes was inadequate. Since the erection of the first mill, the Bunker Hill plant has gone through many changes and improvements. In succeeding years, the flowsheet was enlarged somewhat with Huntington and Chilean mills, vanners, tables, ball mills, and flotation cells being progressively added. In the transition period to the present day practice, the object was not only to improve the equipment for better mechanical performance, but to add new equipment whenever practicable so as to improve both grade and recovery of the product at a lower cost. For the modern plant and flow-sheet, see Figs. 1 and 2. The history of flotation at Bunker Hill dates back to 1913 when experimental work was done in a cell in which the pulp was circulated through a centrifugal pump set outside the cell and discharged against a baffle in the cell. The cell was built with a spitzkasten, and the quiet zone in the latter tended to let the froth break down before it got to the overflow weir. To remedy this situation, a belt drag with light metal flights was arranged for drawing off the froth over a ramp in each cell. Recovery of lead was not as high as present day practice, but was so much better than the vanners that the cells were multiplied gradually to eight cell units. The mill men called them mud hens. Flow between the cells was controlled by a weir, which determined the pulp level in each cell. These early machines were called Bunker Hill machines. The tailings were later treated in pneumatic scavenger flotation machines in a series, producing concentrates for retreatment in the Bunker Hill cleaners as well as the final tailings. This system of flotation served with minor changes until the Fahrenwald machine was developed at Bunker Hill during the period from 1921 to 1925. In 1925, the Fahrenwald flotation machine gradually replaced the Bunker Hill machine, although the Hearing pneumatic scavengers for retreatment of the tailings from the Fahrenwald machines were retained in the lead section until 1928. The use of the Hearing pneumatic machine for cleaning the rougher concentrates was begun in 1924. In 1941, plans were made for remodeling and modernizing the concentrator, the goal being a modern concentrator employing the latest equipment and milling practices. The old wooden building was to be replaced by a steel and concrete structure. Shortly after preliminary work got under way, the United States entered World War II and the demands for metals increased. To meet this requirement, the management decided to increase the capacity of the concentrator to 1800 tpd. This was accomplished in record time by the construction and operation of a sink-float plant. Installation of this plant was the first step in the general plan for modernization. In 1947 the concentrator (including the sink-float plant) was enlarged to handle 3000 tpd. Milling operations today are at the rate of 2300 tpd, five days per week. During 1959, 453,000 tons of ore was beneficiated. The Bunker Hill concentrator is in three sections: the crushing plant, a 3000-ton capacity storage bin, and the concentrator building. The crushing plant is of wood design with a gunnite coating on the exterior.

Jan 1, 1961

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

With either sulfonate or oleate as collector, quartz responds to flotation with moderate additions of calcium only at moderately high pH, where some portion of the activator has hydrolyzed to caOH+ . Calculations of the concentrations of various ionic and precipitated species of calcium and collectors suggest that the products of [(CaOH+) (RSO3)] and [(CaOH+)(01-)] determine whether flotation is obtained under specific conditions. Ion products on the order of 10-12 were calculated for both the sulfonate and oleate systems. The activating effect of calcium ion in nonmetallic flotation systems is of considerable interest because of the normal presence of calcium in natural water. As a result, this phenomenon has received quite some attention in the past. Kraeber and Boppel1 showed that quartz could be activated by calcium above pH 10 with sulfonate as collector. The feasibility of selectively separating quartz from hematite with calcium activation at relatively high pH was demonstrated by Clemmer, Clemmons, Rampacek, Williams, and stacy.2 Cooke and Digre3 showed with a bubble pick-up method that the minimum quantity of calcium ion required as activator for complete pick-up of particles occurs at pH 11.5 for an addition of 20 mg per liter sodium oleate. They also showed that larger additions of calcium (10-fold increase per unit decrease of pH) must be added for complete bubble pick-up as the pH is reduced. Schuhmann and Prakash,4 using a vacuum flotation technique, found that quartz could be floated with moderate additions of calcium chloride and oleic acid at neutral pH, providing the metal ion was present in stoichiometric excess over the quantity needed to form the normal soap with oleic acid. They also reported that calcium will function as an activator only in basic media. More recently, Eigeles and volova5 have shown that essentially complete flotation of quartz is obtained with 6 x 10-4 mole per liter calcium chloride and 1.7 x 10-5 mole per liter sodium oleate at pH 11.6. while no flotation is obtained at about pH 10.9 and below. The importance of adsorption of activator and collector at the air-liquid interface is also demonstrated in these systems. The important role that metal ion hydrolysis assumes in quartz activation systems was also demonstrated recently.6-8 A detailed investigation of metal activation in sulfonate flotation of quartz was undertaken in one system7 and yielded a number of interesting and important observations. Quantification of the data of this system7 to the extent desired was not possible, though, because certain species could neither be ignored nor accounted for accurately. These difficulties can be circumvented when calcium is involved as activator. This detailed analysis was undertaken to obtain a more quantitative explanation of calcium and metal ion activation in quartz flotation. EXPERIMENTAL MATERIALS AND METHODS Sodium alkyl aryl sulfonate9 mol wt 450, and pure potassium oleate were used as collectors. All other reagents used were reagent grade in quality, i.e., n-amyl alcohol as frother, KOH for pH adjustment, and calcium chloride. Conductivity water, made by passing distilled water through an ion exchange column, was used in the investigation. Quartz was prepared by leaching the sized sample (48 x 150 mesh) with HC1 until no iron could be detected in the leach liquor. The experimental equipment and procedure were the same as that described previously.6,10 EXPERIMENTAL RESULTS As the presence of precipitates was noted in all of the systems to which ca++ and collector were added, experiments were undertaken to determine the solubility products of calcium sulfonate and calcium oleate using a nephelometer. With this technique, collector is titrated into a known solution, which in this case was 5 x 10-5 mole per liter CaCl2 at pH 5.5. Upon precipitation of the calcium-col lector salt, e.g., calcium oleate, light is scattered and detected

Jan 1, 1967

By A. F. Colombo, R. T. Sorensen, D. W. Frommer

An analytical method has been developed and used in batch and continuous tests to provide initial insights into the effect of soluble calcium ion in anionic flotation of silica from iron ores. A deficiency or excess of calcium ion in pulps subjected to this type of flotation produces an impairment in flotation efficiency. Data developed in the investigation are from a single ore, but show that total calcium requirement is met from water, ore and reagent sources. In batch studies, starch additions were shown to exert no influence on the calcium level, but fatty acid and sodium hydroxide had subtractive effects. The investigation was next extended to continuous flotation testing where the general observations from batch testing were confirmed. Factorially designed experiments were performed showing that the relationship of calcium additions to flotation response was not a simple one, but tended to shift to higher additions with increased fatty acid. Nevertheless, the range of soluble calcium levels measured remained fairly constant in areas of good flotation response. It is concluded that techniques developed should be useful in removing empiricism from flotation operations in this and other systems requiring activation for success. Further development should aid in process control both through methods outlined and through investigations aimed at studying the effects of other soluble reagents. Of the major mineral commodities, iron ore was one of the last to succumb to flotation processing. Commenting on this fact, one of the speakers 1 at the convention celebrating the 50th Anniversary of flotation in the U.S.A., said ". . . in 1951 froth flotation as a means of beneficiating nonferrous minerals was a might giant and could be characterized by 'you name it, we float' - but not iron ore". This tardiness in applying flotation to iron ores was the result of a number of factors, including economics and ready availability of direct shipping ores. Even now, when the need for beneficiation can no longer be denied, proposals for flotation of iron ores are apt to be viewed with some misgivings. This disquiet results from knowledge that wide variability, which would necessitate corresponding adjustments in reagents, often exist within even a single ore deposit. Coupled with ore variability is the fact that, at present, there is no way of quickly and precisely anticipating the direction and magnitude of reagent changes required for uninterrupted production of acceptable concentrates. Despite the long delay, iron ore flotation is now practiced commercially, although the chief applications2-4 are to specular hematites having relatively coarse liberation and low slime content. In addition, the application of flotation in upgrading magnetic concentrates seems assured. However, these flotation methods do not appear applicable to substantial tonnages of fine grained, earthy, and frequently more complex semitaconites and nonmagnetic taconites. The U.S. Bureau of Mines, in devoting its attention to the semitaconite and nonmagnetic taconite class of materials, has demonstrated that anionic flotation of calcium-activated silica can be utilized with an expected high frequency of success.' Even so, it has been recognized that the flotation method is a complex one, and that the reagent suite must be carefully tailored to the feed. These conditions would be expected to offer difficulties in treating ores with the wide range of variability previously discussed. Accordingly, research was initiated to develop methods to remove much of the empiricism from the flotation operation by measuring the reagent variables within the pulp for a correlation with flotation response. A simple illustration of the technique sought would be in the well established use of pH meters to provide feedback control to reagent feeders and thereby limit acidity or alkalinity to a range producing optimum results. The first efforts in this area have been concerned with measurements of soluble calcium ion. Bureau experience in both laboratory and pilot plant over the past five years has proved that the amount of calcium

Jan 1, 1965

By R. C. Weed

LEACHING in place at Cananea began in the 1920's on a limited scale. The first plants were small wooden boxes located underground in the Capote and Oversight mines, and output was low. Scrap iron was used as precipitant. Later several other small plants were started up. Leaching was limited entirely to underground areas, and production remained low. The Veta precipitating plant, the first on a larger scale, was put into operation in the late 1920's. After a period of intermittent service it was rebuilt in 1946 and has since been running continuously. In 1943 work on the Colorada open pit was started. This produced 40 million tons of waste, which averaged 0.20 pct capper. When these dumps had aged a few years it was decided to leach them on a large scale, and after considerable research and investigation a leaching system of pumps and pipelines was laid out and installed. At the same time the Ron-quillo precipitation plant was built to treat the copper solutions from the pit dumps. This plant started operation in March 1953. Leaching areas consist of the Colorada pit dumps and various mined out and caved underground stopes. The principal mineral, chalcocite, leaches readily in a weak sulfuric acid solution containing ferric iron. Enough pyrite is contained in the areas to produce an acid-ferric iron solution strong enough to dissolve the copper. Each area is studied to determine the best method of distributing the water over it and the best method of collecting the copper-bearing solutions after leaching has taken place. Water is generally distributed by means of ponds, ditches, and sprays. Spraying has proved the best method of spreading the water evenly; however, in a dry climate some water is lost through evaporation. After passing through the leachable material the solution is collected underground by dams and bulkheads and on surface by dams. An effort is made to give the maximum storage capacity practicable.

Jan 1, 1957

By James E. Lawver, James L. Wright

This paper describes the design of a cell used to measure the electrical conductivity, or the reciprocal resistivity, of granular materials. It also establishes a quantitative relationship between the electrical conductivity and the electrical separation of selected mineral samples. It is estimated' that the free world produces about 13 million tons of mineral concentrates a year by electrostatic or electrodynamic methods. At least 95% of these concentrates are produced by corona discharge type machines that separate minerals by exploiting the differences in their electrical conductivity. But, what is meant by the electrical conductivity of a mineral? How large a difference in the electrical conductivity of two mineral species is necessary to effect a separation? Just what is the role of conductivity in electrical separations with regard to capacity and selectivity? There is a considerable amount of information in the literature on both the mechanisms of conduction and the values of the electrical conductivity of pure single-crystal mineral specimens. In addition, some insight into the role of mineral conductivity can be obtained by referring to a Bureau of Mines publication' in which 95 minerals have been grouped as conductors or insulators as a function of temperature. The classification was made on the basis of the percent of weight that reported as "conductors" in an electrical separator. But, as is often the case in the mineral industry, practical application leads theory — and there is little useful information in the literature giving any kind of quantitative relationship between electrical conductivity and electrical separations. The problem is a bit complex because electrical conductivity per se — which can be rather easily measured on large, single grains — is not sufficiently informative to permit quantitative predictions of the separability of the mineral system. Practical "high-tension"* separations are made at such feed rates that the band of minerals on the rotor is multilayered in thickness. Thus, the effective conductivities of minerals as seen by a high-tension machine include particle-to-particle contact resistance, which is often a controlling factor in making a separation. This is particularly true when the mineral grains are coated with slime or with a dielectric surface film. From a high-tension separation viewpoint, one is actually concerned with the time required for corona-charged grains to lose enough charge so that the centrifugal force acting on the particles exceeds the image force holding the grains to the rotor. This calculation requires a knowledge of the effective mineral conductivity s. (The conductivity of the metal rotor is sufficiently greater than that of the minerals so that it need not be considered in charge decay calculations.) The charge density at time t can be approximated by the relaxation equation.3 where t 5 charge density at time t r, - charge density when t = 0 cr e total electrical conductivity £ = permittivity The time required for the charge density to fall to l/e of its initial value is called the relaxation time and is numerically equal to €/n. AS defined here, u is the conductivity of an element of homogeneous grains making contact with each other and with a grounded rotor. This paper deals with the design of a conductivity cell used to measure a, or its reciprocal p, the electrical resistivity. CELL DESIGN Initial test work was performed with a cell recommended by ASTM.4 Since the reproducibility with the ASTM cell was not satisfactory, it was decided to

Jan 1, 1970

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By A. S. Joy

Vacuum flotation tests carried out according to the method of Schuhmann and Prakash1 have shown that the acid limiting edge of the flotation area for a Brazilian hematite, in the presence of dodecylamine, is a straight line of slope approximately unity. The flotation tests and the structure, surface ionization, and electrokinetic properties of this hematite have been fully discussed elsewhere. 2 The numbers of adsorbed H+, and OH- ions and collector molecules have been determined in the range of critical conditions defining the acid limit of the flotation area. The total amine adsorption is shown in Fig. 1 as a set of contours superimposed on the straight part of the flotation area. These were determined by spectro-photometric estimation of the complex formed between residual amine and the anionic dye Orange II.3 The ionic strength of the solutions was kept constant and at a value in sufficient excess of the amine concentration to ensure that the diffuse part of the double layer was populated predominantly by indifferent ions. Adsorption of H+ and OH- in the absence of amine was found by potentiometric acid-base titration in the presence of hematite.4 The form of the adsorption-titration curves suggested that the hematite surface behaves in an amphoteric manner, and that positively charged sites arise through loss of a hydroxyl ion, and negatively charged sites by loss of a hydrogen ion. On this basis, curves 8 and 9, shown in Fig. 2, were calculated from the titration curves. These show the variation of the numbers of positive (s +) and negative (s—) sites with pH. To distinguish between ionic and other modes of adsorption, zeta-potential determinations have also been made using the streaming potential technique. The influence of amine on the zeta-potential of the mineral over a wide pH range is shown in Fig. 3. The change in charge ?s, due to adsorption of amine ions in the Stern layer, was calculated by applying the Gouy-Chapman theory, as modified by Stem, to the results shown in Fig. 3. The validity of this treatment has been shown by Ottewill and Jaycock.5 The three full lines, Nos. (1) through (3), included in Fig. 2 show the relationship between ?s and pH at three amine concentrations. The general conclusion from the zeta-potential curves is that adsorption of amine causes a reduction in the negative charge on the surface. Comparison of the values shown in curves (1) through (3) of Fig. 2 with the - curve shows that the magnitude of the reduction is comparable with that of s - at any given

Jan 1, 1964

By J. M. Steininger

The mechanism of flotation of sphalerite with sulthydryl compounds of different acidities has been interpreted in terms of chemisorption of unionized collector molecules on hydroxylated zinc surface sites. At a constant collector concentration of 1.5 x 10 -4 M, the upper pH limit of flotation has been found to be directly and closely related to the negative logarithm of the collector acidity constant, pKa The experimental results indicate that flotation is prevented when about 95% of the collector is ionized. At low pH, collector adsorption decreases because of the reduction in number of hydroxylated zinc surface sites. These two effects explain why sphalerite does not float with the more acidic sulfhydryl compounds. Typical anionic collectors for sulfide minerals are short chain xanthates or dithiophosphates which are nearly completely ionized in the generally alka- line flotation circuits.1,2 Flotation mechanisms have therefore been usually interpreted in terms of ionic collector adsorption and competition with ionic depressants of the same sign as the collector. A few attempts have been made, however, to consider adsorption of neutral collector molecules. Contact angle measurements obtained by Wark and cox3 have been interpreted by Cook and Nixon4 as representing adsorption of the free acid molecules on sulfide minerals such as pyrite, galena, chalcopyrite, and sphalerite. Cook and Nixon tried to show that the free acid concentration of the collector was substantially constant along the curves relating collector addition and minimum pH for bubble adherence. However, in a review of the subject,5 Sutherland presented new data on the flotation of pyrite with mercaptobenzothiazole which did not support the hypothesis of constant concentration of either collector ions or collector free molecules. He further showed that, with the assumption of constant occupation of surface sites, both theories lead to relations between pH and collector addition having similar mathematical expressions. In fact, with the more likely assumption of variable occupation of surface sites which results in more complex relationships, it appears impossible to

Jan 1, 1968

By M. D. Hassialis, C. G. Myer

THE nature of a collector-coated mineral surface has been the subject of some experimentation and much speculation. Various aspects of the problem have been isolated and studied; it is probable, however, that there exist aspects as yet unrecognized, which are needed for a complete description of the system. The recognized elements of the problem are: 1—The nature of the chemical units comprising the film and the nature of the forces holding these chemical units to the mineral surface; 2—The orientation of the chemical units at the surface; and 3—The disposition of the collector units at the mineral surface. This report touches upon all three aspects of the broad problem; it deals, however, with a new and heretofore unrecognized element—the mobility of the collector units in the two dimensions of the solid-liquid interface. Regardless of the viewpoint adopted as to whether the chemical units comprising the collector film are collector ions or molecules, held either by chemical or physical forces, the writers on flotation theory are in apparent agreement as to the orientation of the chemical units at the mineral surface. Wark and Cox' showed that the contact angle at a stable collector-coated surface in a constant chemical environment is independent of the nature of the mineral and of any resurfacing agent; it depends solely upon the collector used. They showed further that the value of the contact angle increases with the number of carbon atoms in a homologous series of compounds. Taggart, Taylor, and Knoll2 postulated that the collector ion is oriented with the chemically reactive, polar end toward the mineral surface and the non-polar, hydrocarbon-like end toward the liquid phase. In support thereof they cited the abstraction of p-p1 dihydroxydiphenylthiourea and of glycol xanthate8 by galena without a concomitant increase in the contact angle above that of clean galena. It was argued that in these cases the water-repellent layer which normally would be presented to the liquid phase owing to the orientation of the collector ions is overlain by a layer of water-avid groups, the hydroxyl groups. Implicit is the analogous orientation exhibited by a monomolecular layer of an insoluble substance at an air-water interface. It should be pointed out that these facts also are consistent with the concept that the collector units be flat on the surface of the mineral and that the response of the surface to an air bubble is determined by the ratio of the area covered by water-avid groups to the area covered by water-repellent groups; the areas being weighted according to the relative watef-avidity or water-repellency of the groups involved. The disposition of the chemical units of the collector film at the mineral surface has not received the attention it merits. Taggart4 concluded from the observed stoichiometry of the ions entering and leaving a surface that the stoichiometry is obeyed in detail at the surface of the mineral. Thus, in the case of xanthate-coated galena, pairs of xanthate ions are attached to surface lead ions. In the case of chemisorption, Gaudin and Preller postulated an agreement between the atoms or ions of the substratum (the mineral surface) with those of the superstratum (collector film); in the case of the physical adsorption their position is not clear. In one place they state "the first type (physical adsorption) does not require any agreement between the atoms or ions of the substratum with that of the superstratum" and in another place that "It seems, in fact, as though some slight influence of the substratum is formed even in physical adsorption." The present work shows that in some cases of chemisorption the chemical units of the collector film are in constant surface-bound motion and may not be assigned to any site of the surface, hence there is no permanent association between the chemical units of the film and the atoms or ions of the solid. The physical property of the collector film that is of paramount importance is its surface energy. Although this property is not susceptible to direct experimental determination, the related contact angle is. Much of the knowledge of the properties of a collector film derives (by deductive and inductive reasoning) from the behavior of an air bubble thereat

Jan 1, 1952

By J. H. Polhems, R. B. Brackin, D. B. Grove

THE heavy media separation process plays an outstanding role in the concentration of 4000 tons of zinc ore per day at the Mascot mill of the American Zinc Co. of Tennessee. Of the total tonnage, 72 pct is treated in the heavy media separation plant to reject 56 pct of the ore as a coarse tailing, which has a ready market. Concentrates from this separation are beneficiated further by jigging and flotation. Approximately 25 pct of the total zinc concentrate production is made in the jig mill. Jig tailings are ground and pumped to the flotation circuit where the balance of the production is made. Fig. 1 shows a generalized flowsheet of the mill. The Mascot ore is a lead-free, honey-colored sphalerite in dolomitic limestone, with lesser amounts of chert and some pyrite. A mineralogical analysis is given in Table I. After 10 years of successful operation with galena medium and treatment of nearly 10,000,000 tons of ore, a decision to convert to ferrosilicon was made early in 1948 because of the increasing price of galena and consequent high operating costs. The conversion was made on Nov. 6, 1948, and the results obtained since that time have shown remarkable improvement over those made with galena. The Table I. Mineralogical Analysis of Mill Feed, Pct Calcium carbonate 49.5 Magnesium carbonate 35.2 Iron oxide and aluminum oxide 1.5 Zinc sulphide 4.5 Insoluble 9.3 100.0 Table II. Comparative Data, Galena and Ferrosilicon Ferro- Diner-Gelenaa siliconb ence Operating costs per ton milled, ct. 21.21 9.12 12.09 Medium consumption per ton milled, lb 0.80c 0.15 0.65 Reagent consumption per ton milled, lb 0.45 0.02 0.43 Tailing assay, pct Zn 0.310 0.297 0.013 Concentrate. oct Zn 12.08 10.33 1.75 Heavy medla ieparatlon recovery. pct 89.38 90.22 0.84 Mill feed rate, tons per hr 153 166 13 Heavy mesa separation feed rate. tons per hr 100 10 0 Tons milled per heavy media separation man shift 350 620 270 Mill feed to coarse tailings, pct 51.0 56.7 5.7 Lost mill time, pct 5.6 5.0 0.6 Power consumption, kw-hr per ton 2.06 1.92 0.14 a 1947. " First 6 months of 1950. c Net consumption after deducting credit for reclaimed waste galena. Consumption of new galena was 1.320 lb per ton milled. For entire life of galena operation, a credit of 40 pct of the value of the new galena added was realized from the sale of waste galena. comparisons given in this report cover the first 6 months of 1950 as representing the ferrosilicon operation, and the year 1947 as representing the galena operation. This was the last full year in which galena was used exclusively and is representative of the best work done during the 10 years of operation with this medium. After only 2 years' operating experience, with ferrosilicon and treatment of 1,807,585 tons many advantages have been revealed and are summarized in Table 11. Development Prior to the introduction of the heavy media process, all the mill feed was crushed through 5/8 in. and treated by jigging. A finished tailing assaying 0.66 pct Zn was made on rougher bull jigs, and cleaner jig tailings were ground for treatment by flotation. The first test work on the sink-and-float method of mineral beneficiation was carried out at Mascot in 1935, using a 3-ft cone and galena medium for batch tests. The following year a 6-ft cone was installed for pilot-plant work. This unit became a part of the mill circuit on March 1, 1936, and handled a gradually increasing tonnage in the next 2 years as the process developed to the point where it could treat all the + 3/8-in. material in the mill feed. Coarse jigging was then discontinued on March 1, 1939, and all coarse tailings have been made by the heavy media separation plant since that time. Feed Preparation: The original feed preparation plant consisted of a drag washer followed by two 4x10-ft Allis-Chalmers washing screens. A surge bin and two additional 5x12-ft AC washing screens were added in 1943. Use of primary and secondary washing screens was found essential to provide the cleanest possible feed for the cone and thereby avoid excessive contamination of the galena medium. Improved washing was obtained by replacing the drag washer with a 7x20-ft Allis-Chalmers scrubber, shown in Fig. 2, which has been in service since May 1944. Throughout the life of the galena operation, delivery of extremely muddy ore to the mill overloaded the medium cleaning system, and it frequently was necessary to cut off the feed and clean the medium for several hours until its normal viscosity had been re-established. The cleaning circuit

Jan 1, 1952

By John M. Karpinski, David S. Cahn

A new device based on a constant rate of loading has been developed for the compression testing of green and dry pellets. This device has a lower test variation for green pellets than the existing constant deformation rate tester. While the test variations for dry pellets are similar for both types of machine, the new tester is sensitive enough to measure dry pellet deformations of 15 µ in. for cyclic loading studies. In the study discussed here, the compressive load at which a pellet breaks proved to be a normally distributed random variable, and the load-deformation relationship was found to be linear for dry pellets but not for green pellets. The paper emphasizes the need for a standardized test procedure throughout the industry. For many years the green and dry compressive strengths of pellets have been considered as possible criteria of pellet performance during induration. Little attention, however, has been given to standardizing the unfired pellet compressive strength test or to determining the characteristics of this test. The purpose of this study was to compare the constant deformation rate test with constant loading rate

Jan 1, 1969

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 J. Mizrahi, A. M. Baniel, A. Mitzmager, S. Star

This article gives an account of the latest developments in the use of TBE for sink-float separation of silicate minerals. A number of minerals for which the process has been tested and found suitable are reviewed. Specific examples are discussed in relation to the particular problems of sink-float separation and recovery of TBE as compared with non-silicate minerals. The economic aspects are outlined. Silica and silicates comprise the most common and abundant group of minerals, and since they are present in variable amounts in most ores, they constitute an appropriate subject for the review and discussion of the recent developments in the use of TBE for mineral concentration. The silicates considered as a group have an extremely wide range of chemical and physical properties. Some of these properties, which are common to all minerals, are of importance to concentration by TBE and are discussed in some detail in this paper, as well as the development of methods and apparatus for concentration and recovery of TBE and several examples of concentration of silicates. PROPERTIES OF MINERALS WITH RESPECT TO CONCENTRATION BY TBE The properties considered as most important to concentration by TBE are specific gravity, structure, surface and shape. The specific gravity of a mineral is a fundamental property and is constant between very close limits for the same species of different sources if particles are pure and free from cavities or cracks. In many species, however, solid inclusions and changes in chemical composition due to replacement of elements cause variations in the specific gravity. The plagio-clase series of feldspars between albite and anorthite illustrate this point (specific gravity from 2.60 to 2.77). Fig. 1 shows the specific gravity distribution of some of the common silicate minerals.' The highest specific gravity is about 4.7 (zircon) and the lowest is 2.0 (sepiolite), with most of them ranging between 2.6 and 3.5. Although most of the silicates given in Fig. 1 are found as constituents of rocks, those generally considered as rock making minerals are comparatively few. They include quartz, feldspars, nephelite, sodalite, leucite, mica, pyroxenes, amphiboles, chrysolite, kaolin, talc and serpentine. These minerals may appear in igneous rocks as coarse crystals in granites, syenites and diorites or as very fine grained dispersions in basalt and felsite. When the latter rocks occur in an ore, their specific gravities depend on the minerals present and structure. The same is true for the sedimentary and metamorphic rocks (sandstones, shales, gneiss, slate, schists). A comparison of specific gravities for the minerals listed in Fig. 1 with that of TBE (see Table I) shows that the important rock-forming minerals, with the exception of pyroxenes, amphiboles and chrysolite (olivine), lie in the specific gravity range of less than 2.96 and can therefore be separated by TBE. The ore structure is important both in determining the apparent specific gravity and the efficiency of TBE recovery. The structure can be compact (as in

Jan 1, 1963