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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 Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

FOR successful electrostatic separation, the behavior of the component minerals of the tested ore under various conditions should be known. In the course of investigation, it was found that these are difficult to measure accurately by the presently available methods. The method used by Johnson1 is handicapped by the fact that particles in the extreme front and rear end of the falling stream are difficult to control or observe accurately. Furthermore, this method is not applicable when the tested sample consists of more than one mineral. The deflection method, as reported by Fraas,' is inadequate to show the differential amounts of mineral particles deposited at the various horizontal positions under the grounded roll of the separator. This weakness is further noted when the tested sample contains more than one mineral. The purpose of this paper is to introduce a new method capable of measuring the behavior of mineral particles under various conditions, and of defining the optimum condition for separation. The main factors which influence the behavior of mineral particles in electrostatic separation are also described. dividing gauge and the two attached collecting chutes of a Johnson separator.~ This distribution analyzer consists of sixty or more 1.0-cm wide cells as shown in part in fig. 1; it has been fully described in a previous article." Except where otherwise stated, the testing procedure used in this paper is as follows: (a) calculate the weight of feed for a l-min run at a constant feed rate of one layer of closely packed particles falling on the grounded roll; (b) weigh and dry the feed at 100°C for 12 hr; (c) place the top of the distribution analyzer level 4 cm below the bottom of the grounded roll, so that the cells are longitudinally parallel to the grounded roll and the zero edge of a suitable cell is directly under the front vertical tangent to the grounded roll (see fig. 1); (d) with the charged roll at a negative potential of 15.5 kv, the dry hot sample is introduced into the feed chute at a constant feed rate; and (e)

Jan 1, 1951

By Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

FOR successful electrostatic separation, the behavior of the component minerals of the tested ore under various conditions should be known. In the course of investigation, it was found that these are difficult to measure accurately by the presently available methods. The method used by Johnson1 is handicapped by the fact that particles in the extreme front and rear end of the falling stream are difficult to control or observe accurately. Furthermore, this method is not applicable when the tested sample consists of more than one mineral. The deflection method, as reported by Fraas,' is inadequate to show the differential amounts of mineral particles deposited at the various horizontal positions under the grounded roll of the separator. This weakness is further noted when the tested sample contains more than one mineral. The purpose of this paper is to introduce a new method capable of measuring the behavior of mineral particles under various conditions, and of defining the optimum condition for separation. The main factors which influence the behavior of mineral particles in electrostatic separation are also described. dividing gauge and the two attached collecting chutes of a Johnson separator.~ This distribution analyzer consists of sixty or more 1.0-cm wide cells as shown in part in fig. 1; it has been fully described in a previous article." Except where otherwise stated, the testing procedure used in this paper is as follows: (a) calculate the weight of feed for a l-min run at a constant feed rate of one layer of closely packed particles falling on the grounded roll; (b) weigh and dry the feed at 100°C for 12 hr; (c) place the top of the distribution analyzer level 4 cm below the bottom of the grounded roll, so that the cells are longitudinally parallel to the grounded roll and the zero edge of a suitable cell is directly under the front vertical tangent to the grounded roll (see fig. 1); (d) with the charged roll at a negative potential of 15.5 kv, the dry hot sample is introduced into the feed chute at a constant feed rate; and (e)

Jan 1, 1951

By Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

0. C. Ralston and F. Fraas—Dr. Sun and associates have presented an interesting paper not all of which is comprehended by us. The data assembled measure the deflections of particles in an electrostatic field as a function of a number of independent variables and some dependent variables that are not sharply differentiated. These data are all based on a Johnson type machine of definite, well-described geometry, something not often done in electrostatic separation literature. One new fact brought out by this technique is the effect of coal dust on admixed pyrite and quartz. The effects are opposite in character, as should be expected and we do not agree with the authors that these effects are negligible. Fraas' also used a multiple cell "distribution analyzer" and gives in fig. 5 of his paper a straight line plot with no humps or curves. This is not necessarily at variance with Sun's results because Fraas used a larger gap between electrodes and had no evidence of particles adhering to or dropping off the charged electrode. The section of Sun's paper on effect of surface conductivity contains a speculation that the dielectric constant "represents more or less the electrical conductance of the bulk body instead of the surface of the mineral particles." A simple picture of the meaning of the dielectric constant is that it is the specific inductive capacity of a dielectric when used as the dielectric between the plates of a condenser. It is at once evident that the above speculation confuses capacity with conductance—two definitely independent variables. We ask the authors to state in what group or subdivision their garnet belongs; what method and units were used in calculating the data of col. A, table I and their meaning; what was the temperature of the carrier roll and, finally, has any effort been made to investigate the effects of particle shape on distribution in the electrostatic field? S. B. Hudson—I have read this article with great interest. We have been engaged in research work on the principles of electrostatic separation in this laboratory for some time now, and our findings agree with those of the authors in many respects. The article shows evidence of careful and valuable research in the field of electrostatic separation. A "distribution analyser," very similar to that described in an earlier article by one of the authors," was incorporated in an inclined plate-type electrostatic separator designed and built in the Melbourne University laboratory in 1948 for investigation purposes.22 The actual splitting edges were machined from y! in. per-spex, and the paper hoppers were supported on linen thread immediately below the perspex dividers. These dividers fitted into machined slots in a framework to give accurate Yz-in. spacings. The hoppers (staggered) fed directly into a rack of test tubes, which is supported on a vertical pantograph arrangement. The rack was positioned with guides on the horizontal pantograph stand, and this ensured positive alignment when replacing the rack after making weighings. In later work, when much heavier feed rates were used, of the order of 30 to 40 Ib per in. per hr a rack fitted with rectangular metal containers and similarly aligned was used. Some work was done here on comparing the distributions of minerals when passed separately and when passed as a mixture, and it was found that there was quite an appreciable difference in the two results.= However, in our separator the particles do not pass down the plate in a single layer, and this difference is probably caused by collisions of one mineral particles with the other mineral particles. In most of the investigational work here, the change of the center point of the distribution is measured to establish the effect of a variable, such as voltage. Two minerals (zircon and rutile) have been studied rather exhaustively, and it was found that their distributions are very nearly normal. Owing to the sharpness of the distribution curves, the usual method of obtaining the mean or median was inaccurate, and was not used; instead the mean (also the median), calculated on the assumption of a normal distribution, was used to locate the center point of each distribution and proved satisfactory. The effect of polarity becomes very apparent in the plate-type separator where frictional charges play a very important part when using highly resistive minerals such as zircon. With rutile, a comparatively conductive mineral, polarity of the electrode has little effect. On the other hand, the magnitude of the voltage has a far greater effect on conductive than on resistive minerals. Shiou-Chuan Sun (authors' reply)—Thanks are extended to Drs. Ralston and Fraas for their keen interest in this paper. Their questions concerning coal dust,

Jan 1, 1951

By Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

0. C. Ralston and F. Fraas—Dr. Sun and associates have presented an interesting paper not all of which is comprehended by us. The data assembled measure the deflections of particles in an electrostatic field as a function of a number of independent variables and some dependent variables that are not sharply differentiated. These data are all based on a Johnson type machine of definite, well-described geometry, something not often done in electrostatic separation literature. One new fact brought out by this technique is the effect of coal dust on admixed pyrite and quartz. The effects are opposite in character, as should be expected and we do not agree with the authors that these effects are negligible. Fraas' also used a multiple cell "distribution analyzer" and gives in fig. 5 of his paper a straight line plot with no humps or curves. This is not necessarily at variance with Sun's results because Fraas used a larger gap between electrodes and had no evidence of particles adhering to or dropping off the charged electrode. The section of Sun's paper on effect of surface conductivity contains a speculation that the dielectric constant "represents more or less the electrical conductance of the bulk body instead of the surface of the mineral particles." A simple picture of the meaning of the dielectric constant is that it is the specific inductive capacity of a dielectric when used as the dielectric between the plates of a condenser. It is at once evident that the above speculation confuses capacity with conductance—two definitely independent variables. We ask the authors to state in what group or subdivision their garnet belongs; what method and units were used in calculating the data of col. A, table I and their meaning; what was the temperature of the carrier roll and, finally, has any effort been made to investigate the effects of particle shape on distribution in the electrostatic field? S. B. Hudson—I have read this article with great interest. We have been engaged in research work on the principles of electrostatic separation in this laboratory for some time now, and our findings agree with those of the authors in many respects. The article shows evidence of careful and valuable research in the field of electrostatic separation. A "distribution analyser," very similar to that described in an earlier article by one of the authors," was incorporated in an inclined plate-type electrostatic separator designed and built in the Melbourne University laboratory in 1948 for investigation purposes.22 The actual splitting edges were machined from y! in. per-spex, and the paper hoppers were supported on linen thread immediately below the perspex dividers. These dividers fitted into machined slots in a framework to give accurate Yz-in. spacings. The hoppers (staggered) fed directly into a rack of test tubes, which is supported on a vertical pantograph arrangement. The rack was positioned with guides on the horizontal pantograph stand, and this ensured positive alignment when replacing the rack after making weighings. In later work, when much heavier feed rates were used, of the order of 30 to 40 Ib per in. per hr a rack fitted with rectangular metal containers and similarly aligned was used. Some work was done here on comparing the distributions of minerals when passed separately and when passed as a mixture, and it was found that there was quite an appreciable difference in the two results.= However, in our separator the particles do not pass down the plate in a single layer, and this difference is probably caused by collisions of one mineral particles with the other mineral particles. In most of the investigational work here, the change of the center point of the distribution is measured to establish the effect of a variable, such as voltage. Two minerals (zircon and rutile) have been studied rather exhaustively, and it was found that their distributions are very nearly normal. Owing to the sharpness of the distribution curves, the usual method of obtaining the mean or median was inaccurate, and was not used; instead the mean (also the median), calculated on the assumption of a normal distribution, was used to locate the center point of each distribution and proved satisfactory. The effect of polarity becomes very apparent in the plate-type separator where frictional charges play a very important part when using highly resistive minerals such as zircon. With rutile, a comparatively conductive mineral, polarity of the electrode has little effect. On the other hand, the magnitude of the voltage has a far greater effect on conductive than on resistive minerals. Shiou-Chuan Sun (authors' reply)—Thanks are extended to Drs. Ralston and Fraas for their keen interest in this paper. Their questions concerning coal dust,

Jan 1, 1951

By K. A. Natarajan, I. Iwasaki

Platinum electrodes are not inert as often thought to be. The reactivity of platinum electrodes can explain their erratic behavior in many electrochemical measurements of metallurgical interest, e.g, in flotation systems, streaming potential measurements, contact-angle measurements, and in leaching systems. The anomalous behavior of platinum electrodes in redox potential measurements in aqueous systems was studied through Eh and pH measure ments in water-oxygen, iron-water-oxygen, and manganese-water-oxygen systems. Stability relations between Fe++ and Fe (OH), and between Fe (OH), and Fe (OH), were studied to judge the correspondence between experimental and theoretical equilibrium lines. The practicality of redox potential measurements in estimating ferric-ferrous ratios in aqueous systems was investigated along with their suitability as indicators in leaching operations, e.g., the removal of iron by aeration from manganese leach solutions. Platinum electrodes have often been used in the measurement of dissolved oxygen concentrations and of redox potentials (Eh) in a variety of fields, e.g., analytical chemistry,' corrosion," geology and mineralogy,,'' biology,"' sewage treatment,' * hydrometallurgy,"I" and flotation."la The effectiveness of Eh-pH diagrams, first reported by Pourbaix' in 1949, has contributed much towards the theoretical understanding of numerous problems encountered in the metallurgical industry. Not many references are available in the literature, however, wherein attempts have been made to confirm Eh-pH diagrams from experimental measurements. One reason might be that, in spite of the apparent simplicity of the electrochemical technique, the direct measurement of Eh involves complex practical problems.' Factors such as the purity of the solution, the type of electrodes used, the history of the indicator electrode, and the type of atmosphere (namely, oxidizing, reducing, or inert) do have effects on the measured Eh values. The influence of mixed potentials cannot be underestimated. The poisoning of platinum electrodes by organic and inorganic impurities present in the solution may lead to erratic results. Platinum, commonly thought to be an inert electrode material, is not really so, as attested by a number of previous investigators who advised caution concerning the anomalous behavior of platinum electrodes in various electrochemical measurements.'" In the present article, a few pertinent experiments related to Eh-pH measurements in systems of interest in the metallurgical and water pollution fields are described in an attempt to correlate such information with what is already known, especially in the electrochemical literature. Iron-water and manganese-water systems were selected with a view of studying the correspondence between experimentally observed and theoretically established equilibrium lines. The work included an investigation of the behavior of platinum electrodes with respect to pretreatment and adsorption characteristics, the measurement of dissolved oxygen concentrations and their relation to Eh, the determination of the electrode potential of the ferric-ferrous couple at different pH, and the measurement of oxidation potentials in iron-manganese leaching systems. Experimental Procedure A rotating platinum electrode was used in many of the measurements to study the effect of rotation on measured Eh values. The electrode made by the Pine Instrument Co. consisted of a stainless steel rod with a platinum disk soldered to the end. It was covered with a Teflon insulation along the sides, so that only the circular tip of the electrode was exposed to the solution. Prior to its use, the platinum surface was brightened on a metallurgical polishing wheel with alumina as an abrasive, unless specified otherwise. The electrode was rotated with a Sargent cone-drive motor at 350 rpm. The contact of the electrode with the external circuit was made by filling a notch at the top of the stainless steel shaft with mercury and by dipping a copper wire into the mercury pool. The performance of the rotating platinum electrode was compared with the performances of a Beckman platinum button electrode and a platinum wire electrode. All the potentials were measured with respect to a saturated calomel electrode. A saturated KC1-agar bridge was used to minimize the liquid junction potential. A Beckman Zeromatic pH meter together with a Beck-man electrode switch was used to measure both the Eh and pH. A double-walled, all-Pyrex jar with a capacity of about 1 liter and themostated by circulating water of constant temperature was used for a reaction cell. Four equally spaced ports in the cover provided access for a glass electrode, a salt bridge connecting the saturated calomel electrode, a dispersion tube for bubbling gases into the cell solution, inlet and outlet tubes for passing the desired gases over the solution, and a graduated burette for introducing reagents from outside. The rotating platinum electrode was inserted through an opening in the top-center of the cover, and a positive gas pressure was maintained inside the cell to prevent air from entering into the cell compartment. A magnetic stirrer was used to mix the solution inside. For the determination of dissolved oxygen in the test solutions, the polarographic techniquex was used.

Jan 1, 1971

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 R. L. Bleifuss

Most pellets swell only slightly during reduction, but some swell so enormusly that their increase in size is termed catastrophic. Since catastrophic swelling produces irregularities in blast furnace operation, an understanding of the phenomenon is of practical as well as theoretical interest. Observations have shown that catastrophic swelling is characterized by the growth of a network of fine iron filaments. Evidence is presented to show that the growth of these filaments is promoted by the presence of calcium in the magnetite lattice of the original ore. During reduction, the limited mutual solubility of CaO and FeO produces an unstable condition that inhibits the uniform nucleation of metallic iron. Other elements in the magnetite lattice whose oxides are continuously soluble in FeO in the critical temperature range have no effect. Magmatic segregation or pyrometasomatic replacements are the most likely deposits to contain calciferous magnetite. Over the past few years there has been an increasing concern throughout the iron and steel industry over the tendency of pellets to swell during reduction. Although the tendency for swelling is only one of several attributes that determine the behavior of pellets in the blast furnace, swelling-reduction tests are now part of the normal evaluation procedure of many pellet consumers. During a swelling-reduction test, pellets display either the moderate increase in size known as normal swelling, or the enormous expansion known as catastrophic swelling. This paper is concerned with the origin and control of catastrophic swelling and how it is enhanced or inhibited by the presence of minor elements. Swelling Behavior The swelling behavior of pellets during reduction is indicated in Fig. 1. The lower curve represents normal swelling and the upper curve catastrophic swelling.' During the initial stages of reduction, that is, up to about 20-25% reduction, the two curves are considered to be equivalent. Beyond this point normal pellets undergo a gradual decrease in size with increased reduction. By contrast, pellets which exhibit catastrophic swelling show a second dramatic expansion between 25 and 70% reduction. The initial normal swelling develops as the hematite of the pellet is successively reduced to magnetite, and then to wustite. Work by Edstrom' and others has shown that the reduction of hematite to magnetite is accompanied by a 25% volume increase, whereas the magnetite to wustite stage shows a volume decrease. The volume increase overlaps the wustite field because magnetite is formed simultaneously with the wustite during topochemical reduction of the pellet. The catastrophic expansion is clearly related to the wustite, to

Jan 1, 1971

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 T. D. Tiemann

The caustic extraction of silica from Wisconsin and Minnesota taconite was investigated by bomb digestion over the temperature range from 312 to 40S°F at caustic concentrations from 25 to 500 gpl. Residual concentrates containing as much as 65 pct iron were obtained. The high rate of silica extraction is related to a microcrystalline variety of quartz, since it is shown that the rate of extraction of silica from quartz is directly proportional to its surface area. Increasing dependence on foreign ore reserves makes the need for developing the technology necessary for treatment of our low grade ores even greater than it was at the time of World War 11, when much research was conducted on the extraction of alumina from clays. Representative of these were the studies carried on at the University of Illinois.' More recently, the U. S. Bureau of Mines investigated the production of alumina from anorthosite, and currently the aluminum industry is showing renewed research interest in this field. The studies reported here are part of a continuing research program based on the above concepts and aimed at the eventual development of processes for the treatment of low grade ores, particularly those of iron. The fact that large reserves 3 of low grade nonmagnetic taconite are contained in the iron formation of the Wisconsin Gogebic range have made this ore an important and attractive target for investigation. The idea of removing silica by caustic extraction is not new.495 The digestion in caustic solution of ground flint, diatomite, and other forms of chalcedonic silica is the basis for the production of soluble sodium silicates by the wet-process. Upgrading bauxites by the caustic extraction of silica has been investigated6 and is the subject of a German patent.7 Although mention is made by Roe' of the caustic extraction of silica from low grade Michigan and Wisconsin iron ores, and he lists a number of patents and references on the subject, recent work in the field does not appear to be extensive. As far as the author knows, there has been no work in this field on these particular ores. It is hoped that this paper will stimulate further investigation which will eventually lead to an econclmic means of processing these large low grade reserbes. TACONITES The taconites investigated in this study were from Minnesota and Wisconsin. That from Minnesota, a semi-taconite, Nas obtained from the Oliver Iron Mining Division, U. S. Steel Corporation, and represents a sample of an oxidized iron formation from their King Mine on the Mesabi range. The Wisconsin taconite came originally from the trench sampling of the U. S. Bureau of Mines, and represents a composite of the Norrie, Pabst, Yale, Plymouth, and Pence members of the Gogebic iron formation. Assays on these samples are given in Table I. A typical calcu1atr:d average analysis of the Wisconsin taconite is given in Table 11. Chemically, the Minnesota semi-taconite sample contains about 39 pct iron and 39 pct silica, while the Wisconsin taconite contains 30 pct iron and 53 pct silica, both on the dry basis.

Jan 1, 1962

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 L. J. Boucher

The beneficiation process at Northampton is described and reasons are given for installing a flotation plant. The economics of running the plant, the difficulties of operation, and subsequent remedial action are discussed. CEMENT rock from a quarry operated by the Northampton plant1, 2 of the Universal Atlas Cement Co. has a relatively low average CaCO³ content. To meet specifications required for cement manufacture it was formerly necessary, over succeeding years, to add increasingly larger amounts of high-lime, low-alumina rock, which was obtained from a quarry operated by the company at a point along the New Jersey-New York state border, 79 miles from Northampton. Previous to 1940, 8 pct of total kiln feed consisted of high-lime rock. From 1940 to 1943, this figure jumped to 13 pct. Today it would be necessary to use 18 pct of this high-lime to meet current specifications. Increasing transportation costs made this an expensive operation. In addition, this high-lime rock quarry was contaminated with granite and high magnesia pockets. As there was no available high-lime rock in the Lehigh Valley, it was decided in 1936 to install a pilot plant where experimental work was done on processing local cement rock with flotation and burning the resulting product in a small kiln. This plant was in operation for six months and at the end of that period a suitable method of flotation had been established. From these studies it was also found that if the proportion of shipped-in rock rose slightly above 10 pct it became economically advisable to use the flotation process. It is the object of this paper to present briefly the operation of the flotation system at the Northampton plant. It seems advisable to comment on the raw material used in this process. The Northampton quarry lies geologically in what is known as the Jacksonburg formation of Trenton limestone.' In Pennsylvania this extends in a southwesterly direction from Portland, through Nazareth, Bath, and Northampton and across the Lehigh River to Ormrod. This argillaceous limestone was formed from marine deposits laid down during the early Ordivician period when a shallow sea covered most of Pennsylvania. Carbonaceous matter in the original deposit formed graphitic carbon under later metamorphic forces. From this graphite the cement rock of the Lehigh Valley derives its dark color. Most of the graphite must be removed before lime flotation can proceed. The quartz in this rock is desirable from the chemist's point of view, as it furnishes him with a form of silica necessary in the calcining process, but its abrasive properties damage pumps and piping. Aluminum silicates of potassium, sodium, iron, and magnesium are found in the form of sericite. Cal-cite averages from 69 to 76 pct. Iron in the form of iron pyrites is also present and has no deleterious effect. Talc is present in minute quantities. The

Jan 1, 1954

By C. B. Daellenbach, T. D. Tiemann

The prevention of quartz flotation by chelation of activating ions with the tetrasodium salt of EDTA in oleic acid flotation was investigated using a small vacuum flotation cell. It was found that flotation of quartz could be entirely suppressed in the presence of activating ions during complete flotation of hematite. It was further shown that the concentration of activating ions liberated from Wisconsin Gogebic taconite is sufficient to float quartz in an oleic acid system. Large tonnages of low grade siliceous iron ores occur on the iron ranges of the Lake Superior district. Although the lower grade nonmagnetic types are currently not of economic importance, they constitute a large future potential source of iron and steel, and hence have been the subject for extensive research on their beneficiation. TACONITES Of particular interest to the Dept. of Minerals and Metals Engineering at the University of Wisconsin have been the low grade nonmagnetic taconites of the Wisconsin Gogebic Range. These occur as five members of the iron formation and have been described in detail elsewhere.' Essentially they consist of an intimate mixture of hematite, goethite and quartz, with minor amounts of siderite, magnetite, iron containing silicates, and manganese minerals. A typical mean analysis based on the Bureau of Mines analyses' is: SiO2-52.50%, Fe2O3-43.40%, (Fe-30.38%), A12O3-1.15%, P2O5-0.07%, Mn2O3-0.29%, CaO-0.07%, (S-0.03%), Loss on ignition-2.00%. The quartz in these ores is of a microcrystalline variety,2 and so intimately associated with the iron oxide minerals that extremely fine grinding, in the order of minus 325 mesh, is required for liberation. During grinding and subsequent flotation, the concentration of quartz activating ions in the pulp becomes sufficiently high to activate quartz which floats along with the iron minerals in the presence of oleic acid. cook3 and co-workers concluded that oxidized iron ores released appreciable quantities of quartz-activating ions in water pulps. This is shown to be true in the present investigation. As a matter of fact, application of this phenomenon, as well as the intentional addition of activating calcium ions, is made use of in the an ionic flotation of silica from the iron oxide minerals.4,5 In view of the above, it becomes almost impossible to effect a sharp separation of quartz in the oleic acid flotation of the iron oxide minerals from these taconites, even with satisfactory liberation. This fact led to the current investigation to find a means of linking the activating ions with a chelating agent. The aim was to establish basic data for the quartz-water-oleic acid system which might not only be applicable to iron ores, but also to the flotation of other siliceous ores as well. EXPERIMENTAL METHODS Vacuum Flotation Cell: The vacuum flotation used in this investigation is similar to that used by previous investigators.6,7,8 The cell consisted of a 100 ml graduated cylinder with the lip removed. Vacuum from a water aspirator was applied to the cell through a trap; suitable tubing, and a rubber stopper inserted in the top of the cell. Constant absolute pressure was maintained by means of a manometer connected to the trap with a tee. Flotation: The required amount of activating ion solution was pipetted into a beaker. This was then diluted to 99.0 ml with water having a pH below that desired for the test, or so adjusted with HC1. After addition of 1.0 ml of oleic acid solution (2.0 mg oleic acid per ml), the pH was readjusted to the desired value with NaOH, and the entire solution transferred to the cell containing 1.5 gm of quartz. After vigorously agitating the cell and its contents for 2 min, the vacuum was applied to the cell and the estimated degree of flotation observed for a 3-min period. Activating ions studied included Ca++, Mg++, Al+++, and Fe+ + + added as chloride solutions. In addition,

Jan 1, 1964

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