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By T. M. Morris

A SEARCH of the literature reveals that no measurements have been made of the forces acting between a small solid particle whose surface is hydrophobic, and an air bubble to which the solid adheres, both immersed in water. Analyses have been made of the forces acting to support a greased solid on the surface of water, and the forces acting to cause a solid, whose surface is hydrophobic, to adhere to an air bubble in water. The latter analysis T. M. MORRIS, Junior Member AIME, is in the Department of Metallurgical Engineering and Mineral Dressing, School of Mines and Metallurgy, Rolla, Mo. New York Meeting, February 1950. TP 2734 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received May 16, 1949. This paper is the result of work done for part of a doctor's thesis at the Missouri School of Mines and Metallurgy. is often incomplete, however, because the internal pressure of the bubble has been neglected. It will be demonstrated that the internal gas pressure is not a negligible factor when dealing with bubbles of the size encountered in flotation. A study of the forces acting between an air bubble attached to a large flat surface is informative. It must be borne in mind, however, that this is not the condition present in a flotation cell, where the particle is small compared to the size of the bubble. The bubble is allowed to spread to its maximum contact angle on a large flat surface in the first case, whereas in the second case, the spread of the bubble is limited to the surface of the small particle which is presented to the bubble. Kabanov and Frumkinl studied the forces acting to cause adhesion of bubbles of hydrogen to a large surface of mercury, which served as an electrode in a dilute sulphuric acid solution. The force acting to hold the bubble to the mercury surface was found to be the vertical component of the surface tension between hydrogen and the sulphuric acid solution. The forces tending to cause the bubble to separate from the mercury surface were found to be: (1) the force exerted due to the internal pressure of the bubble acting upon the area of contact between the bubble and the mercury surface, and (2) the buoyant force of the bubble minus the hydrostatic force acting at the base of the bubble. These investigators photographed bubbles that were just on the verge of separating from the mercury surface. From these photographs, they measured the contact angle between the mercury surface and the tangent to the hydrogen-solution interface at the point of contact between bubble and mercury surface. They calculated the volume of the bubble and the internal pressure of the bubble. The equivalence between the upward acting and downward acting forces was remarkable. Wark2 pursued an investigation similar to that of Kabanov and Frumkin, and at the same time. His deductions verified those of Kabanov and Frumkin. He also considered the conditions present in flotation and was aware of the effect of the internal pressure of the bubble. Further, he proposed several conditions under which a small solid particle would adhere to an air bubble in water. In 1922, Edser," an English physicist, made the following statement. "It must be remembered that no particle could float stably, but for the possibility of variation of the contact angle, for if this were a constant, a slight tilt would inevitably cause the particle to sink." Wark criticized this statement, maintaining that the contact angle does not vary. The experimental data to be presented indicates that Edser was correct. Experimental Procedure: Briefly, the experimental procedure was as follows. A bubble of air was generated in distilled water. A rod of known diameter, one end of which was water repellent, was attached to this bubble. The weight of the 'rod was measured. The internal pressure of the bubble was measured with a manometer. The hydrostatic head from the surface of the water to the bottom of the rod was measured. The angle between the horizontal projection of the end of the rod and the tangent to the bubble at the circle of contact be-

Jan 1, 1951

By T. M. Morris

G. L. Simard and D. J. Salley—The authors and ourselves" independently came to similar conclusions both as to the value of tracer methods for the study of flotation and the general nature of collector-mineral interaction. The dodecylamine-quartz system appears to be simpler than the dithiophosphate-galena one, and the results therefore appear clearer. It is evident from a comparison of the two papers, however, that care must be used in extending the results to other more complicated systems. In the case of dodecylamine-quartz, the possible influence of micelle formation is of interest. There is no a priori reason why complete monolayer formation should necessarily occur at the concentration for micelle formation, since sorption depends on the nature of the substrate as well as on the properties of the solute in solution which alone determine micelle formation. Indeed, the fact that the onset of increased adsorption, point B in fig. 2, occurs at a concentration one tenth the critical value for micelles might imply that micelle formation does not have an influence. Interestingly enough, the data may be handled by the methods used for multilayer adsorption of gases on surfaces. Thus, if the amount adsorbed is plotted versus concentration, a sigmoid type curve is obtained, similar in form to a so-called Type I1 isotherm13 generally attributed to multimolecular adsorption. A monolayer value of about 0.12 mg per g may be estimated from the knee of this isotherm. On the other hand, if a plot is made according to the multilayer gas equation for adsorption to an unrestricted number of layers, eq 38,13 assuming the critical micelle concentration as the saturation concentration, the data above 120 mg per liter fall on a straight line from which a monolayer adsorption of 0.3 mg per g results. These considerations suggest that the monolayer may occupy an area somewhat less than the geometric value of 0.4 mg per g, and that micelle formation may indeed be a factor in the adsorption. the angle between the solid-liquid and liquid-gas interfaces, measured through the liquid. The value of the equilibrium contact-angle is uniquely determined by the value of the interfacial energies of the three intersecting interfaces and is independent of other forces in the system. When the liquid-gas interface intersects the solid at an edge—as was the case for all the experiments reported in this paper—the orientation of the solid-liquid interface is indeterminate or varies through 90" for a right-angle edge. Mr. Morris has called the angle between the horizontal and the liquid-gas interface for this edge condition a "static contact-angle." I feel that this term is unnecessarily misleading. In the first place, "static contact-angle" sounds too much like "equilibrium contact-angle." In the second place, the magnitude of the "static contact-angle," which I would prefer to call the "supporting angle," is determined by the forces in the system other than those derived from the interfacial energies, hence "contact angle" is misleading. If Mr. Morris had said that the "supporting-angle" is variable and depends on the weight of the particle and size of the bubble and that it has a maximum possible value equal to the equilibrium contact-angle, his discussion would have been more accurate. T. M. Morris (author's reply)—Mr. Kellogg puts forth a reasonable criticism of some of the terminology used in the paper. I agree that the term "static contact angle" may be misleading. Substitution of the term "supporting angle" or "vector angle" may be more suitable. F. X. Tartaron—In this paper the author presents a very interesting mathematical development of the forces present when a mineral particle adheres to an air bubble. Excellent concordance is obtained between mathematical formulation and experimental results. It is the writer's understanding that when the mineral surface presented to an air bubble is greater than the area of contact, the maximum contact angle is obtained. However, this contact angle represents distortion of the bubble, the normal shape of which is spherical or in cross-section, circular. This distortion produces a force that acts in opposition to the force of adhesion between the bubble and particle. Hence, at maximum contact angle, the force of adhesion between bubble and particle is at a minimum for static conditions. However, when the size of bubble is increased, the size of particle remaining the same (and all other conditions remaining the same), the contact angle decreases, the distortion of the bubble decreases and the force in opposition to adherence of bubble and particle also decreases. Thus, there is stronger attachment between bubble and particle. When this situation is applied to actual flotation conditions, it is doubtful that it has any significance. The reason is that the bubbles are normally so much larger than the particles, that in substantially all cases, it is probable that negligible distortion of the bubble takes place. In table IV, the author makes computations for bubbles from 0.50 to 2 mm diam. This is from 0.02 in. to 0.08 in. Certainly the bubbles generated in a flotation machine are far larger than this. The surfaces of the author's bubbles range from 0.79 to 12.6 sq mm. The area of contact of the glass rod (0.15 mm diam) is 0.017 sq mm. Thus, ratio of bubble surface to mineral area of contact ranges from 47 to 741 in round numbers. If we take a %-in. diam bubble and a 65-mesh (0.208 mm) particle of cubical shape, the ratio of bubble surface to mineral area of contact is 2931. Mineral particles do not readily become attached to air bubbles. Taggart has shown that in pneumatic flotation machines collector-coated particles are only temporarily attached to bubbles. They keep falling off and down in the froth but at a delayed rate as compared with gangue. Spedden and Hannan's motion pictures confirm the difficulty of attaching particles to air bubbles. In the agitation froth process, according to Taggart, air is precipitated from the water selectively on to the collector-coated particles and these

Jan 1, 1951

By T. M. Morris

G. L. Simard and D. J. Salley—The authors and ourselves" independently came to similar conclusions both as to the value of tracer methods for the study of flotation and the general nature of collector-mineral interaction. The dodecylamine-quartz system appears to be simpler than the dithiophosphate-galena one, and the results therefore appear clearer. It is evident from a comparison of the two papers, however, that care must be used in extending the results to other more complicated systems. In the case of dodecylamine-quartz, the possible influence of micelle formation is of interest. There is no a priori reason why complete monolayer formation should necessarily occur at the concentration for micelle formation, since sorption depends on the nature of the substrate as well as on the properties of the solute in solution which alone determine micelle formation. Indeed, the fact that the onset of increased adsorption, point B in fig. 2, occurs at a concentration one tenth the critical value for micelles might imply that micelle formation does not have an influence. Interestingly enough, the data may be handled by the methods used for multilayer adsorption of gases on surfaces. Thus, if the amount adsorbed is plotted versus concentration, a sigmoid type curve is obtained, similar in form to a so-called Type I1 isotherm13 generally attributed to multimolecular adsorption. A monolayer value of about 0.12 mg per g may be estimated from the knee of this isotherm. On the other hand, if a plot is made according to the multilayer gas equation for adsorption to an unrestricted number of layers, eq 38,13 assuming the critical micelle concentration as the saturation concentration, the data above 120 mg per liter fall on a straight line from which a monolayer adsorption of 0.3 mg per g results. These considerations suggest that the monolayer may occupy an area somewhat less than the geometric value of 0.4 mg per g, and that micelle formation may indeed be a factor in the adsorption. the angle between the solid-liquid and liquid-gas interfaces, measured through the liquid. The value of the equilibrium contact-angle is uniquely determined by the value of the interfacial energies of the three intersecting interfaces and is independent of other forces in the system. When the liquid-gas interface intersects the solid at an edge—as was the case for all the experiments reported in this paper—the orientation of the solid-liquid interface is indeterminate or varies through 90" for a right-angle edge. Mr. Morris has called the angle between the horizontal and the liquid-gas interface for this edge condition a "static contact-angle." I feel that this term is unnecessarily misleading. In the first place, "static contact-angle" sounds too much like "equilibrium contact-angle." In the second place, the magnitude of the "static contact-angle," which I would prefer to call the "supporting angle," is determined by the forces in the system other than those derived from the interfacial energies, hence "contact angle" is misleading. If Mr. Morris had said that the "supporting-angle" is variable and depends on the weight of the particle and size of the bubble and that it has a maximum possible value equal to the equilibrium contact-angle, his discussion would have been more accurate. T. M. Morris (author's reply)—Mr. Kellogg puts forth a reasonable criticism of some of the terminology used in the paper. I agree that the term "static contact angle" may be misleading. Substitution of the term "supporting angle" or "vector angle" may be more suitable. F. X. Tartaron—In this paper the author presents a very interesting mathematical development of the forces present when a mineral particle adheres to an air bubble. Excellent concordance is obtained between mathematical formulation and experimental results. It is the writer's understanding that when the mineral surface presented to an air bubble is greater than the area of contact, the maximum contact angle is obtained. However, this contact angle represents distortion of the bubble, the normal shape of which is spherical or in cross-section, circular. This distortion produces a force that acts in opposition to the force of adhesion between the bubble and particle. Hence, at maximum contact angle, the force of adhesion between bubble and particle is at a minimum for static conditions. However, when the size of bubble is increased, the size of particle remaining the same (and all other conditions remaining the same), the contact angle decreases, the distortion of the bubble decreases and the force in opposition to adherence of bubble and particle also decreases. Thus, there is stronger attachment between bubble and particle. When this situation is applied to actual flotation conditions, it is doubtful that it has any significance. The reason is that the bubbles are normally so much larger than the particles, that in substantially all cases, it is probable that negligible distortion of the bubble takes place. In table IV, the author makes computations for bubbles from 0.50 to 2 mm diam. This is from 0.02 in. to 0.08 in. Certainly the bubbles generated in a flotation machine are far larger than this. The surfaces of the author's bubbles range from 0.79 to 12.6 sq mm. The area of contact of the glass rod (0.15 mm diam) is 0.017 sq mm. Thus, ratio of bubble surface to mineral area of contact ranges from 47 to 741 in round numbers. If we take a %-in. diam bubble and a 65-mesh (0.208 mm) particle of cubical shape, the ratio of bubble surface to mineral area of contact is 2931. Mineral particles do not readily become attached to air bubbles. Taggart has shown that in pneumatic flotation machines collector-coated particles are only temporarily attached to bubbles. They keep falling off and down in the froth but at a delayed rate as compared with gangue. Spedden and Hannan's motion pictures confirm the difficulty of attaching particles to air bubbles. In the agitation froth process, according to Taggart, air is precipitated from the water selectively on to the collector-coated particles and these

Jan 1, 1951

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

In this paper, oxidation potentials measured in the presence of various concentrations of cyanide, ferro-cyanide, and ferricyanide and ethyl xanthate at various values of pH are related to flotation response. Eh-pH diagrams are presented to show that the formation of surface ferric ferrocyanide is probably responsible for depression when cyanide is added. The influence of cyanide on the depression of pyrite with xanthates as collector has been the subject of a number of investigations,'-6 and several theories on the mechanism of depression have evolved from these studies. Wark and Cox7 and Gaudin8 have suggested that the depressing effect is due to a competition of cyanide ion with xanthate ion for the surface. Cook and his colleagues9-11 have explained this phenomenon in terms of competition between hydrocyanic acid and xanthic acid. Sutherland 12 has shown that although both of these theories accurately describe the relation between pH value and cyanide addition at constant collector addition, they fail to describe the relation between pH value and the amount of collector required to cause flotation. Taggart 13 suggested that depression in these systems is due to the formation of a reaction product between ferric ion at the pyrite surface and ferrocyanide ion derived from solution. Majumdar4,6 has attempted to prove this hypothesis by measuring the contact angles of pyrite in the presence of 25 mg per liter ethyl xanthate and different concentrations of potassium ferrocyanide and ferricyanide. In all cases the contact angles were quite high up to pH 10. These results indicate that pyrite should not be depressed by either potassium ferrocyanide or ferricyanide. In view of these facts, Majumdar has assumed that the compound Fe(CN)2 forms at the surface. Gründer and Bornl4 have stated that depression may be due to the formation of the compound K2Fe(II)Fe(CN)6 at the pyrite-solution interface. This compound is thought to be an interaction product between the K2Fe(CN)6-2 ion from solution and the Fe++ ion at the pyrite surface and, accordingly, K4Fe(CN)6 should depress pyrite at least as effectively as KCN. This was proven experimentally, but there was no simple relation between the depression of pyrite and the concentration of either KCN or K4Fe(CN)6 in solution. In view of the many mechanisms that have been proposed for pyrite depression by cyanide, it is apparent that a clear understanding of the phenomena occurring in these systems is lacking. One reason for this may be the fact that the species responsible for pyrite flotation in the presence of xanthate is not the xan-thate ion but rather dixanthogen.15 Since the oxidation of xanthate to dixanthogen is dependent on the oxidation potential of the solution, it would seem that knowledge of these potentials would be a requisite to understanding the pyrite-xanthate-cyanide system. It is the object of this paper to measure both the oxidation potential and pH of the pyrite systems in the presence of various concentrations of cyanide, ferrocyanide, and ferricyanide and xanthate and to relate these values to flotation response. EXPERIMENTAL MATERIALS AND 'TECHNIQUES In the experiments discussed here, pure potassium ethyl xanthate was used as collector, and reagent grade potassium cyanide, potassium ferrocyanide, and potassium ferricyanide were used as depressants. Reagent grade HC1 and KOH were added for pH adjustment. Conductivity water, made by passing distilled water through an ion exchange column, was used in all experimental work. Two natural samples of pyrite were used in the investigation. Sample preparation for flotation included dry grinding with a mortar and pestle and sizing the product to 100 x 200 mesh. Prior to flotation, a 0.75-gm sample of pyrite was added to a solution containing a known amount of depressant at the desired pH value, and the system was conditioned for 4 min. Following this, a known amount of collector was added and the system was conditioned for another 4 min. The pH — termed flotation pH - was measured; the pulp was transferred to a Hallimond cell, and flo-

Jan 1, 1969

By R. S. Kinasevich, D. D. Crabtree, D. W. Fuerstenau, T. P. Meloy, A. L. Mular

This paper presents details of the concept that size reduction in comminution machines takes place by three mechanisms; namely impact, abrasion, and chipping grinding. Experimental evidence is presented to show that each of these comminution mechanisms produces particles of a characteristic size distribution. The actual product from a comminution device has been postulated to be the result of a composite of these three types of mechanisms. Composite curves are demonstrated for a material following the Gaudin-Schuhmann, Gaudin-Meloy, and Rosin-Rammler size distributions. About forty years ago, Gaudin1 showed that grinding may be considered a combination of impact and attrition. Since that time, major efforts have been devoted to studies, both experimental and theoretical, of impact and compression fracture, while attrition has been almost wholly neglected, at least quantitatively. A significant attempt to distinguish between attrition and impact grinding is the work of Tanaka2 in Japan. By means of an instrumented abrasion grinder and a drop weight device, he determined coefficients relating the increase in surface area of various materials to the energy expended by the two methods of comminution. He showed that physical properties of a material determine its response to attrition and impact grinding. By extending his ideas to ball milling, he concluded that to a large extent the basic comminution process in a ball mill is due to the impact of falling media. Because Tanaka used a mass of particles in his drop weight device, however, he did not obtain pure impact size reduction. The purpose of this paper is to define basic mechanisms of size reduction, to verify that each has a characteristic and meaningful distribution curve, and to show how a combination of these mechanisms results in a size distribution very like that produced by grinding mills. DEFINITION OF COMMINUTION MECHANISMS The product from a grinding device will be considered to be the sum of the contributions resulting from the action of three basic mechanisms of size reduction which operate in a ball and rod mill. These are impact grinding and two forms of attrition, namely abrasion grinding and chipping grinding. Each is defined below. Impact Grinding: If a straight line is obtained when the cumulative weight fraction finer than a given size is plotted on log-log paper as a function of that size, the slope of the line is defined as the distribution modulus. A number of investigators 3,4,5 have shown that impact fracture of single specimens of different composition and shape results in a distribution of daughter particles that exhibit a distribution modulus of unity. This has also been true for materials broken in slow compression. Moreover, under slow compression or impact fracture conditions, theoretical distribution equations6*' which approximate the Gaudin-Schuhmann distribution when the distribution modulus is close to unity agree reasonably well with the experimentally obtained products. If one now considers a comminution process where many impact or compression fracture events are taking place, then any rational summation of the individual events yields, in the finer size range, a Gaudin-Schuhmann distribution whose distribution modulus is also unity. If the size distribution of the product from a mill has a distribution modulus of unity, it is implied that particles are fractured in a manner which results in a comminution event having a distribution like that obtained for the breakage of single specimens. These impact events would occur when a particle is smashed between balls or between a ball and the mill liner. Attrition, on the other hand, is that form of grinding which cannot be called impact grinding, and comprises both abrasion and chipping grinding. Abrasion Grinding: Abrasion is a form of surface wear resulting when particles rub against each other or against balls or the mill liner of a tumbling mill, somewhat like a piece of wood being rubbed over a piece of sandpaper. The surface wears at a more or

Jan 1, 1964

By R. E. Pray, M. C. Fuerstenau, J. D. Miller, B. F. Perinne

Quartz cannot be floated with potassium amyl xanthate as collector at any pH. Complete flotation is achieved with certain minimal additions of amyl xanthate and Pb from pH 5.8 to 8.5 and with amyl xanthate and Zn from pH 7.5 to 8.1. The active species of these metal ions responsible for activation are shown to be PbOH and ZnOH No flotation was effected when Cu and Mg were added as activators. Previous work1,2,3,4,5,6,7 has shown that quartz may be floated in the presence of sulfonate or soap when the following two criteria are met: 1) some portion of the added metal ion has hydro-lyzed to its first hydroxy complex, and 2) a precipitate of the metal-collector has formed. Since sulfonate and soap, two substantially different anionic collectors, conform to these criteria, this phenomenon may well be expected to be of a general nature. In this view then, it would appear that any anionic collector, including xanthates, should also float quartz when these two criteria are met. The object of this paper is to demonstrate the conditions under which quartz will respond to flotation in the presence of potassium amyl xanthate. EXPERIMENTAL MATERIALS AND TECHNIQUES Quartz was prepared by leaching the sized sample (48 x 150 mesh) with HC1 until no iron could be detected in the leach liquor. Conductivity water, made by passing distilled water through an ion exchange column, was used in the investigation. Potassium amyl xanthate was selected as collector. Purification of this reagent was effected with the following method: 1) a laboratory sample of potassium amyl xanthate was dissolved to saturation in hot ethyl alcohol and filtered several times while still hot, 2) the filtrate was cooled, during which time potassium amyl xanthate crystallized out of solution; the xanthate was then separated from the solution by filtration, 3) the amyl xanthate was dissolved in fresh hot ethyl alcohol and filtered, 4) step (2) was repeated, 5) the crystalline potassium amyl xanthate was washed with ethyl ether five times, and 6) the xanthate was dried for 24 hours under continuous vacuum. Inorganic impurities are insoluble in hot ethyl alcohol whereas dixanthogen is soluble in ethyl ether. Five-gram charges of 48x150 mesh quartz were conditioned for 3 min in 125 cc of solution containing the reagents and were floated with the small-scale cell and procedure described in previous papers.478 Reagent grade HCl and NaOH were used for pH adjustment, while reagent n-amyl alcohol was added as frother. Frother addition is necessary with this technique, but it should be mentioned that data, similar to those shown in Fig. 2 for 1 x10-4 mole per liter Pb++ and 1 X 10-4 mole per liter AX-, were also obtained with a Hallimond cell in the absence of frother. EXPERIMENTAL RESULTS Clean quartz could not be floated with amyl xanthate at any pH. Flotation was achieved under certain conditions, however, when various metal ions were added to the system. The activating ions used were lead, zinc, copper and magnesium, added as reagent grade chloride salts. Lead as Activator: The effect of Pb++ as activator was examined at five different additions of amyl xanthate (Fig. 1, 2 and 3). When 1x 10-5 mole per liter xanthate was added in the presence of 1 x 10"4 mole per liter Pb ++, no flotation was obtained at pH 6.5 or below, while a recovery of about 60% was effected at pH 7. When the pH was increased to 7.8, no flotation was again obtained, Complete depression was noted from pH 7.8 to about 10, at which point a recovery of about 30% was achieved. At about pH 11, the precipitate of lead amyl xanthate which was present below this pH dissolved and complete depression occurred. When 2.5x 10-5 mole per liter xanthate was added with the same lead addition, no flotation was obtained at pH 5.8 and below, whereas a recovery of

Jan 1, 1965

By Oscar Johnson

The industry has been canvassed for information on methods of handling grinding steel at milling plants, an important subject overlooked in most descriptive articles on milling plant arrangements. The data collected are offered in this paper as a guide to designers and planners who have loading and storage problems to solve. WITH capacities of milling plants expanding from a few hundred tons per day to many thousand, replenishment of grinding media is an important phase of mill operation. Information on this aspect of milling procedure, solicited from several mining companies, is summarized here. At International Nickel's Copper Cliff concentrator the normally strenuous job of changing rods is considerably simplified by the use of the Garrow rod charging machine, Fig. 1. The overhead crane, Fig. 2, brings a bundle of rods, weighing four to a ton, to the loading platform of the machine, positioned in front of the mill. The rods roll one by one down the slight incline of the platform onto a conveyor which sends them into the mill. A modified version of the Garrow machine, Fig. 3, is used at the Federal mill of St. Joseph Lead Co. An overhead crane places the portable steel structure in position in front of the discharge end of any one of the five 9x12-ft rod mills. It then raises a bundle of about twenty 3-in. rods to the discharge platform of the charging machine. The rods are 11 ft 7 in. long and weigh about 278 lb. By tripping a lever the operator allows one rod at a time to roll onto the inclined rollers. A slight push by the operator sends the rod smoothly into the mill. The Union Miniére du Haut Katanga rod charging machine, Fig. 4, is mounted on a pair of wheels at the forward end with swiveling castors at the back for steering it to position for use or storage. Wheels are equipped with locking brakes to maintain the machine in position while it delivers rods to the mill. The device is made in two parts. The first consists of the base, which carries the supporting wheels, grooved rollers, and drive. The other part serves as the rod deck, which can carry up to six 3-in. x 12-ft rods. A short boom built integral with the base reaches through the mill trunnion and delivers the rods into the mill barrel. Three grooved guide rollers ,are provided to receive the rods from the sloping charging deck. One rod at a time is fed to the rollers and thus delivered into the mill. Two of the rollers are driven by 11/2-hp motor through a V-belt drive. The third roller is driven off the first by a single V belt in a groove below the rod groove and on the same center line. Rods are delivered to the charger by an overhead crane in bundles of three to six as needed. The power is supplied by cables plugged into power outlets, suitably located near the rod mill discharge end. The Oliver Iron Mining Co. uses a rod charger, Fig. 5, suspended from a monorail by cables. It is constructed .of structural steel and has two decks, each capable of carrying five rods of 4-in. maximum diameter. The rods ride in grooved free-turning rollers. Legs extending down from the mill end of the mechanism act as supports when the unit is ex-

Jan 1, 1955

By A. M. Gaudin, T. P. Meloy

This paper presents the theoretical derivation of the distribution equation for single fracture. The equation is shown to accord with experiment. Several empirical size-distribution equations have been proposed to describe the size of broken materials. Among them are the Gaudin-Schuhmann equation and the Rosin-Rammler equation.3*4 The Gaudin-Schuhmann equation relates the relative fraction of undersize to relative size —. In this relationship Mo is the total weight of crushed sample, M(x) the weight of undersize contained, x the size considered, K a quantity called the size modulus, which together with the slope parameter, m, characterizes the product. The slope parameter is usually close to unity. The Rosin-Rammler equation relates the same relative fraction of undersize to the relative size modulus by an equation of different form In this relationship, e is the base of natural logarithms. In the fine size-range when m is unity both equations give Gilvarry and Berstrom have sought a size distribution equation having a sound theoretical basis. They have come to an equation taking into account the effect of volume, surface and edge flaws in fragmentation. If the effect of the volume and surface flaws is neglected, the equation taking only edge flaws into account has the form of the Rosin-Rammler equation. In the present paper, a theoretical size-distribution

Jan 1, 1962

By J. H. Schulman, M. H. Buckenham

Although much interest has been taken in the use of mixed collectors in flotation, this investigation is probably the first in which oppositely charged collectors have been considered. The results obtained and the techniques used could well be applied in the behavior of other mixe the behavior of other mixed collector systems and could possibly explain the effective flotation results which have been reported when various collector combinations have been used. Collectors conventionally employed in flotation systems are soluble in water, but many of their homologs of longer hydrocarbon chain length are insoluble. The properties and behavior of these insoluble compounds may be conveniently studied when they are spread as monolayers at the air/water interface on the Langmuir-Adam trough. Although some extrapolation is involved in applying effects observed at this interface to the solid/liquid interface where the action of collectors is most important in flotation, the technique nevertheless is most useful for the study and interpretation of flotation phenomena. Of particular interest in this field are investigations on the insoluble monolayers of the fatty acids, alkyl sulfates and alkyl amines; the penetration of these insoluble monolayers by soluble surface active components and their interaction with metal ions from the underlying solution.'-' Unfortunately, the insoluble xanthates and related compounds do not exhibit stable monolayer properties due to rapid oxidation at the air/water interface and, therefore, cannot be studied directly in this way. These approaches have enabled Leja and Schulman8 to suggest mechanisms explaining dynamic attachment of mineral particles to air bubbles involving collector-frother interaction and the adsorption of collectors onto mineral surfaces. In this and other papers 9-12 considerable evidence is presented to support the concept of collector-frother and metal ion-monolayer interaction and to emphasize the value of monolayer techniques in elucidating the flotation system. Molecular interactions resulting from monolayer penetration are determined by the nature of and the association between both the polar and nonpolar groups of the two molecules concerned. Stable complexes are formed on penetration when interaction takes place between both the polar and nonpolar groups, and although the resultant complexes are not necessarily true chemical compounds, the properties of the molecularly associated complexes are very different from those of the individual components. Shortening of, or changes from, a straight hydrocarbon chain in one of the reacting molecules reduces considerably the stability of the associated complex as do changes in temperature, pH and substrate composition. The effect of hydrocarbon chain length becomes most important when observed effects are to be applied in flotation studies where soluble homo-logs and, therefore, much shorter hydrocarbon chain lengths are used. The present investigation is based on Langmuir trough work carried out by Bowcott,l3 who studied the penetration of insoluble amine monolayers by potassium lauryl xanthate. Bowcott observed that cetyl amine monolayers were readily penetrated by xanthate, but found that with time the high film pressures resulting from penetration and association fell off, indicating decomposition and ejection of xanthate from the film. This decomposition was pH dependent. Of particular interest in this work were studies on an equimolecular mixed monolayer of the fully N-substituted ammonium compound, for example, cetyl trimethyl ammonium iodide (CTAI) and potassium cetyl xanthate. Despite the fact that the CTAI is water soluble and the xanthate on its own is incapable of forming a monolayer, the mixed film was stable up to a surface pressure of 67 dynes per cm. In this condensed state the xanthate was stabilized against atmospheric oxidation, although iodine injected beneath the film caused an immediate collapse. The stabilizing of xanthate against oxidation by association with an alkyl trimethyl ammonium bromide (TAB) could be important in flotation, and a study of the performance and behavior of mixed collectors therefore was undertaken. A modified Hallimond tube operated under controlled conditions was used to evaluate flotation. Surface tension measurements were made to investigate the properties of the soluble

Jan 1, 1963

By R. B. Tippin, J. S. Browning

The feasibility of multistage heavy liquid cyclone beneficiation of spodumene was successfully demonstrated in this investigation. The indicated recovery for the multistage circuits was mathematically estimated as continuous test runs would be necessary to determine the true influence of a middlings recycle. In the three-stage circuit, both the concentrate analysis and spodumene recovery exceeded 90%. The two-stage circuits were unable to simultaneously yield an acceptable produce with a satisfactory recovery. With the Nation's growing industrial mineral requirements, the U.S. Bureau of Mines has undertaken the task of investigating new mineral dressing techniques and improving the existing concentrating methods of ore beneficiation. One such project is the gravity concentration of minerals using heavy liquid separation (HLS).1-3 Heavy liquid separation is based on the simple principle that a material, whose specific gravity is less than the density of a solution, will float on top of the liquid and a material with a greater specific gravity will sink. Heavy liquid separation techniques were developed in the laboratory over half a century ago but the process has yet to be commercially adopted. In the past, commercial operations have failed because of numerous problems.4 Industrial interest has recently been reactivated because of engineering advances. The number and quality of the heavy liquids suitable for mineral processing has increased while the cost has sharply dropped. New and improved liquid recovery systems5,6 are under consideration with present emphasis on 1) solvent extraction, 2) thermo-evaporation, and 3) water washing. The toxicity of the heavy liquids7 presents a problem, but throughout the chemical and mining industries, many similar hazardous materials are safely handled. A distinct advantage of using heavy liquids is its ability to provide a near perfect separation of fine-sized mineral particles even when the specific gravities are close. Effective separations have been made on minerals down to 200 mesh and with a density

Jan 1, 1968

By D. L. King, B. A. Schepman

A number of unique thickener mechanism designs with the operating results achieved in various installations are discussed. These unusual designs include new types of raking arms, lifting devices, and feed-wells employed to obtain higher underflow densities and minimum operator control or improved overflow clarity. Several installations of large thickeners in which the underflow lines pass upward through the center to eliminate tunnels are reviewed. The location of the lines and pumps with the maintenance and operating records are included. Several types of automatic lifting device controls are used to provide storage in the thickener. Some of the installations employing underflow density controls are achieving excellent operating results. Thickeners or sedimentation basins are used in minerals beneficiation plants for three basic purposes. These are 1) Concentration, or to raise the density of a liquid-solids slurry. 2) Clarification, or removal of finely divided solids from a slurry. 3) Hydroseparation, or removal of a specific fraction of solids from a slurry. The concepts that control the size and configuration of basins used for these purposes are different. Likewise, different equipment is properly employed to accomplish these various objectives. The pattern of liquid and solids flow in the clarification zone of a sedimentation basin used for thickening or raising the density of a pulp normally is not critical. That is, the solids normally will drop or plunge to the bottom quite readily, and the liquid separates itself from these solids so as to overflow the basin in a relatively clear condition. When clarification or removal of finely divided particles is to be accomplished, the flow patterns in the basin are extremely important. The velocity of the incoming stream must be dissipated and the liquid flow distributed ]In the tank so that the feed is not divided into streams that flow directly to the overflow carrying the fine:ly divided solids that are to be removed. Similarly, the configuration of the arms that provide the raking and thickening of the pulp is not as important in basins primarily devoted to clarification. In thickening, the design of the arms and the manner in which the pulp is moved or worked, can contribute appreciably to the thickening process and achievement of a higher density underflow. THICKENING As the grade oi ore bodies throughout the world generally decreases, and chemical processes, hydro-metallurgy, and other more refined methods of minerals beneficiation are required, problems in thickening become more acute. Difficult operating conditions, previously encountered only occasionally in thickening metallurgical slurries, or clarifying process streams in the mill, are now the rule rather than the exception. Fig. 1 illustrates some of the different types of raking arms that are successfully achieving the desired goals in pulp thickening. The conventional raking arms consist of a relatively simple truss with the blades attached to the bottom chords. This design provides proper Concentration of many 'normal' slurries. A modificaiion of this design as shown in the center is being used with excellent results on large thickener basins where it is desirable to have a

Jan 1, 1962

By M. Earl Volin, M. Adnan Goksel

The hydro thermal agglomerating process described (U.S. Pat. No. 3,235,371, 1966)* is different from the cotzventional high-temperature processes. Green pellets or briquettes made with mixtures of moist iron-ore concen- trates and burned or hydrated lime fines are hardened in an azttoclave at 198°C and under steam pressure for 112 to 5 hr, with or without previous drying. Under the neutral conditiotzs, oxidation of magnetite does not occur as in the conventional processes. The resulting azzlomerates-- are hard and resistant to abrasion and weathering, and they do not disintegrate in the Under test. The equipment required for the process is not complicated, and favorable operating costs are projected because of the low temperamre and efficient heat recuperation. A low capital investment makes the process attractive even when the annual pellet production is below a million tons.

Jan 1, 1970

By R. Adair, W. R. Hudspeth, W. T. McDaniel

D. C. Ralston (U. S. Bureau of Mines, Washington, D. C.)—Flake mica can be beneficiated by a wider variety of methods than almost any known mineral. However, most of these methods are not recorded. It is therefore a pleasure to find a written record of another application of a new method of beneficiation. The fact that the Humphries spiral gives a concentrate containing the mica flake mixed with the slime and small grit has forced a further separation by screening, in which the finest of the flake will pass through. Fine flake has its uses and if mixed with slimes or clay is recoverable reliably only by flotation. This loss must be charged against the spiral. On the other hand, the new muscovite mica paper recently brought out by General Electric Co., can be made only with grit free flake larger than 50 mesh. It is of interest to learn from the authors whether the washing of the flake on a 50 mesh screen can take the grit content below 1 pct, to meet this new demand. Must another degritting process be used on the flake, like electrostatic treatment of the dried material? H. D. Snedden (The Humphreys Investment Co., Denver)—When recovering mica in the spiral, gangue particles coarser than 16 mesh can be rejected more readily when operating at a high pulp density (35 pct solids) or a low volume rate (—15 gpm). A launder screen as used by Lehigh Coal Co. has proved very satisfactory to partially dewater the mica concentrate and to reject any gangue slimes it may contain. Under certain conditions mica cleaning in the spiral can be improved by reagentizing the spiral feed, the object being to insure wetting of the gangue minerals and to agglomerate or prevent wetting of the mica particles. R. B. Adair (authors' reply)—The grit content of the mica recovered on Humphreys Spirals varies with the fineness of the launder screen used to clean the mica product. Normal practice has been the use of 80 mesh screen, which will produce a mica product of acceptable grit content for most purposes. By using a 50 mesh screen, and increasing the screen area per ton of feed, it is felt by the authors that a mica of 99 pct plus purity could be obtained. This statement would hold only in the case of mica feed which was well de-laminated, and free from "clay balls," (semikaclinized feldspar), which tends to follow the mica through the process.

Jan 1, 1952

By D. C. Seidel, E. F. Fitzhugh

The cementation of nickel from acidic solutions by metallic iron is discussed. The cementation is carried out in pressure vessels at temperatures above 100°C. The results from bench scale studies on variables such as retention time, temperature, and solution pH are presented. A continuous processing flowsheet is proposed. Cementation is one of the oldest known hydrometal-lurgical reactions. The cementation of copper on iron was recorded by 'Paracelsus the Great' in about 1500,1 and by 1600 the technique was being used to recover copper at the Rio Tinto operations in Spain.2 At least one early author felt that cementation may have been one of the primary reasons for the alchemists' belief in the transmutation of metals.' The early writings state that when iron was placed in the clear waters from some mountain springs, the iron disappears and copper is found in its place. To the alchemists this may well have been one of the most convincing proofs that transmutation could and did occur. Since these early times the recovery of copper from acidic solutions by cementation has been practiced in plants throughout the world. Probably the cementation technique in some form has been common to more copper mining and milling operations than any other single recovery process. During current hydrometallurgical extraction studies, which were sponsored by the Republic Steel Corp. at the Colorado School of Mines Research Foundation, Inc., it was found that under the proper conditions, metallic nickel could be cemented from acidic solutions by powdered iron. The reactions are apparently similar to those that occur during the cementation of copper, but the nickel cementation had not been anticipated because iron and nickel are nearly adjacent in the electromotive series. The potential difference between iron and copper is approximately 0.78 v, while the potential difference between iron and nickel is less than 0.21 v. The cementation of nickel with iron at room temperature is almost negligible, but when the reaction is carried out in a closed vessel at temperatures in excess of 100°C, the nickel can be cemented almost quantitatively. The part played by this discovery in a practical method of nickel recovery is set forth in a separate paper.3 The following paragraphs are a discussion of experimental studies that were made to investigate this cementation reaction. The technique has been designated the HTC or High Temperature Cementation procedure. The cementation work was part of a study on hydrometallurgical techniques for the extraction of nickel from the garnierite or silicate type nickel ores. A hydrothermal extraction procedure had been developed, and this technique produced an acidic pulp or solution that contained both nickel and appreciable amounts of magnesium.3 Small quantities of ferric and ferrous iron were also present along with cobalt, manganese, and chromium. The potential for recovering the nickel from these acidic solutions by ion exchange or solvent extraction did not appear to be promising because of the relatively high magnesium content. The Ni ++ and Mg++ have nearly identical ionic dimensions and tend to be co-absorbed or extracted during ion exchange or solvent extraction treatments. Preliminary tests indicated that the nickel could be precipitated as a sulfide when using the high pressure H2S precipitation technique developed for the Moa Bay operations of the Freeport Nickel CO.4 This technique gave good recoveries, but the nickel sulfide product requires considerable additional processing before a marketable form of nickel is realized. The process also requires clarified feed solutions, and the solids-liquid separations on the garnierite residues are difficult. A program was initiated to investigate alternate procedures that might shorten the route to a marketable nickel product, and hopefully also permit bypassing the difficult solids-liquid separation steps required for the H2S precipitation technique. It was during these studies that the nickel cementation reaction with iron was encountered. EXPERIMENTAL EQUIPMENT AND PROCEDURES The bench scale precipitation tests which were conducted during this experimental program were made in 2-liter stirred autoclaves.* A photograph showing the form and arrangement of the autoclave

Jan 1, 1968

By S. A. Falconer

The benefication of non-sulfide or non-metallic minerals by froth flotation has been practiced commercially for at least 30 years. In the pioneer stages of development of this flotation art, most of the promoters investigated were fatty acids or soaps that had been found to float sulfide minerals. When commercial application of froth flotation started to assume importance in the late 1920' s and 1930's, a great deal of fundamental research work was done in the laboratory (11-14, 19, 22, 23, 26, 29, 32). Among other subjects, this work included studies of the relative effectiveness of the different fatty acids and their salts on the flotability of various non-sulfide and non-metallic minerals. The successful present-day use of fatty acid and petroleum sulfonate type promoters results from better understanding of the principles of non-sulfide flotation plus the developnent of effective selectivity modifying agents. Many investigators have contributed to a greater appreciation of the numerous interrelated factors involved in fatty acid and soap flotation of non-sulfide minerals. Of these, mention should be made of recent fundamental studies by Tagart and Arbiter (31); Wolstenholme and Schulman (35); Schulman and Smith (28); Wark and his co-workers(3,4) Du FIietz(l5); and J. Leja(24). According to J. Rogers, K. L. Sutherland and E. E. and I. W. Wark(27), "the basic requirements for a 'collector' in flotation are that it shall form a film over the surface of the mineral to be floated and that this film should be oriented so that hydrocarbon (non-polar) groups form the effective surface." When modified in that way, the mineral surface possesses some degree of air avidity. All investigators would agree with Taggart, Taylor and Ince (33)that such a film can be produced only if the collector compound contains an active (polar) group to be attracted to the mineral surface, as well as the non-polar group. The commonly used anionic-type promoters for non-sulfide, non-metallic and metallic oxide ores include: 1. Fatty acids, and their soaps and emulsions. 2. Fatty acids residues — "pitches" and "foots." 3. Sulfonated and sulfated fatty acids and so-called "petroleum sulfonates." 4. Paraffin chain salts — sulfonated and sulfated higher alcohols. 5. Cyclic organic acids or salts, such as naph-thenic acid and naphthenates. Certain fatty acids are more effective than others as promoters for particular minerals under certain specified conditions. Some of the more costly fatty acids or collectors of the paraffin chain salt type may give improved selectivity in some instances and are being used commercially in a limited number of operations. Local availability may dictate the choice of the fatty acid used — for example, the use of palmoil fatty acid for flotation of copper carbonate ores in the Belgian Congo. Coconut oil fatty acids have also found limited use.

Jan 1, 1961

By Jack White

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the l l-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the ll-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 1% in. and a further gain of work accomplished of 2.1 pct when we went from 1 % in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By Jack White

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the l l-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the ll-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 1% in. and a further gain of work accomplished of 2.1 pct when we went from 1 % in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By J. M. Cases

Study of the normal interaction between flotation collectors and the silicates through measurement of the electrokinetic potential carried out by the streaming potential and flotation recovery methods indicates that the interaction is wholly 07 partly of electrostatic origin. In some cases, it partly consists of the chemical affinity of the polar head of the collector for the active surface sites. Reagents composed of a polar head and a paraffinic chain and forming micelles in solution were used as flotation collectors. They are sodium dodecylsulfate (anionic collector) and dode-cylammonium chloride (cationic collector). The adsorption isotherms on biotite of alkylamine chloride having 10 to 18 carbon atoms in their chain all show a sharp change in slope relevant to the condensation of alkylammonium ions into bidimensional phase.1-1 Condensation is possible because of the relative homogeneity of the adsorbing surface and of the very strong lateral binding forces through Van der Waals type bonds in the adsorption layer. The different isotherms are seen to coincide when, instead of the equilibrium concentration, the under-saturation ?µ = kT In C/C, is plotted in abscissa, C representing the concentration of the solution with which the adsorbed layer is in equilibrium and C, the saturation concentration of micelles in solution or. critical micellar concentration. More particularly, the undersaturation ?,u1*, corresponding to a surface coverage 6 of 0.5, is equal to Since silicates immersed in aqueous solutions show a surface charge which is a function of pH, it is of primary importance to study the influence of this parameter on the adsorption of collector ions acting as Strong electrolytes in the pH range where they are utilized. The aim of this present paper is to determine by studying electrokinetic phenomena and recovery curves effectuated with a cationic collector and a anionic collector, the nature of the normal adsorbate-adsorbent interaction. Tables 1 and 2 sum up part of the results previously obtained'H11 for the determination of the point of zero charge of the silicate studied by the streaming potential method. Throughout the remainder of this paper, the surface charge will be considered positive for all values of pH below the point

Jan 1, 1971

By R. W. Smith, R. W. M. Lai

Although several recent papers have shown that equilibrium contact angle in dodecylamine solution-quartz systems can often be directly correlated with flotation behavior, it appears that under conditions of relatively high amine strength and pH such a correlation can't be made. Under these conditions, when there are precipitated amine dipoles present, it appears that a dynamic contact angle formed immediately after introduction of fresh gas bubbles into the system is also of importance. The dynamic angle decreases in magnitude as a function of time. Apparently the decrease is the result of changes in interfacial tensions (liquid-gas and solid-gas) as a function of time. Dynamic angles of the type described can't be observed in trimethyldodecylammonium chloride solution-quartz systems. Several recent papers1-3 have shown quite conclusively that contact angle data can often be correlated directly with flotation behavior in cationic surfactant solution-quartz systems. At the same time, Iwasaki, Cooke, and Colombo4 have shown data for a dodecylammonium chloride solution-goethite system where, in the vicinity of pH 12 and a collector concentration of 10" M, 100% flotation occurs when equilibrium contact angle appears to be zero. A clue to this discrepancy can be found in the work by smith2 who reported the occurrence of dynamic, time dependent, contact angles in a quartz-dodecyl-amine solution system at relatively high pH values and at surfactant concentrations greater than or equal to 10-4 M. It was found that under these conditions, if a fresh bubble was brought into contact with a properly conditioned and polished quartz surface, a large angle was obtained. The large angle, however, instantly started to diminish in magnitude, sometimes to a smaller stable angle and sometimes to its complete disappearance. Also, if the bubble was preconditioned for some minutes (or sometimes even seconds) in the same amine solution, the smaller or sometimes zero stable angle was obtained when the bubble was brought in contact with the quartz surface. One is immediately led to the thought that perhaps it is these dynamic angles when they exist, not stable angles, that are of primary importance in flotation. The present investigation attempts to further study the relationship between contact angle and flotation

Jan 1, 1967

By E. B. Fitch, E. C. Johnson

The operating behavior of liquid-solid cyclones is outlined, together with the nature and range of the process results obtainable, to serve as a background for engineers wishing to consider application of this new process tool. BY now most engineers are familiar with the liquid-solid or Dutch State Mines cyclone. However, it should be helpful to know exactly what it is that the equipment does and what its limits are. Without going into cyclone theory, this paper will describe the operating characteristics of Dutch State Mines cyclones. These are manufactured under license in this country and sold under the trade-marked name of DorrClone. The physical construction of the liquid-solid cyclone has been covered in many papers,'-' the DorrClone in particular being described in some detail by Weems. Fig. 1 shows the unit in cross-section. The feed enters at C. The coarse, heavy particles are thrown centrifugally to the periphery and make their way down the wall to the apex where their rate of discharge as underflow is controlled by an adjustable rubber apex valve. As the apex diameter is decreased the solids build up behind the valve, producing a denser underflow. Meanwhile the fine particles are swept into the upward flowing vortex stream which exits as overflow through the vortex finder, F. Flexibility to produce the specific result desired in a particular process is achieved by providing means for varying the areas of the entrance, vortex discharge, and apex discharge. The entrance area may be varied by insertion of special shims. Vortex discharge area may be changed by use of different-sized vortex finders which are interchangeable. Similarly, the different sizes of apex valves are interchangeable and in addition each apex valve is variable down to about 60 pct of its maximum diameter. A most significant primary distinction to make is that although liquid-solid cyclones have been sometimes called thickeners, they actually are classifiers, and very potent ones. They are almost never thickeners in the special sense that many metallurgical engineers understand the term. There would be no profit in quibbling over the definition of a word, but when the application of cyclones is considered, it will help to understand the difference be- tween two mechanisms, one of which will be called classification, and the other thickening. In what is called thickening the fine solid particles present in the feed hold together by surface attraction during the sedimentation process. The loose network of particles thus held together constrains all particles to settle at approximately the same rate, the larger ones dragging the smaller ones down. As a result, pulp settles with a sharp line of demarcation between solids and a relatively clear supernatant liquid. Essentially all the solids, regardless of their fineness, pass into the thickened underflow, and a clear overflow is separated. In classification, on the other hand, the interparti-cle forces are relatively insignificant as compared to the settling force on the individual particles, and are insufficient to prevent independent movement of the particles. The coarsest, heaviest particles settle most rapidly through the pulp, passing more slowly settling fines. Particles coarser than the mesh-of-separation essentially all settle into the underflow, but if the feed contains any particles finer than the mesh of separation, at least part of them will appear in the overflow. A clear supernatant or overflow can be obtained only if there are no undersize particles present in the feed. Thus it will be seen that classification is impossible under ideal thickening conditions. The finer particles are pulled down at essentially the same rate as coarser particles, and there is no separation on the basis of particle size. The surface attraction holding the particles together in a thickener is usually feeble. Whenever the sedimentation force on any particle is strong with respect to the interparticle forces, that particle can pass through the tenuous structure and settle independently. There are at least four ways of making the sedimentation force strong, with respect to the interparticle forces, and obtaining classification. First, and most obvious, the particles may be large and heavy. Thus coarse sands settle out in a beaker or Dorr thickener ahead of the rest of the thickening solids. Second, the interparticle forces may be altered by physicochemical means; i.e., it is often necessary to add dispersing agents to destroy the interparticle forces and permit classification to take place. Third, the interparticle forces may be reduced by dilution of pulp. It is well known that to obtain the most efficient separation of

Jan 1, 1954