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By T. D. Tiemann

The extraction of iron oxide from several types of siliceous iron ore by digestion in sulfuric acid solutions was investigated in the temperature range from 80 to 258°F at concentrations from 10-60% H2SO4. Degree and rate of extraction were found to be critical with respect to both temperature and acid concentration, requiring temperatures in excess of 250° F and acid concentrations of 50% to effect 80-90% extraction of iron oxide in 60 min. Rate of extraction is directly related to surface area and apparently controlled by the rate of chemical reaction. The activation energy was found to be approximately 25,000 cal per g mol. Extraction characteristics of the several ores were found to be similar. The Department of Minerals and Metals Engineering at the University of Wisconsin conducts a continuing research program to develop processes for the utilization of low-grade iron ore deposits such as the nonmagnetic taconites of the Wisconsin Gogebic range. Methods under study include high-intensity magnetic and high-tension electrical concentration, flotation, direct reduction and chemical extraction. Concentration of a number of low-grade siliceous iron ores by the extraction of silica in sodium hydroxide solutions has been reported by the writer and others in the literature.1,2,3,4 Residual concentrates containing as much as 65% iron have been obtained from taconites containing less than 30%. Currently an alternative approach is under study, namely the dissolution of iron oxides from siliceous gangue by extraction in acid, with subsequent recovery of the iron from solution. Such a process for extracting iron oxide from the low-grade oclitic deposits of Alberta with hydrochloric acid was investigated and described by Gravenor, Govett and Rigg.5 The current investigation, which is entirely devoted to extraction with sulfuric acid, is the subject of this paper. In part, the objective of this research was to determine the possibility of leaching iron ores in situ with sulfuric acid. The investigation of a recovery method is left to later research. The ores investigated contained hematite (Fe2O3) as the principal iron-bearing mineral and minor amounts of goethite (Fe2O3 .H2O), magnetite (Fe3O4), and iron-bearing silicates. The gangue was essentially quartz occurring as single crystals and microcrystalline aggregates. Detailed description of the ores may be found in the literature.1,2,4,6 A comparison of the specific ores used in this investigation is shown in the partial analyses of Table I. EXPERIMENTAL METHODS All extractions were made at atmospheric pressure using beakers with a mechanical stirrer for those tests run below the boiling point, and Erlenmeyer flasks for those at the boiling point. A thermostatically controlled hot plate was used for heating, and this maintained the mean temperature within the solution to within ±3°F during extraction. The sample of ore for extraction was weighed to + 0.02 g. For tests below the boiling point, the sample was added to the beaker containing the sulfuric acid solution at the desired temperature. For the boiling-point tests, the sample was added to the Erlenmeyer flask followed by distilled water and the

Jan 1, 1968

By P. F. Allen

WITH capacity of 51,000 tpd, the Morenci concentrator produces approximately 49,000 tons of tailing for final deposition. Disposal involves distribution of thickened tailing to dams in such a manner that solids can be used for dam construction and water can be impounded for reclamation. During early operation, although tonnage treated was lower, it was necessary to use orthodox disposal practices in the areas available for tailing. Trestles mounted on a series of dams, constructed of borrow material, carried the pipelines from which tailing were spigoted behind the dams. When adequate disposal areas were finally established, it became desirable to look for more efficient methods. The idea of placing the distribution pipe or manifold directly on the crest of the dam was advanced in 1945. After a period of testing, that procedure in modified form replaced the older trestle arrangement. Marked economies resulted, and time required to prepare dams for succeeding storage cycles was greatly reduced. The tailing disposal area for the Morenci concentrator lies south and southwest of the concentrator building. Ample storage space is available. A wide valley contains three sharply defined ravines lying on an average slope of 4 pct. The more central of these ravines, and the largest of the three, is known as Morenci canyon, which provides substantially more tailing storage area than the other two. Stargo canyon lies to the west; the other is appropriately named East canyon. Total area of the tailing disposal site is nearly 31/4 sq miles, or about 2000 acres. The storage area in use was somewhat more than 850 acres at the end of 1954, or 16 acres per 1000 tons of tailing to be stored daily. At the present time none of the east side dams are in regular use but they are available in emergency. There are five dams on the west side (in Morenci and Stargo canyons) as well as two auxiliary dams adjacent to No. 4 West and No. 5 West. This has made it possible to concentrate men and equipment in a smaller area with consequent savings in both manpower and machinery. Tailing Thickening: Tailing from the flotation process reaches the thickeners with a solid content of about 18 pct. The launder carrying this pulp is subdivided at a central point into seven launders carrying a proportional share to each of seven tailing thickeners, which are located in Morenci canyon close to the concentrator. Five of these thickeners are 300 ft diam and two are 325 ft diam. Thickeners are of the peripheral traction type and rakes travel at the rate of one revolution in 37 to 39 min. The entire installation provides 500,000 sq ft of settling area, or approximately 101/2 sq ft per ton of tailing to be thickened daily.

Jan 1, 1957

By Edwin S. Smith

Current research in the phenomenon of liquefaction has indicated some of the in situ material properties which should be considered in the planning of waste deposits. Practical experience in tailings disposal, however, must be combined with the results of theoretical and laboratory studies to develop the most economical and safe techniques against liquefaction failure. One of the many responsibilities of mining engineers and mill superintendents is providing for the safe and economic disposal of wastes (or tailings) that remain after a mineral has been extracted from an ore. The cost of disposal of these tailings exerts a considerable influence on the minimum grade of ore extracted from the mine. It varies to a great extent with each project and is dependent on many diverse factors, such as the local topographical, meteorological and seismological conditions together with the applicable water pollution laws of the area. The process used to extract a mineral from an ore determines the physical characteristics of the tailings and, thus, is one of the more important factors to be considered in the planning of disposal facilities. Today's trends in ore-processing operations complicate the problems of waste disposal: recently developed extraction methods result in the production of finer tailings, the use of lower grade ores produces greater volumes of wastes, and the enhanced land values in the vicinity of industrial plants preclude the former practice of spreading wastes laterally and force the expansion of waste deposits upwards to greater heights. The frequent occurrence of stability problems in waste deposits indicates that many mining organizations may not be employing current soil mechanics concepts and practices. One of the most spectacular, and often most disastrous, types of tailings dike failure is that resulting from liquefaction of the fine particles combined with a breach of the main retaining dike. TAILINGS DISPOSAL Since the many variations in site conditions, ore bodies and extraction processes dictate the general disposal procedures and mill superintendents incorporate special facilities to provide the most economical disposal method consistent with accepted safety standards, it is only natural that there are almost as many different disposal techniques as there are disposal areas. Methods used in the construction of retaining dikes include: (1) transporting granular material from an adjacent borrow pit and (2) mechanically segregating the total effluent and allowing the coarse fraction to form a retaining dike by hydraulic deposition. The location of a plant and the surrounding topography exert the greatest influence on the layout of tailings disposal facilities. Most of the above-ground waste deposits can be classified as side-hill, (Fig. 1,) cross-valley, (Fig. 2,) or stockpile deposits. Side-hill disposal areas are probably the most common type of layout because of the natural advantage of gravity flow, without the disadvantage of river diversion usually needed for tailings ponds formed behind cross-valley dikes. Stockpiles, which are placed on relatively level ground, require four retaining dikes and the effluent generally has to be pumped up into the disposal area. Changing production techniques or developments adjacent to disposal areas have produced variations or partial combinations of these three categories. In a series of articles, Lenhart describes many of the problems encountered in the disposal of tailings and certain innovations used to overcome them. He discusses weaknesses in design and layout and shows that disposal techniques are often learned the

Jan 1, 1970

By R. W. Jigins

OPERATIONS of the Braden Copper Co. are in the Chilean Andes, southeast of Santiago. Most remote of the company communities is Sewell, a town of 12,000 people, 7000 ft above sea level at the junction of the Coya and Teniente rivers. Four miles down the Coya River is the smelter town of Caletones, elevation 4500 ft, with a population of about 3000. At the junction of the Coya and Cacha-poal rivers 24 miles below Sewell is Coya, elevation 2500 ft, location of the company's two hydroelectric plants. The principal accounting offices; steel, iron, and brass foundries; railroad shops; and major mechanical repair and fabricating shops are at Rancagua in the central Chilean valley, 44 miles from Sewell. The company owned narrow-gage railroad connects with the Chilean national railroad at Rancagua and above Coya is the only modern means of travel. El Teniente mine at Sewell is the largest underground copper mine in the world. The mill, at the portal of the main tunnel, has a capacity of 34,000 tons of crude ore daily, although production is lower during the winter beacause of electric power restrictions. When operating at capacity the mill produces 2000 tpd of copper concentrates and 32,000 tpd of tailings. Tonnage of tailings is equivalent to about 100,000 tons of pulp at normal water to solid ratio. Because the streams flowing from the Andes in the area of Braden's operations are used for irrigation, it is impossible to dump tailings directly to the river. The tailings pulp must be separated into solids and decanted clear water, and the clear water returned to the same water shed from which the original water was removed. Remaining sands and slime must be deposited in a reservoir where they will not cause trouble to farmers.

Jan 1, 1958

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

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

Jan 1, 1961

By M. Yoshinuga, S. Watanabe

Some iron ores, sinters and pellets occasionally show abnormal behavior during reduction which makes them undesirable as blast furnace burden. These may be divided into the following three types: (1) decrepitation of iron ore, (2) swelling of iron ore and pellet, (3) size degradation of sinter. In this paper reduction experiments and microscopic examinations were made to clarify the mechanisms and causes of these abnormal phenomena. Decrepitation occurs mainly when compact hematite ores containing small amounts of limonite are heated to about 500°C. Abnormal swelling of ores and pellets is caused by the rapid growth of fibrous metallic iron from the surface of wustite grains during rapid reduction and is fundamentally related to their original microstructure. Size degradation of sinter occurs through the development of cracks in the reduction stage from hematite to magnetite. About 80% of the iron ore used in Japan's blast furnaces comes from abroad, and the iron and steel industry must import this raw material from distant overseas places at high freight rates. The blast furnace operators, therefore, have made strong efforts to cut costs under these restrictive raw materials conditions by increasing productivity and simultaneously decreasing coke consumption. If, however, a high proportion of the ore burden consists of unsuitable materials furnace irregularities often result, establishing the need for detailed investigations of the basic natures of the ore constituents and their reactions during the reduction process. Recent investigations '-' of various ore burdens suggest that some iron ores, sinters and pellets are occasionally abnormal in appearance and behavior during reduction studies and may prove to be undesirable for use in the blast furnace. In the present study reduction experiments and microscopic observations were made to clarify the causes and the mechanisms of the abnormal behavior observed in some constituents of the burden and their effect on blast furnace performance. REDUCING APPARATUS The four kinds of reducing apparatus used in this investigation were the following: 1. The JIS* (Modified Gakushin') apparatus shown in Fig. la is most commonly used in Japan to evaluate the reducibility of ores under static conditions. It consists of a stainless steel tube of 60 mm inside diam suspended from a balance. 2. Linder's Reduction Test apparatus (Fig. lc) has a reducing barrel 130 mm diam and 200 mm long which rotates at 30 rpm. Some modifications were made in test procedures. 3. Gakushin Standard Test apparatus for determining the reducibility of granular ores (Fig. Id) consists of a horizontal silica tube 32 mm diam, 370 mm long and a movable electric furnace which enables rapid heating and cooling of the specimens. 4. The test apparatus for reduction under load (Fig. lb) gives a comprehensive evaluation of the behavior of ores under conditions simulating the static loading and the chemical reduction conditions encountered in. a blast furnace stack. A stainless steel tube of 50 mm diam and a carbon rod of 35 mm diam were used as the sample container and the support for the weight, respectively. In each investigation care was taken to select the most suitable method and apparatus to distinguish the abnormal nature of the ore in question. SIZE DEGRADATION OF SINTER DURING REDUCTION Sinter makes up the major portion of the blast furnace burden and its behavior during reduction should bear a close relationship to the iron productivity. Recently, some of the sinter was found to show serious size degradation during reduction at a relatively low temperature range. Because this phenomenon was thought to be a possible source of furnace irregularities, tests were made to clarify the phenom-

Jan 1, 1969

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

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

Jan 1, 1951

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

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

Jan 1, 1951

By Robert C. Emmett, Donald A. Dahlstrom

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

Jan 1, 1954

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

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

Jan 1, 1951

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

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

Jan 1, 1951

By C. H. Snyder

Tlie paper deals with details of construction of the Colmol, including improvements in design that will be incorporated in new models. These improvements are results of problems encountered and worked out in experimental operation with the unit. Also included is a descriptive plan of operating where roof and air conditions are of the worst. A summary of production results from the machine is presented. THE various units used in conventional mining are well-built. Generally the universal under -cutter, driven by approximately 50 hp, weighs approximately 25,000 lb, and is a very strong piece of machinery. The drilling machine is not as husky as an undercutter, generally is driven by approximately 15 hp, weighs about 9000 lb, but is a well-constructed, sturdy mechanism. The loader, usually 60 hp, weighs approximately 24,000 lb, and certainly is a strong, rugged unit. The working parts of all three of these sturdy conventional units have been made of tough steel alloys for some time. With the use of three machines having a combined horsepower of approximately 125, with a combined weight of approximately 60,000 lb, and with the possible inherent advantage of single purpose machines performing a single function, the use of considerable explosives has been required to complete the conventional mining cycle. It was therefore our thought, upon conception of the first Colmol,* a machine which was to do the work accomplished by the three conventional units plus explosives, that it must be a rugged mechanism. Now, after building several units, after going through many and varied experiments, and after the expense of some million-odd dollars, we find we have observed and learned nothing that contradicts our original thinking. A continuous miner, if it is to give satisfactory performance, must be a rugged piece of machinery. Ruggedness has been, and is constantly, in the minds of our engineers in designing the Colmol. Improvements where they have been required have been in the direction of more weight, more horse power, better alloys, bigger hydraulic pumps and motors; in short, always, toward more torque, greater strength, more ruggedness. Now not only the finest steel alloys are utilized but more than 100 Timken roller bearings are used. The following is some pertinent data on models being built and to be built: The low unit: Cutting height, 30 to 42 in. By changing the upper gear case the maximum cutting height can be increased to 61 in., giving a variable cutting height of from 30 to 61 in. Width of cut, 9 ft 6 in. Width of body, 75 in. over tracks. Height of body, 271/2 in. Overall length, 23 ft. Driven by three 50-hp motors or a total of 150 hp. Weight, 26 tons. The intermediate unit: Cutting height, 451/2 to 72 in. Width of cut, 9 ft 8 in. Width of body, 77 in. over tracks. Height of body, 43 in. Overall length, 26 ft. Driven by three 70-hp motors or a total of 210 hp. Weight, 35 tons.

Jan 1, 1951

By C. H. Snyder

Tlie paper deals with details of construction of the Colmol, including improvements in design that will be incorporated in new models. These improvements are results of problems encountered and worked out in experimental operation with the unit. Also included is a descriptive plan of operating where roof and air conditions are of the worst. A summary of production results from the machine is presented. THE various units used in conventional mining are well-built. Generally the universal under -cutter, driven by approximately 50 hp, weighs approximately 25,000 lb, and is a very strong piece of machinery. The drilling machine is not as husky as an undercutter, generally is driven by approximately 15 hp, weighs about 9000 lb, but is a well-constructed, sturdy mechanism. The loader, usually 60 hp, weighs approximately 24,000 lb, and certainly is a strong, rugged unit. The working parts of all three of these sturdy conventional units have been made of tough steel alloys for some time. With the use of three machines having a combined horsepower of approximately 125, with a combined weight of approximately 60,000 lb, and with the possible inherent advantage of single purpose machines performing a single function, the use of considerable explosives has been required to complete the conventional mining cycle. It was therefore our thought, upon conception of the first Colmol,* a machine which was to do the work accomplished by the three conventional units plus explosives, that it must be a rugged mechanism. Now, after building several units, after going through many and varied experiments, and after the expense of some million-odd dollars, we find we have observed and learned nothing that contradicts our original thinking. A continuous miner, if it is to give satisfactory performance, must be a rugged piece of machinery. Ruggedness has been, and is constantly, in the minds of our engineers in designing the Colmol. Improvements where they have been required have been in the direction of more weight, more horse power, better alloys, bigger hydraulic pumps and motors; in short, always, toward more torque, greater strength, more ruggedness. Now not only the finest steel alloys are utilized but more than 100 Timken roller bearings are used. The following is some pertinent data on models being built and to be built: The low unit: Cutting height, 30 to 42 in. By changing the upper gear case the maximum cutting height can be increased to 61 in., giving a variable cutting height of from 30 to 61 in. Width of cut, 9 ft 6 in. Width of body, 75 in. over tracks. Height of body, 271/2 in. Overall length, 23 ft. Driven by three 50-hp motors or a total of 150 hp. Weight, 26 tons. The intermediate unit: Cutting height, 451/2 to 72 in. Width of cut, 9 ft 8 in. Width of body, 77 in. over tracks. Height of body, 43 in. Overall length, 26 ft. Driven by three 70-hp motors or a total of 210 hp. Weight, 35 tons.

Jan 1, 1951

By Edward Martinez

The reactions occurring in the copper segregation process were studied by heating mixtures of chryso-colla, salt, and a reducing agent. The techniques used in this investigation were differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). The initial chloridizing of the oxide copper occurs below 500°C and cupric chloride may be the compound formed. Volatilization and reduction of the copper chloride take place above 550°C, although these reactions do occur to some extent below 550°C. A difference in the distribution of the metallic copper in the reducing agent was found when sodium chloride was used compared to magnesium or calcium chloride. Segregation was obtained with mixtures of cupric oxide, salt, and charcoal with no quartz or silicate present. Preliminary tests indicate that segregation may occur even in the absence of water vapor. The DTA and TGA techniques were found to be of great value in studying these reactions. The copper segregation process is a method for treating oxide copper ores. The ore is heated, most commonly in a kiln, with a reducing agent and a halide salt, generally sodium chloride, to approximately 700°-750°C and held for 30 to 60 min. The addition of the salt results in the migration and reduction of the oxide copper to metallic copper on the surface of the reducing agent rather than in the original mineral structure. The metallic copper can then be recovered by a subsequent leaching or flotation step.1-4 Experimental work in 1923 on oxide copper ores from the Sagasca mines in Chile led to the discovery of the reduction process. It was found that during the reduction of the copper mineral with coal, at about 700°C, the copper migrated to the surface of the coal instead of remaining disseminated in the ore. The unusual behavior of the copper, which caused it to migrate from the ore particles to the coal, was traced to the presence of a small quantity of sodium chloride occurring in the ore. Further experiments with oxide copper ores proved that the presence of a halide, either occurring naturally or admixed with the ore, caused the copper to migrate from the ore particles and to segregate in the charge. The ratio of salt to reducing agent may vary somewhat, but in general 0.5% to 1.5% sodium chloride and 0.5% to 1.0% coal or coke have been mixed with the ore. Less than stoichiometric amounts of salt are required to convert all the oxide copper to a chloride, so that a cycle of reactions appears to be occurring. Several investigations reported in the literature have proposed reactions to explain the mechanism of the process.5-10 There appears to be general agreement that water is necessary for the reactions to occur; that once the sodium chloride reacts with silica or silicates in the ore, the hydrochloric acid formed will convert the cupric oxide in the mineral structure to cuprous chloride. However, there appears to be a conflict of opinion on the reactions involved in the reduction of the cuprous chloride to metallic copper on the surface of the carbon. Rey has indicated that salt, water vapor, and

Jan 1, 1968

By J. Steininger

The Hallimond tube flotation of sphalerite and pyrite with sulfhydryl collectors has been found to be inhibited near neutral pH in the presence of copper salts. Alkalimetric titration studies of the reactions between collector and activator indicate that this effect is apparently related to the formation of basic copper collector complexes. The conditions of formation of these compounds and their stability are in good agreement with the depression observed in flotation experiments. The general procedure employed in the flotation of zinc sulfide minerals requires activation of the mineral surface by heavy metal ions, usually copper, which react more easily than zinc with the common short-chain ionic sulfhydryl collectors.' Commercially, zinc flotation is carried out under alkaline conditions (pH 7 to 10.5), but small-scale experiments indicate that zinc sulfide minerals can also be activated and collected under acid conditions. Selectivity, however, becomes poor at low pH and undesirable materials such as pyrite, silicates and carbonates tend to report in the float. Several years ago, Wark and Sutherland~, measured the contact angles of air bubbles on copper-activated sphalerite immersed in various sulfhydryl collector solutions. When represented, as in Fig. 1, on pH-sodium cyanide diagrams, the results are characterized by 'islands' of non-contact near neutral pH and

Jan 1, 1969

By R. C. Ferguson, Harlowe Hardinge

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

By R. W. Ludt, C. C. DeWitt

The use of froth flotation for the separation of minerals has become one of the most important of ore dressing processes. Its particular adaptability to the enrichment of low grade ores has made the process an important factor in the national economy. The methods have been extended to the recovery of a great number of minerals. Among the few minerals which have resisted efforts toward industrial flotation is chrysocolla, a hydrated partly colloidal copper silicate. Chrysocolla, being a product of natural oxidation, has been found to occur in small quantities with many ores which are recovered by flotation methods. In present practice, these small quantities of copper silicate pass off with the tailings and are lost. The advantages to be gained by a satisfactory process for the recovery of chrysocolla is apparent. Any application of principles which points a way toward the satisfactory industrial flotation process for copper silicate would be of advantage. This paper presents an attack on this problem. Two methods for the recovery of chrysocolla have been developed by the United States Bureau of Mines.1,2 They have been successful on a laboratory scale but have been seriously restricted in industrial application by critical requirements in the procedure. In one of the Bureau of Mines methods,' the ore is activated with sodium or hydrogen sulphide in an aqueous solution at a pH of 4. Amy1 xanthate is then used as a collector with pine oil as a frother in the flotation process. An excess of sulphide acts as a depressant and the state of optimum conditions is difficult to control industrially. In the second Bureau of Mines method,2 soap is used as the collector at a pH of between 8 and 9. The diffi- culties with this process are that soap is not a specific collector, that heavy metal or alkaline earth ions cause the formation of insoluble soaps, and that a more acid solution causes the formation of a free acid which does not act as a collector for chrysocolla. The problem of recovering chrysocolla by flotation involves the selection of a suitable collector. The collector molecule must be composed of an active polar group that has an attraction for chrysocolla, and of a hydrocarbon chain. Certain dyes have been shown to have an attraction for certain minerals. Suida3 found that hydrated silicates are colored by basic dyes. Dittler4 showed that chrysocolla, among other colloidal minerals of acid reaction, preferentially takes up such basic dyes as fuchsin B, methylene blue, and methyl green. Endell5 gave information to show that the colloidal material in clay may be determined by its selective adsorption of fuchsin. A simple experiment, likewise, illustrates the difference in the adsorptive power of chrysocolla and of silica for the basic triphenyl methane dyes. When a mixture of chrysocolla and silica is immersed in a very dilute dye solution, less than 5 ppm, the chryso-colla is rapidly dyed and the silica is dyed more slowly. The difference is substantial but one of degree. Dean2 showed that the dyes, crystal violet and toluidine blue, are taken up by quartz in an adsorption type process. The difference in the adsorptive power, however, offers the means by which a new collector may act. To form such a collector, a hydrocarbon chain must be attached to the dye molecule. This involves a process of organic synthesis. Butyl, hexyl, and octyl hydrocarbon chains were selected for substitution in the malachite green molecule. For the purpose of identification, the alkyl-substituted dyes formed are called: butyl-malachite green; hexyl- malachite green; and octyl-malachite green. An outline of the procedure for their synthesis is given in the appendix. It is generally recognized in the preparation of this type of dye that the chemical structure of some of the dye molecules varies. However, a uniform formula is attributed to the dye. Such a procedure has been followed in specifying the structure of these alkyl-substi-tuted malachite green dyes. The structure is given on the basis of their properties as an homologous series of dyes, on their method of preparation, and on the purity of intermediates used. Structure of substituted alkyl malachite green is: C6H4 N(CH3)2 p-R C6H4 CH C6H4 N(CH)2 Procedure The flotation cell is a Bureau of Mines 100-g, batch unit provided with an air inlet at the bottom above which is a variable speed agitator. The agi-

Jan 1, 1950

By Strathmore R. B. Cooke

On the basis of experiments conducted on quartz using a bubble pick-up method, it was shown in an earlier paper1 that this mineral will preferentially adsorb hydrogen, calcium, or sodium ions, depending on the relative concentrations of those ions in the solution in which the quartz is immersed. For quartz particles ranging in size from 0.2 to 1 mm, it was demonstrated that the concentration of calcium in solution (assumed present as ions) necessary to completely activate quartz for flotation is given by the expression: Ca++ = [H+] X 106 + [Na+] X 10-3 In this expression, ionic concentrations are in mols per liter. It was further shown that the pick-up method is apparently more sensitive to changes in reagent concentration than the standard captive-bubble method, and that induction times are apparently much reduced. Since completing these earlier tests, a new type of cell has been constructed; this is shown in Fig 1 and 2. In Fig 1, A is a ground joint, B is a central tube reaching to within a few millimeters of the bottom of the cell, and C is a stopper which may be removed for reagent addition, and serves the further purpose of excluding carbon dioxide from the air during the test. The entire cell is constructed of Pyrex glass, and does not give the trouble experienced with the earlier cell, in which activating ions were released from the glass at high pH values. The only critical factor in the construction of the cell is the clearance between the central tube and the bottom of the cell. This clearance should be sufficiently small that the bubble can be pressed directly on the mineral grains lying on the bottom of the cell. In the pick-up tests to be described in this paper, the reagents used were all of C. P. grade, except the sodium oleate, which was Merck's "neutral powder." The quartz employed was water-clear vein quartz, sized on screens, cleaned with both hydrochloric acid and sodium hydroxide, and given a thorough final washing with distilled water. Experimental procedures were the same as described in the earlier paper. EFFECT OF SIZE OF QUARTZ ON PICK-UP To ascertain the effect of particle size on the adsorption of calcium ions, the quartz was sized from minus 14 plus 20 mesh through the intervening screen sizes to minus 270 mesh plus 400 mesh. Each size was thoroughly cleaned, and then tested in the cell at different calcium chloride and sodium hydroxide concentrations, and at a constant sodium oleate concentration of 20 mg per liter. All particles, within the size range given, exhibited complete pick-up within the curve expressed by the equation above. This presumably means, when the conditions imposed by the equation are satisfied, that this maximum is independent of particle size. However, it was found that the range through which partial particle pick-up occurred progressively broadened as particle size decreased. This is shown in Fig 3, in which curves B, C, and D show the limits at which pick-up just commences (as the pH is increased) for particles of minus 14 plus 28 mesh, minus 65 plus 100 mesh, and minus 270 plus 400 mesh size, respectively. These results indicate that for satisfactory activation, at any given pH, a lower calcium ion concentration is required for fine particles than for coarse particles. EFFECT OF HIGH ALKALINITY ON PICK-UP At calcium concentrations of between 1 and 10 mg per liter, and at high alka-linities, it was noticed that pick-up ceased as soon as calcium hydroxide commenced to precipitate. This effect was investigated at other calcium concentrations, with the same results. Solutions of calcium chloride, containing 105, 104, l03, and l02 mg of calcium per liter were made alkaline with sodium hydroxide until calcium hydroxide just started to precipitate, according to the following equation: CaCl² + NaOH -+ Ca(OH)2 + 2NaCl The beginning of precipitation was taken as that point at which either a faint opalescence appeared in the solution, or a Tyndall cone became ap-

Jan 1, 1950

By F. M. Lewis, E. C. Johnson

Used as a classifier in a plant expanding capacity or changing to closed-circuit operation, the liquid-solid cyclone offers the advantage of being adaptable to existing conditions. This paper presents the costs involved and experience gained from operating a unit of this kind. BY now application of the liquid-solid cyclone to various classification problems in the chemical and metallurgical fields is more or less familiar. A less publicized application receiving wider consideration as time goes on is its use as a classifier in closed-circuit grinding operations. This paper presents the costs involved and the experience gained from operating a unit of this kind. In May 1951 the Tennessee Copper Co. installed in its London mill at Copperhill, Tenn., a liquid-solid cyclone in closed circuit with a 6x12-ft ball mill. The circuit in which this cyclone is used is the regrind or secondary grinding circuit of a copper flotation plant. Originally it consisted of a 7x1-ft conical ball mill in closed circuit with a bowl classifier. Changes made in the primary grinding circuit of the mill necessitated a regrind circuit of greater capacity. The bulk concentrate to be ground was raised from 900 or 1000 tpd to 1200 or 1300 tpd, and to accommodate this increase a 6x12-ft ball mill which had been in the primary grinding circuit was moved into the regrind circuit. When this was done it was found that the classifier was too small to give efficient operation. Space in the plant was at a premium, and since use of the cyclone offered considerable economy the installation shown in Figs. 1-3 was made. The unit installed is a 24-in. diam cyclone classifier with a 20" cone angle. Diameter of the vortex finder is 8 in., the feed entrance area 20 sq. in. Maximum diameter of the adjustable apex valve is 6 in. and minimum diameter approximately 3.5 in. To refer again to Figs. 1 and 2, the cyclone* is mounted motor. It is charged with 18 tons of 1-in. balls. The overflow, or fines, passes by gravity directly to the copper flotation plant. The ball mill discharge and the new feed combine in a sump from which they are sent to the cone by an 8x6 SRL pump at a rate of 497 gpm. Pressure at the cyclone is 5 psi. For reasons dictated by the nature of the ore treated, particularly the bonding of fine grains of chalcopyrite to pyrrhotite, the froth from the second cell of the copper rougher circuit is returned to the regrind mill. This flow, which somewhat complicates the standard closed-circuit calculation, is seen in Fig. 1 to be 156 tpd of solids. This combines with the cyclone classifier underflow to constitute the ball mill feed. The requirement of the regrind product is that it be 77 to 80 pct —200 mesh. The liquid-solids cyclone operating in conjunction with the 6x12-ft mill adequately met this requirement. Of primary importance is the fact that this arrangement provided sufficient classification capacity to replace the undersized classifier and was capable of building up a circulating load to 87 pct, large enough to give an efficient grind. In this instance the circulating load is taken as the ratio of sands to new feed. Any comparison of performance between this liquid-solid cyclone and the replaced classifier would of course be unfair, since the bowl classifier was

Jan 1, 1955