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By W. Mitchell, B. H. Bergstorm, C. L. Sollenberger

A unique compression testing machine was constructed to load individual 1/8 to I-in. spheres of glass, etc., at rates from 100 to 100,000 lb per min. During loading the applied load was continuously plotted us the deformation of the sphere-platen system. The load-deformation curve conforms to the theory of elasticity. The spheres were surrounded by steel retaining rings or by gelatin (to prevent secondary fracture). The product size modulus was determined from screen analyses. The data shows: 1) Energy input per unit mass is inversely proportional to the product size modulus (or directly proportional to the surface area per unit mass). 2) Increased loading rates slightly increase the energy input per unit mass required for fracture without affecting the preceding relationship. Theory and experiments show that much of the input strain energy is transformed into kinetic energy of fragments at fracture, this being a reason efficiencies based on "surface energies" are exceedingly small. This large kinetic energy component is usually dissipated as heat but may be partially reclaimed because impact of fragments against the surface of steel retaining rings causes additional comminution. There are currently several theories of crushing, all of which can be derived from the general energy equation of Gillilandl as was pointed out by Mitchellet al in 1954'and more recently by Rose and Sullivan. If the exponent of the equation dE-x-"dx is assigned values of 2,1, or 3/2, one obtains by intergration between suitable limits, the equations of Rittinger, kick,' and Bond,' respectively. Other investigators prefer to let the exponent be a variable parameter. To help resolve the controversy, the Allis-Chal-mers Research Laboratories set up a series of carefully controlled experiments in which the level of energy input required to induce the comminution of a single particle by slow compression loading was very accurately measured. The energy input requirement, in conjunction with the size analysis of the fragments produced at fracture, yielded an accurate energy-product size relationship for the conditions studied. APPARATUS The experimental apparatus was designed to apply a compressive load on opposite sides of an "ideal particle." Spherical shapes were chosen, because point contact loading is typical of the mechanics of conventional crushers. The bulk of the experimental data was obtained with glass spheres because glass is a fairly homogeneous, brittle material for which the properties follow definite statistical distributions. The spheres ranged from 1/8 to 1-in. diam. The compressive load was applied to the specimen at a uniform rate by a loading mechanism (Fig. 1) which acted through a system of levers. The uniform rates, ranging from 100 to 100,000 lb per min, were obtained by metering water into a bucket suspended at the end of the secondary level. Point contact was made with the sphere through tungsten-carbide platens. This latter material is extremely hard and has a high modulus of elasticity, thereby having a minimum of distortion when loaded. The actual load on the sphere was measured independently of the lever system by a load cell, Fig. 2, which was designed so that the voltage output of the load cell would be proportional to load. The deformation of the specimen was indirectly measured by a pair of transformers which were attached to opposite edges of the upper platen. The cores of the transformers, which were attached to the lower platen, moved relative to the transformer so that the voltage output was proportional to the deformation of the sphere-platen system. The load was plotted against the deformation on an X-Y recorder. The area under the curve, as shown on Fig. 3, was proportional to the energy input of the

Jan 1, 1961

By R. Schuhmann

Distribution of material in the fine sizes of a comminution product generally is well represented by the empirical equation' y = 100 (x/k)a [1] in which y — cumulative percent finer, x = particle size, a -= distribution modulus, and k = size modulus. Charles3 found that the energy consumption in comminution is usefully expressed by another empirical relation, E = Ak(1-D) [2] in which E = energy input per unit volume of material, A = a constant, k = size modulus based on Eq. 1, and n = a constant; (1-n) is the slope of a plot of log E vs log k. Holmes3 has presented energy equations similar to Eq. 2. The constants a and n in Eqs. 1 and 2 have been shown to depend both on the nature of the material and on the comminuting device. Moreover, Charles showed that within experimental error a and n are a — n+1 = 0 [3] Combining Eqs. 2 and 3, E = A k-a [4] In the first sections of this article it is shown that the energy equation, Eq. 4, can be derived directly from the size distribution equation for the fine sizes, Eq. 1. The derivations are made without assuming any of the specific relationships between energy and particle size which have been common in previous literature. For comminution processes in which Eqs. 1 and 4 adequately represent the experimental data, the constant A in Eq. 4 is found to be a simple and useful inverse measure of grindability. That is, A is the energy consumption per unit volume of comminution product finer than unit size as determined from the straight line portion of the log-log plot of the size distribution. These considerations all lead to a unifying hypothesis of comminution mechanism from which both Eq. 1 and Eq. 4 can be derived. Finally, it is pointed out that this hypothesis raises serious questions as to the significance of the Rittinger hypothesis, the Kick hypothesis, and other theories in which energy numbers are systematically assigned to various size fractions of comminution products in order to calculate theoretical energy consumptions. Derivation of the Energy Equation from the Size Distribution Equation: For simplicity, consider the comminution of 100 volumes of a feed material of relatively uniform particle size. The comminution process may be considered as the summation of many individual and independent comminution events. The extent of comminution is most easily expressed as the number of comminution events, z. In the first derivation, the key assumption is that the characteristics of the comminution events in a given crushing or grinding process are substantially constant and do not vary with the progress of the comminution process. Accordingly the characteristics of an average comminution event may be defined. In one such event, a quantity of energy $E is applied to a single particle of size f and volume $v. The crushing of this particle produces fine particles with a size distribution similar to that given by Eq. 1: yw=100(x/ka)a0 [51 In this equation yo, a0, and k0 are used to characterize the product of an individual comminution event rather than the product of the comminution process as a whole. In using Eq. 5, we will not be concerned with values of x close to the feed size f and will therefore assume only that the equation is applicable to the finest sizes of the material. In 100 volumes of total product, the actual volume of product finer than x from a single comminution event, or dy, is given by The total volume of material below size x, resulting from z events, is then given by y = z(dy) =z(dv) (x/ka)a0 =" Eq. 7 reduces to Eq. 1 when we let a, = a and [8a] z =100/dv (ka/k )1 or k = ka (zdv/100)-1/a [8b] Eq. 8a shows that the distribution modulus of the comminution product is the same as for the product of an individual comminution event. Eq. 8b shows how the size modulus of the comminution product k varies with the extent of comminution as measured by the number of events, z, or as measured by the fraction of the feed actually subjected to com- minution1 zdv/100. The energy input to 100 volumes of total feed, or 100E, is the sum of the energy inputs for all the comminution events: 100E =z (dE) [9]

Jan 1, 1961

By I. C. Edwards, G. E. Agar

INTRODUCTION Many research efforts over the past years have been devoted to the study of energy consumption in comminution and the characteristics of the broken material. The old "laws" have been superseded by more general methods of analysis1,2,3 but these methods have not received much attention outside the laboratory. The work to be described here was done in order to obtain scale-up information on wet rod milling of sylvinite. EXPERIMENTAL A 30 x 51 in. Denver rod mill with a scoop feeder was used for these tests. Ore was fed from a variable speed screw conveyor into the feed box along with saturated brine. The mill was filled to approximately 45% of its volume with 2400 lbs of a mixed rod charge. The minimum possible feed rate was 2.5 tph and the practical maximum for this mill was 8 tph. Individual tests were run by setting the solid feed rate, then adjusting the brine flow rate to give the desired pulp density. After steady state, as evidenced by constant discharge pulp density, had been reached, samples of the discharge were taken at 15-minute intervals. These samples were filtered, washed with alcohol, dried, then screened on Tyler screens using a Ro-tap shaker. The sylvinite feed was regular plant feed that had been scalped at 20 mesh and was tapped from the coarse ore bins. It had been previously passed through a roll crusher set at 3 inches and after scalping was essentially minus 1 in. plus 20 mesh. RESULTS The screen analysis of samples from each series of tests was plotted on double logarithmic scale paper to determine the reliability of the series. In some cases, a faulty sample or screening procedure dictated that the results of one test be discarded. The criterion is a simple one — if the slope of one plot was noticeably different from the others in the series, it was discarded. The only cases that caused difficulty were those in which the brine was not thoroughly displaced from the solids. When this occurred agglomerates were easily identified on the screens and the whole of the sample was retained on 3 or 4 screens. There were no border line cases. The number N given in Table 1 is the number of size analyses that were pooled to form the data shown in Fig. 1. While examining the reliability of the sizing and sampling by plotting the data, it became apparent that there was no significant difference between grinds made at 50 and 60% solids. For the analysis given here, these tests were regarded

Jan 1, 1967

By P. L. De Bruyn, R. J. Charles

THE transfer of kinetic energy of translation into other forms of energy by impact is a fundamental process in most crushing and grinding operations. During and after the impact process the original source energy may be accounted for in any of the following possible forms: 1) Kinetic energy of translation of both the impacted and impacting objects. 2) Kinetic energy of vibration of the components of the impact system. 3) Potential energy as strain energy of the components of the system or in the form of residual stresses. 4) Heat generated by internal friction during plastic deformation or during damping of elastic waves. 5) New surface energy of fractured materials. At any instant during the impact process only the strain energy of the components of the system can contribute directly to the brittle fracture process. If fracture is the desired result, as in comminution, it would seem advantageous to choose or arrange the conditions of impact so that a maximum amount of the original kinetic energy could be converted to strain energy at some moment during a single impact. The present work deals with determination of these desirable conditions for a simple case of impact and application of the principles involved to general cases of impact. Experimental Method: Longitudinal impact of a rod with a fixed end was chosen as the impact system for investigation. The rod was mounted horizontally and the fixed end was formed by butting one end of the rod against a rigidly mounted steel anvil. The rod, of pyrex glass, was 10 in. long by 1 in. diam with both ends rounded to a 6 in. radius. The rounded ends permitted reproducible impacts on the free end of the rod and assured a symmetrical fixed end. Pyrex was selected as the rod material because of the marked elastic properties of such glass and the similarity of fracture between pyrex and many materials encountered in crushing and grinding operations. The frequency of natural longitudinal oscillation of the rod was 10 kc, and thus simple electronic equipment could be used for observation of strain changes occurring in the rod at this frequency. As shown in Fig. 1, impacts on the free end of the rod were obtained either by a pendulum device or by a spring-loaded gun. Relatively heavy hammers (100 to 600 g) of mild steel were used in the pendu- lum impacts, while fairly light projectiles (20 to 80 g) were fired from the spring-loaded gun. One of the main objects of the experimental work was to obtain the strain-time history of the rod as a function of the mass and kinetic energy of the impacting hammers. For this purpose a technique involving wire resistance strain gages and a recording oscilloscope was employed. Five gages were applied at equidistant sections along the rod, and by means of a switching arrangement the strain-time history at any section, and for any impact, could be obtained in the form of an oscillograph with a time base. The equation relating strain and voltage change across a strain gage through which a constant current is flowing is as follows: e = ?v/iRF [1] ? = strain, ?v = voltage change, i = gage current, R = gage resistance, and F = gage factor (from manufacturer's data — SRA type, Baldwin Lima Corp.). With the above equation an oscillograph depicting voltage change vs time on a single trace can be converted directly to a strain-time diagram if a calibration of the vertical response on the oscilloscope screen for specific voltage inputs is available. In the present case the calibration was obtained by photographing precisely known audio frequency voltages on the same oscillograph as that on which a voltage-time trace from a strain gage had been made. Synchronization of the beginning of the single trace with the beginning of the impact was accomplished by permitting contact of the impacting objects to close an electrical circuit from which a voltage pulse, sufficient to initiate the trace, was obtained. The struck end of the rod was lightly silvered for purposes of electrical conduction so that it would form one of the electrical contacts. Markers every 100 micro-seconds on the traces served for a time base calibration. Determinations of the kinetic energies of translation prior to impact were made in the case of the pendulum hammers by measuring the height of fall of the hammer and in the case of the projectiles by measuring the exit velocity from the gun barrel by means of an electrical circuit employing light sources, slits, and phototubes.' During the experimental work it became evident that the time of contact between the impacting object and the rod was an important variable in the impact process. Measurements of the times of contact were made, therefore, for every impact for which a strain-time record was obtained. The time of contact was determined by permitting the impacting components, when in contact, to act as a closed switch and discharge a condenser at relatively constant voltage. The discharge was observed and photographed with a time base on the oscilloscope screen.

Jan 1, 1957

By J. P. Zannaras

Referring to the article by R. J. Charles and P. L. de Bruyn, let us assume that W = weight of glass bar; P = weight of hammer; e = total deformation; K = unit of deformation; K = potential stress energy; E = modulus of elasticity; L = length of the bar; 7 = coefficient of inertia; h = height, ft; V = velocity, fps; and a = cross section area. The portion of kinetic energy which is effective in producing stress energy in a fixed bar struck horizontally is given by the formula" 1 + 1/3 W/P K =1+1/3w/p/(1+1/2w/p)2 P.V2/2g = Ph where 1 + 1/3 W/P ? (1 + 3W/P/(1=1/2w/p)2 [8] Putting e e = W/P =------------ From the above equation it can be seen that the maximum transfer of kinetic energy to stress energy is when e = 0 or W/P = 0 which indicates that the weight of the hammer must be very large as compared with the weight of the impacted rod. Eq. 8 diametrically opposes the conclusions reached by the authors of this article. In fact, if their suggestions were followed to the extreme when e = co when P = 0, there would be no transfer of kinetic energy to stress energy at all, as 7, becomes zero. Eq. 8 presumes that the velocity with which the stress is propagated through the bar is infinite, whereas the authors claim that the compression waves reflected are reaching the struck end of the bar prior to the complete transfer of the kinetic energy to cause such modification of the conditions there as to make them reach the reverse conclusions demonstrated by the above formula. That such interference exists is unquestionably demonstrated by the authors and others. However, if my observations are correct, such interference for this specific experiment and also for practical comminu- tion is insignificant, and the conclusions of the authors are in error and must be reversed to comply with Eq. 8. Eq. 5, w. = AE 2/2, given by the authors on page 51, is derived from the following equation (Eq. 9): K = 1/2Pe, where P = Sa, S = ?E, e = EL, and L = 1. The above formula, Eq. 9, cannot be applied in this case. This formula is applicable for static loads where the load increases from zero up to its final value, P, in such a way that the deformation at different instants is proportional to the loads acting at those instants and actually represents the area of a right triangle in the strain load diagram of base e and height P. The typical photographs shown in Figs. 3 and 4 represent the familiar strain load diagrams, and since the line of the wave marks the existence and intensity of the strain with the unquestionable conclusion that such strain has been caused by the action of a load acting continuously all along the wave until it reached the horizontal axis, the work stored at this point is represented by the area under the wave line and the horizontal axis and not by the area of the fictitious triangle given by the authors. Then if this is correct, even visual estimation of these areas at gage stations given in the typical photographs of Figs. 3 and 4 suffice to contradict the authors' calculation given in Figs. 6a and 6b and Figs. 7a and 7b. The typical photographs presented by Charles and de Bruyn show a considerable variation of the intensity of the strain at different stations but very small variation of areas which actually represent the stress energy at the corresponding stations. And, apparently, by squaring the small quantities, the authors magnified their error tenfold. J. M. Frankland's paperV iscusses the relative strain intensity and not the total energy for different types of impact loading. He states in his paper, "The reader is explicitly warned not to confuse the results in this report with those obtained when the load is applied by a blow as from a hammer. In this case the peak load rises to very large but mostly unknown values. The accompanying large deflections and stresses are the result of high values of P, not of the dynamic load factor n. According to Frankland "the dynamic load factor" is the numerical maximum of the response factor. It therefore appears that the authors followed the same procedure in obtaining the relative strain energy ab-

Jan 1, 1957

By G. E. Agar

The energy required for dry grinding of phosphate rock in ball and roller mills has been analyzed by the method suggested by Charles and others. The findings are not wholly in agreement with Charles' predictions; but, nevertheless, the method of analysis is shown to be sound. Some of the deviation is probably attributable to the characteristics of a dry grinding system. Many laboratory and pilot plant scale experiments have been made to determine the energy requirements for size reduction. This topic was recently reviewed by Agar and Brown' and it was apparent that very little if any data exists in the literature which permits one to apply the Charles 2-Holmes3 analysis to industrial size systems. More recently, the idea has been put forward that the postulate of comminution by impact used by Charles and others was not sufficient to explain size-reduction throughout the grinding operation. The opportunity exists in the grinding operations of International Minerals and Chemical Corp. (IMC) to analyze the energy-size relationship for dry grinding phosphate rock in both ball and roller mills. This situation arises because the mills are used for custom grinding and make products from 50% to 85% minus 200 mesh. BACKGROUND A correlation of particle size in a comminuted material with the energy expended in size reduction was proposed by charles2 to be of the form

Jan 1, 1965

By S. R. Mitchell, M. Weissman, J. H. Brown

It has been demonstrated that for fine grinding operations, the energy input (E) and the product size modulus (k) are related by an equation of the form E = Ak1-n (1) where "A " and "n" are constants.(1) Also, it was suggested that "A" is determined by the comminuting device and that "n" depends primarily on the rock. It was the purpose of this investigation to study these constants by determining the form of an energy-size reduction relationship over an extreme range of energy inputs, and at the same time, to evaluate the minimum particle size produced during grinding. The comminution system consisted of a stainless steel laboratory ball mill in which quartz was ground at 60% solids for times ranging from 1 minute to 700 hours. The mill which measured 10" diameter by 12" long was loaded with about 60 pounds of balls ranging in diameter from 3/8" to 1-1/8" and was driven at 55 r.p.m. Quartz was selected for the tests because it is a brittle material for which size distribution data were expected to show a constant distribution coefficient regardless of product size.(1) This permitted specification of product size by a size modulus alone. The quantities measured were the time of grind (measure of energy input) and the specific surfaces of the products. Since no accurate method is available yet for determining particle size directly in the very fine size ranges specific surface data were used for this purpose. The relationship between these two quantities may be expressed as where Cis a constant which includes a shape factor, p is the specific gravity of the solid,_S is the specific surface and D is an equivalent particle diameter. For a material in which the distribution coefficient does not change with size the term (l/S) may be substituted for k in equation L if the proportionality constant "A" is changed correspondingly. This substitution was made in this work but in addition, screen analyses were made on several products to verify the equivalence of the two methods of size specification. EXPERIMENTAL DETAIL QUARTZ Quartz from two sources was tested. One series of tests was made on a clear, high purity crystalline quartz from Brazil and Arkansas, and another series was run on massive quartz fromNewHampshire. Grinding charges of the high purity material consisted of 2000 gram samples of -3-8 mesh particles, while the charges of the New Hampshire quartz were 500 gram samples of a -35+48 mesh screen fraction. The difference in the sizes of these two series of charges was not significant in view of the extreme size reduction accomplished, and the New Hampshire samples were smaller to permit ginding to finer sizes in shorter times than required for 2000 gram lots. SAMPLING AND ANALYSIS Separate charges were ground for all tests in which the grinding times were less than one hour and for all tests on New Hampshire quartz since these products were difficult to sample accurately without affecting the entire charge. During the long time grinding tests the mill was

Jan 1, 1961

By R. J. Charles

SEARCH for a consistent theory to explain the relationship between energy input and size reduction in a comminution process has accumulated, over the years, an enormous amount of plant and laboratory data. Although some correlation of these data has been possible for purposes of engineering design and for the advancement of research in fracture, there is still great need of a means of predicting behaviour of a solid when it is reduced in size by mechanical forces. The best known hypotheses proposed to describe the energy-size reduction relationships in crushing and grinding stem from a common origin. The present article analyzes problems of comminution in the light of the precepts of this origin. Its object is to reconcile points of difference between these well known hypotheses and to present relationships more widely applicable to comminution studies. Theoretical Considerations: Most existing relationships between energy and size reduction of a brittle solid stem from a single, simple, empirical proposition.&apos; Although this proposition can be demonstrated by observation and experiment, no theoretical derivation is yet possible. Mathematically, the proposition may be stated as follows: dE = -Cdx/xn [1] where dE = infinitesimal energy change, C = a constant, dx = infinitesimal size change, x = object size, and n = a constant. Eq. 1 states that the energy required to make a small change in the size of an object is proportional to the size change and inversely proportional to the object size to some power n. No stipulations are placed on the exponent n in either magnitude or sign. In 1867 Rittinger2 postulated that the energy required for size reduction of a solid would be proportional to the new surface area created during the size reduction. As far as can be determined there is as yet no physical basis for Rittinger&apos;s hypothesis. Rittinger&apos;s hypothesis can be stated mathematically as follows: E, = K(oa-a-0 . [2] Er = energy input per unit volume, K = a constant, <ti = initial specific surface, and o2 = final specific surface. In the size reduction of particles of size x, to particles of another smaller size, x2, Eq. 2 becomes the well known relation: ET = K&apos; {l/x2-l/xx) [3] where K&apos; is a constant. Eq. 3 may be arrived at from the proposition given in Eq. 1 by integrating and by assigning a value of 2 to the exponent n. J dE = J - C dx/x2 E = Kt (1x - 1x) where K&apos; = C. In 1885 Kick3 proposed the theory that equivalent amounts of energy should result in equivalent geometrical changes in the sizes of the pieces of a solid. For example, if one unit of energy reduced a number of equal-sized particles to particles of one half the size, then the same amount of energy applied to the particles resulting from the first test should result again in a size reduction of one half or a final size one quarter the original size. The Kick concept may be expressed as follows: Ek = K" log x1/x2 [41 K" = a constant and E, = energy per unit volume. The expression for Kick&apos;s law may be arrived at by again integrating Eq. 1 and in this case assigning a value of 1 to the exponent n. dE= J - C dx/x Eh = - C In {x/x2) = K" log (x,/x,) where K" = 2.3 C. Application of Kick&apos;s and Rittinger&apos;s laws to comminution has met with varied success. Gross and Zimmerley4 and Piret5 have shown that Rittinger&apos;s equation applies under certain conditions of experimentation. Walker and Shaw6 express the belief that in metal turning and shaping and in grinding of both metals and minerals the production of very fine particles (less than lP) follows Kick&apos;s hypothesis, whereas Rittinger&apos;s concept is valid for the size reduction of coarse particles. For the practical case of crushing and grinding, however, neither of the above hypotheses has received general acceptance. Bond&apos; has lately proposed that since neither Kick&apos;s nor Rittinger&apos;s hypotheses seem correct for plant design work, an energy-size reduction relationship somewhere between the two would be more applicable. The fundamental statement of Bond&apos;s work

Jan 1, 1958

By J. Leja, A. Pomianowski

This paper deals with equimolar solutions of xanthate and alkyl trimethyl ammonium bromide adsorption on copper, nickel and sphalerite powders. After an exposition of the experimental methods used and the results obtained, a thorough discussion leads to the conclusion that some molecular association between xanthates and CnTAB&apos;s does take place in the bulk of the solution as well as at the air/liquid interface. A spectrophotometric study of xanthate solutions alone and their mixtures with equimolar amounts of alkyl trimethyl ammonium bromide (CnTAB) has established the additivity of ultraviolet spectra (190 to 400 mµ) for potassium ethyl xanthate (KEtX) and C12TAB. No obvious effect of C12TAB addition on the decomposition rate of KEtX was observed, the half life of which was established as 29±3 days at 22°C and pH ~ 6. In mixed solutions of longer xanthate homologues (e.g. hexyl and nonyl) and equimolar CnTAB, colloidal precipitates are gradually formed which contain primarily the corresponding xanthate molecules. The kinetics of collector abstraction by copper, nickel and sphalerite powders were studied in 10-5 to 10"3 M solutions of KEtX alone and in equimolar mixtures of KEtX and CI2TAB. Presence of C12TAB caused a two to three fold increase in the amounts of xanthate (equivalent to many monolayers) abstracted by copper and nickel powders (and in their rates of abstraction) but had no effect on the abstraction of xanthate by unactivated sphalerite (equivalent to less than a monolayer). The adsorption of C12TAB itself from the mixed solutions was very low, at most one monolayer only, and was not influenced by the nature of the solid powder used. A change in the rates of abstraction, observed with copper powder, was interpreted as a completion of a multilayer coating (completely enveloping the grains of powder) following which the rate of reaction between solution and substrate becomes controlled by diffusion through this coating. Work of Buckenham and Schulmanl on molecular associations between different collectors has shown unusual synergistic effects in flotation of individual minerals using mixtures of xanthates and alkyl trimethyl ammonium bromides. The objectives of the present work were: 1) to determine whether these molecular associations between the two types of collectors (indicated by strong surface tension lowering in Buckenham and Schulman&apos;s work) take place at the air/liquid interface only or in the bulk of solution as well 2) to establish whether different solids adsorb any species from the mixture of these amphipatic molecules preferentially or as molecular complexes 3) to establish whether there are any differences in the kinetics of adsorption. Metallic copper and sphalerite powder were chosen as adsorbents since they represent two extremes since copper readily adsorbs multilayers of xanthate (and is easily floated with ethyl xanthate) while unactivated sphalerite does not appreciably adsorb xanthate but is known to be readily floated by the amine type collectors. Metallic nickel was included because the infrared spectroscopic work2 on xanthate

Jan 1, 1964

By A. Sutulov

The segregation process, an alternative to benefi-ciation of oxidized and mixed sulfide-oxide copper ores, is studied here with particular emphasis placed on time, temperature, and coke and salt percentage factors. The results show that in a segregated charge, the copper assay of the coarsest fractions normally is of the same order as the copper assay of the tails obtained from grinding and flotation of the same charge. This observation suggests the possibility of limiting or even eliminating the grinding step from the flowsheet of a segregation plant. The segregation process, which offers an alternative to beneficiation of oxidized and mixed sulfide-oxide copper ores, was discovered in 1923 by Moulden and Taplin1,2 who were attempting to apply the Perkins process3&apos;4 to a Chilean ore from the Sagasca mine. After a number of investigations and pilot plant tests,&apos;-&apos; the segregation process became a successful industrial operation.&apos;r9 The process aroused considerable interest in Chile because of the existance of large reserves of oxidized copper minerals which could not be treated by conventional hydrometallurgical methods because of the high consumption and prohibitive cost of sulfuric acid. During the last two years, the Mineral Dressing Dept. of the Institute de Investigaciones Tecnology cas of University of concepci6n (Chile) has carried out a test program on 23 different ore samples representative of the national territory. These tests were carried out on a semi-continuous and batch scale with a rotary, automatically controlled and gas-fired furnace (supplied by Bartlett and Snow, Fig. 1). The results of these tests indicate that out of the 23 samples 20 can be successfully treated by this process with recoveries of copper ranging from 85 to 95%.10,11 The purpose of this paper is to treat some specific problems related to the segregation process. TIME AND TEMPERATURE FACTORS Time and temperature of calcination play an important role in the segregation process. When moderate salt and coke conditions were maintained (each of reagents was added in the quantity of 1 — % wt), systematic testing easily demonstrated the influence of these two factors on segregation and flotation recovery. The pertinent results are summarized in Fig. 2. To determine what specific changes the segregation process made on the ore, the samples, segregated at 700°C for different periods of time and under standard conditions, were granulometrically analysed and

Jan 1, 1963

By R. E. Powers, E. H. Kinelski, H. A. Morrissey

A group of 17 American blast-furnace sinters, an American open-hearth sinter, an American iron ore, and a Swedish sinter were used to evaluate testing methods adapted to appraise sinter properties. Statistical calculations were performed on the data to determine correlation coefficients for several sets of sinter properties. Properties of strength and dusting were related to total porosity, slag ratio, and total slag. Reducibility was related to the degree of oxidation of the sinters. THIS report to the American iron and steel industry marks the completion of a 1949 survey of blast-furnace sinter practice sponsored by the Subcommittee on Agglomeration of Fines of the American Iron & Steel Institute. The use of sinter in blast furnaces, sinter properties, raw materials, and sinter plant operation have been reported recently.1,2 After preliminary research and study," test procedures were adapted to appraise the physical and chemical properties of sinter to determine what constitutes a good sinter. During the 1949 to 1950 plant survey each plant submitted a 400-lb grab sample to research personnel at Mellon Institute, Pittsburgh, Pa. A 400-lb sample was also submitted from Sweden. In addition, 2 tons of group 3 fines iron ore were obtained from a Pittsburgh steel plant. The following tests were performed on the iron ore sample and on the 19 sinter samples: chemical analysis; impact test for strength and dusting; reducibility test; surface area measurements, B.E.T. nitrogen adsorption method; S.K. porosity test; Davis tube magnetic analysis; X-ray diffraction analysis for magnetite and hematite; and microstructure. Results of these evaluations are discussed in this paper and supply a critical look at testing procedures used to determine sinter quality. Sinter Tests and Results Each 400-lb grab sample of sinter was secured at a time when it was believed to represent normal production practice at each plant. It was not possible to use the same sampling procedures throughout the survey; consequently samples were taken from blast-furnace bins, cooling tables, and railroad cars. These were very useful for evaluation of test methods, since they were obtained from plants with widely divergent operations. With the exception of Swedish sinter and sinter sample N, which were produced on the Greenawalt type of pans, all survey sinters were produced on the Dwight-Lloyd type of sintering machines. Sinters submitted for test were prepared in identical manner by crushing in a roll crusher (set at 1 in.), mixing, and quartering. To secure specific size fractions for tests, one quarter of the sample was crushed in a jaw crusher and hammer mill to obtain a —10 mesh size. The remainder was screened to obtain specific size fractions. The group 3 fines iron ore was dried and screened and samples were taken from selected screen sizes to be used for various tests. Prior to testing, each ore sample except the —100 mesh fraction was washed with water to remove all fine material and was then dried. This iron ore, a hematitic ore from the Lake Superior region, was used as a base line for comparing results of tests on sinters. The iron ore did not lend itself to impact testing, since it was compacted rather than crushed in the test, and no impact tests are reported. However, the iron ore was subjected to all remaining physical tests to be described. Chemical Analysis: Table I presents chemical analyses performed on the survey sinter samples. Included in this table are data obtained from determination of FeO and the slag relationships: CaO + MgO and total slag (CaO + MgO + SiO, SiO2 + Al2o3 + TiO2). The percentage of FeO was used as an indication of the percentage of magnetite in the sinter. It was believed that slag relationships could be correlated with sinter properties. During initial determination of FeO great disagreement arose among various laboratories, both as to the results and the methods of determining values. Table I lists the values of FeO resulting from the U. S. Steel Corp. method of chemical analysis,&apos; which reports the total FeO soluble in hydrochloric and hydrofluoric acids (metallic iron not removed) with dry ice used to produce the protective atmosphere during digestion. Use of dry ice was a modification required to obtain reproducible results. In this method, the iron silicates and metallic iron are believed to go into solution and are therefore reported as FeO. This is important, for in the study of the microstructure of sinters, glassy constituents suspected of containing FeO as well as crystallized phases of undetermined identity which may also contain FeO have been observed. Strength Test by Impact: In evaluating sinter quality, one of the properties stressed most by blastfurnace operators is strength. This strength may be described as the resistance to breakage during handling of sinter between the sinter plant and the blast-furnace bins. It is also the strength necessary to withstand the burden in the blast-furnace. After

Jan 1, 1955

By F. Petkovich, M. P. Sudbury

Development of a new method of treating flotation concentrates for the preparation of feed suitable for direct charging to blast furnaces or converters is described. The method takes advantage of the fact that pyrrhotite contained in the concentrates can be made to undergo exothermic reactions which, under properly controlled conditions, result information of oxidation products that serve to cement the concentrate particles together. Thus green concentrates either in the form of pellets or briquettes containing 5 to 8 pct moisture not only were hardened, but also completely dried autogenously by continuous countercu,vrent shaft treatment with air. The process was developed through pilot plant operations ranging from 35 1b per hr to 2.5 tph, culminating in products evaluation tests on a railway-car scale. Falconbridge Nickel Mines Ltd., Falconbridge, Ont., have practised sintering, on a travelling grate machine, for many years to prepare a pyr-rhotite-bearing Cu-Ni sulfide flotation concentrate for blast furnace smelting. In recent years the development of new mines, and the construction of concentration plants some 40 miles from the smelter has increased the quantity of concentrate shipped, and encouraged the search for a method of producing dry, hard pellets or briquettes suitable for direct charging to a blast furnace or converter, and capable of withstanding considerable handling. The cementing properties of pyrrhotite oxidation products were discovered during a search for a suitable agglomeration method for sulfide concentrates, and were exploited in a process subsequently developed for autogenously drying and hardening pellets and briquettes of pyrrhotite-bearing Cu-Ni sulfide flotation concentrate. EARLY TEST WORK The production of hard, dry pellets of concentrate filter cake was attempted in the drying plant of one of the new concentrators. While soft pellets were produced containing 5 pct moisture, attempts at further drying resulted in disintegration of the pellets and intolerable dusting. The plant therefore was controlled to produce a concentrate containing about 6 pct moisture. The partially dried concentrate, which was shipped to Falconbridge for sintering, tended to heat spontaneously in transit. An at- tempt to briquette the moist concentrate at Falconbridge, in a plant incorporating a muller, a 20-in. roll briquetting press, and a steam-heated travelling grate oven, was more successful, although the equipment corroded rapidly due to the acid nature of the concentrate after spontaneous heating had occurred. LABORATORY STUDIES OF SPONTANEOUS HEATING OF CONCENTRATE The spontaneous heating of partially dried concentrate was studied in the laboratory. Moist concen-

Jan 1, 1961

By Z. F. Oszter

This paper discusses the materials handling problems encountered in cement plants in the areas of raw and finished product grinding and cement storage. It illustrates by examples taken from actual installations, how optimum feed rate is maintained in grinding circuits with properly applied bin dischargers (vibrating hoppers, Mexican feed boxes and aerated bin bottoms) for materials with high moisture content and/or difficult handling characteristics. Initial operating problems with the equipment will be discussed and maintenance requirements will be indicated for comparison purposes. In a modern cement plant, properly functioning bins and feeders are as essential as process control, for no process control system, however sophisticated, can perform efficiently if raw materials do not flow into the plant processing facilities at the required time and rate. Most materials used in the manufacture of portland cement - such as silica sand, iron ore fines, crushed limestone, gypsum, clinker or cement itself-have difficult handling characteristics under specific conditions. High moisture content or even dampness due to condensation may create bin hang-ups, arching and ratholing, all of which result in improper functioning of the feeders or proportioning equipment. The occurrence of these problems in existing cement plants spurred a search for the corrective devices discussed in this article. The equipment investigated included bin dischargers in the form of vibrating hoppers for moist raw materials, Mexican feed boxes for clinker and aerated bin bottoms for cement. Operating experience with different types of bin dischargers showed that the actual performance of some did not come up to the expected level under prevailing operating conditions. The aim of this paper is to record the results of this experience and there- by aid designers and/or operators in selecting the most suitable equipment for their requirements. BIN DISCHARGERS (GENERAL) A discharger is a device that causes stored solids to flow out of a hopper without metering the rate of flow as a feeder does. The investigation discussed here encompasses the following types of bin dischargers used in a cement plant: (1) vibrating hoppers (horizontal motion), (2) whirlpool hoppers (horizontal and vertical motion), (3) aerated bin bottoms (air fluidization), and (4) Mexican feed boxes. Most of the bins used in the cement industry are of the plug flow type. Raw materials with difficult handling characteristics such as high moisture content may cause flow problems when fed through small openings (12 x 18 in.), even if two walls of the bin are vertical.&apos; The most appropriate solution to these flow problems would be the installation of separate mass flow bins for the different materials such as silica sand, iron ore fines, crushed limestone and gypsum, with each bin correctly designed from test data obtained by measuring the flow properties of representative samples. However, limited storage capacity and plant space usually preclude this approach. Quantitative design of bins for flow based on measured flow properties of representative samples of the bulk solid to be handled is no longer an "engineering art," thanks to the basic research done on the flow of bulk solids in bins during the last decade.2, Where space limitations exist, however, or where improvements are being sought for equipment already installed, bin dischargers offer a satisfactory solution to flow problems. They can be applied effectively under flat-bottom bins or mounted to existing plug-flow bins with very little modification to these facilities. In both the foregoing cases, plug flow will be maintained, but the flow channel will be considerably increased so that it is larger than the discharger&apos;s bin-side diameter, and material flow can be sustained if the discharger is correctly selected. The main purpose of a bin discharger is to draw the bulk solids across the whole area of the bin outlet, to make the opening fully effective. As long as the bin-side opening of the discharger is large enough

Jan 1, 1969

By James J. Bean

HE task of measuring the specific gravity of the -*- operating medium in a heavy-media separation system has never presented a particularly difficult problem because the medium is fairly stable and the overflow of the separatory vessel, as well as its underflow, can be sampled easily and accurately and the specific gravity of the suspension determined easily by weighing a known volume. However, while this method is simple and accurate it does require the operator to take the sample by hand and to weigh it and there is considerable temptation to avoid the periodic sampling if everything seems to be going well, or if something is occupying the attention of the operator. Furthermore all operators do not sample in exactly the same manner and considerable practice is required for two operators to be able to "check" each other to the last few hundredths, particularly if the sample is cut underneath the drainage screen where location of the point of sampling and load on the screens tends to influence the determination. While none of the above presents much of a problem, we have all recognized that some mechanical method of continuous measurement and recording would be advantageous since the operator would merely have to glance at the meter to check the gravity and to have an indication of the trend of any changes. Also if the instrument were of the recording type, a permanent record would be available for the guidance of the superintendent. The Eagle-Picher Mining and Smelting Co. was the first heavy-media user to actually install such a recording meter. In 1946 they installed in their Central Mill at Cardin, Okla., a specific gravity recorder manufactured by the Bristol Co. of Water-bury, Conn. R. A. Barnes, of the Bristol Co., working with E. H. Crabtree, Jr. and Elmer Isern, of Eagle-Picher, made the application and worked out the problems of sampling and measuring. Their attempts to measure the specific gravity of the medium in the cone itself were not entirely successful and they resorted to an outside sample tube for actually making the determination. Because of the particular flowsheet used, it was possible to tap off from the medium return pipeline a stream of medium and divert it into the sampling tube, which was provided with a constant level overflow and a spigot underflow, and into which the bubbler tubes dipped. The Eagle-Picher installation was successful and its possibilities were recognized by the Mineral Dressing Laboratory of the American Cyanamid Co. It was decided to install a similar unit in the heavy-media pilot plant to investigate further its possibilities. Chief among these was the continuous record which it was felt would be proof of the steadiness of the gravity in a heavy-media cone, something which is not always appreciated by POtential users. Because the heavy-media pilot plant is required to operate at a wide range of specific gravities, it was realized that the unit would have to record all gravities from 1.25 to 3.50, and do it to the nearest 0.01. It would not be necessary to record all of this wide range on a single chart and the method selected was to have 4 bands, each band range overlapping the other a small amount and calibrated so that with standard charts one division would represent 0.01 sp gr. A shift from one band to another could be arranged without alteration of the instrument itself, being accomplished by a simple change in the bubble-tube lengths, as described later. Accordingly, a recording type instrument was purchased and installed. Because there were some advantages in doing so, the first installation attempted to measure the gravity of the cone proper by placing the bubble tubes in the cone. This was not at all satisfactory and the second scheme utilized a fixed vertical screen at the surface of the cone, and an external sample-tube arrangement. We were particularly anxious to make this work as we felt it would be advantageous to measure the top level of medium where the separation was actually being made, but we were doomed to disappointment because it was impossible to keep the screen clean of float. Since the top gravity of the cone is the most convenient place to sample for control, a launder about 2 in. wide was installed longitudinally beneath the

Jan 1, 1951

By James J. Bean

HE task of measuring the specific gravity of the -*- operating medium in a heavy-media separation system has never presented a particularly difficult problem because the medium is fairly stable and the overflow of the separatory vessel, as well as its underflow, can be sampled easily and accurately and the specific gravity of the suspension determined easily by weighing a known volume. However, while this method is simple and accurate it does require the operator to take the sample by hand and to weigh it and there is considerable temptation to avoid the periodic sampling if everything seems to be going well, or if something is occupying the attention of the operator. Furthermore all operators do not sample in exactly the same manner and considerable practice is required for two operators to be able to "check" each other to the last few hundredths, particularly if the sample is cut underneath the drainage screen where location of the point of sampling and load on the screens tends to influence the determination. While none of the above presents much of a problem, we have all recognized that some mechanical method of continuous measurement and recording would be advantageous since the operator would merely have to glance at the meter to check the gravity and to have an indication of the trend of any changes. Also if the instrument were of the recording type, a permanent record would be available for the guidance of the superintendent. The Eagle-Picher Mining and Smelting Co. was the first heavy-media user to actually install such a recording meter. In 1946 they installed in their Central Mill at Cardin, Okla., a specific gravity recorder manufactured by the Bristol Co. of Water-bury, Conn. R. A. Barnes, of the Bristol Co., working with E. H. Crabtree, Jr. and Elmer Isern, of Eagle-Picher, made the application and worked out the problems of sampling and measuring. Their attempts to measure the specific gravity of the medium in the cone itself were not entirely successful and they resorted to an outside sample tube for actually making the determination. Because of the particular flowsheet used, it was possible to tap off from the medium return pipeline a stream of medium and divert it into the sampling tube, which was provided with a constant level overflow and a spigot underflow, and into which the bubbler tubes dipped. The Eagle-Picher installation was successful and its possibilities were recognized by the Mineral Dressing Laboratory of the American Cyanamid Co. It was decided to install a similar unit in the heavy-media pilot plant to investigate further its possibilities. Chief among these was the continuous record which it was felt would be proof of the steadiness of the gravity in a heavy-media cone, something which is not always appreciated by POtential users. Because the heavy-media pilot plant is required to operate at a wide range of specific gravities, it was realized that the unit would have to record all gravities from 1.25 to 3.50, and do it to the nearest 0.01. It would not be necessary to record all of this wide range on a single chart and the method selected was to have 4 bands, each band range overlapping the other a small amount and calibrated so that with standard charts one division would represent 0.01 sp gr. A shift from one band to another could be arranged without alteration of the instrument itself, being accomplished by a simple change in the bubble-tube lengths, as described later. Accordingly, a recording type instrument was purchased and installed. Because there were some advantages in doing so, the first installation attempted to measure the gravity of the cone proper by placing the bubble tubes in the cone. This was not at all satisfactory and the second scheme utilized a fixed vertical screen at the surface of the cone, and an external sample-tube arrangement. We were particularly anxious to make this work as we felt it would be advantageous to measure the top level of medium where the separation was actually being made, but we were doomed to disappointment because it was impossible to keep the screen clean of float. Since the top gravity of the cone is the most convenient place to sample for control, a launder about 2 in. wide was installed longitudinally beneath the

Jan 1, 1951

By James J. Bean

F. M. Lewis—I believe that density recorders are a prerequisite to all well operated ore concentrators. Well designed density instruments are very accurate and give an excellent record of the steadiness of the operation. Controlling the density of a pulp with an instrument, as described by Mr. Bean, is new and the experiences in developing this controller at the American Cyanimid Co.&apos;s Mineral Dressing Laboratory are very interesting and should be extremely helpful to others who are planning on designing or installing controlling density instruments. Instruments for recording the density of the classifier overflow or flotation feed are rather common in ore concentrators, but there are a number of profitable uses for these instruments that are not too well known. Density recorders with bubble tubes submerged to different depth in a classifier pool give an excellent record of the classifier operation. The loading and unloading of the machine, that is so easily masked in the density of the overflow, can be easily recorded if the instrument is installed at some depth in the classifier pool. On some occasions, two instruments with bubble tubes at different depths give a more candid picture of the irregularities in the classifier operation. When cleaner machines are operated in closed circuit with roughers, a record of the density of the cleaner tails will indicate the circulating load in the circuit and is very helpful in coordinating different operators, because quite often the circulating load will be increasing for some time before the analyses of the products will change. Density recorders are used to record the variations in the dilution of the underflow from thickening tanks and in at least one plant the instrument controls the underflow.

Jan 1, 1951

By James J. Bean

F. M. Lewis—I believe that density recorders are a prerequisite to all well operated ore concentrators. Well designed density instruments are very accurate and give an excellent record of the steadiness of the operation. Controlling the density of a pulp with an instrument, as described by Mr. Bean, is new and the experiences in developing this controller at the American Cyanimid Co.&apos;s Mineral Dressing Laboratory are very interesting and should be extremely helpful to others who are planning on designing or installing controlling density instruments. Instruments for recording the density of the classifier overflow or flotation feed are rather common in ore concentrators, but there are a number of profitable uses for these instruments that are not too well known. Density recorders with bubble tubes submerged to different depth in a classifier pool give an excellent record of the classifier operation. The loading and unloading of the machine, that is so easily masked in the density of the overflow, can be easily recorded if the instrument is installed at some depth in the classifier pool. On some occasions, two instruments with bubble tubes at different depths give a more candid picture of the irregularities in the classifier operation. When cleaner machines are operated in closed circuit with roughers, a record of the density of the cleaner tails will indicate the circulating load in the circuit and is very helpful in coordinating different operators, because quite often the circulating load will be increasing for some time before the analyses of the products will change. Density recorders are used to record the variations in the dilution of the underflow from thickening tanks and in at least one plant the instrument controls the underflow.

Jan 1, 1951

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

Understanding the sedimentation of relatively thick suspensions depends on a knowledge of the concentration of the solids at various positions in the pulp and the times required for thickening. Flocculated kaolin slurries were studied with the transviewer, a device utilizing X-ray absorption in conjunction with a scanning mechanism. A mathematical model based on the Schuhmann size distribution in conjunction with Poiseuille&apos;s law yields experimental filtration rates as a hction of pulp density. These calculations show accord with the experimental results. Accurate predictions of sedimentation results depend upon determining the number and size of pores for given conditions, since the voidage of the suspension controls the volume of liquor that can pass through an element of suspension per unit time. A study on the sedimentation of flocculated kaolin slurries has indicated that their sedimentation takes place by two different processes, dependent on the initial concentration of solids in suspension. One process occurs in dilute suspensions (less than about 2 pct solids by volume) in which a multitude of relatively small flocs settle by hindered settling. In this case, maximal settling velocity is independent of container geometry.1 r2 The other process occurs in relatively concentrated suspensions (greater than about 2 pct solids by volume) in which the mass is considered to be settling as an aggregate network or as one large floc, defined as to size by the dimensions of the container. In this case, container geometry affects sedimentation results significantly. Furthermore, it seems that this single floc is in a state of compression from the beginning of settling, and that two phases of compression settling exist during sedimentation. The first occurs during the early stage of settling where liquid exudes easily and rapidly from the floc, while the second occurs in the compacting portion of the floc where water escapes more slowly and with much more difficulty. Sedimentation of relatively thick suspensions may then be thought to be a filtration phenomenon in which a pulp thickens by the filtration of its contained liquor through the aggregate network of pulp above it. This postulate has also been advanced by others.,4 To understand the sedimentation of relatively thick suspensions more fully, it was felt that knowledge of the concentration of solids at all positions in the

Jan 1, 1962

By D. W. Frommer, M. M. Fine

Mineral dressing research showed that iron concentrates of commercial quality could be produced from the Pea Ridge deposit near Sullivan, Mo. Magnetic separation and flotation, on a laboratory scale, yielded concentrates ranging from 60.0 to 71.4 pet Fe at recoveries of 89.9 to 97.6 pct. Early in 1957 St. Joseph Lead Co. announced discovery of three new centers of iron ore deposition in east central Missouri.&apos; The discovery resulted from exploratory drilling in the vicinity of a magnetic anomaly or high. A deep hole, drilled into the Pre-Cambrian porphyry to determine the cause of anomaly, penetrated a magnetite-rich deposit.&apos; The magnetic surveys that revealed the anomalies were begun by the Missouri Geological Survey in 1929 and were continued into the 1930&apos;s.3 At that time an area in Crawford County known as the Bourbon magnetic high was located. A second anomaly, similar in size and intensity, was discovered about six miles to the northeast in Franklin County in the vicinity of Sullivan. This area was not studied in detail until the early 1940&apos;s. To determine the cause of the anomalies, the USBM drilled four holes at the Bourbon site in 1943-1944.4 The most productive hole penetrated four mineralized zones having a total thickness of 127.5 ft. These zones contained magnetite in rhyo-lite porphyry at depths between 1600 and 2000 ft. Additional iron ore below 2000 ft was considered a possibility at the time. Some years later, the USGS, Missouri Geological Survey, St. Joseph Lead Co., and others cooperated in sponsoring an aeromagnetic survey of the Bourbon-Sullivan area which resulted in discovery of a third high, about eight miles distant at Pea Ridge. This deposit, with an estimated reserve of 50 to 100 million tons of iron ore, is considered the main find, although other orebodies in the same general area are of definite interest.&apos; When this discovery was announced, St. Joseph Lead also revealed plans to exploit the deposit jointly with Bethlehem Steel Co. A new corporation, Meramec Mining Co., was established, and 1962 was set as the target date for completing physical plant.2 Annual production goals of 2 million tons of beneficiated and pelletized iron ore were proposed. Preparation of the site for shaft sinking was begun in June 1957. Orebody and Character of Samples: Drilling to date has established the Pea Ridge orebody as a crescent-shaped pipe about 300 ft wide. Some drillholes were bottomed at 2800 ft in very good iron ore.6 he orebody is predominantly magnetite, although hematite and a mixture of hematite and magnetite occur in the upper areas and peripheral walls. Drillhole cores representing ore from different areas and depths of mineralization were grouped into composites for testing. These composites consisted of: 1) medium grade mixed hematite and magnetite, 2) high grade mixed hematite and magnetite, and 3) high grade magnetite. The medium grade ore came from three holes with depths of i350 to 1900 ft. The high grade mixed hematite and magnetite cores came from the 1790 to 2812-ft interval of one hole. Partial chemical analyses of the three samples, given in Table 1, reveal that all are excessively high in phosphorus and sulfur, and sample 1 is

Jan 1, 1960

By H. Kahata, I. Iwasaki, Y. Takahasi

A combined process of segregation roasting-flota-tion or magnetic separation was developed to recover nickel from lateritic iron ores and oxidized nickel ores. An ore was mixed with a halide, such as sodium or calcium chloride, and a solid reductant, and roasted, whereby the nickel content was segregated out as discrete metallic particles. Either by magnetic separation, or by flotation using a sulfhydryl collector together with a heavy metal salt as an activator, the metallic nickel could be concentrated, Effects of such variables as the types and amounts of the halide and solid reductant addition and of the roasting conditions were extensively investigated. The concentration results on several types of ores from Japan, the Philippines, and New Caledonia are presented. Laterite is a residual product of rock decay in tropical and subtropical regions, and consists mainly of hydrated oxides of iron and alumina. Those derived from serpentine sometimes become enriched in nickel, cobalt, and chromium contents. Oxidized nickel ores are of the similar origin, but their nickel as well as silica contents are higher due to gamierite-type silicates. Apparently, nickel enrichment is more pronounced in the transition region between the surface layer of iron laterite and the underlying serpentine. Iron laterites often contain over 50% iron, and their reserve is reported to exceed 22 billion tons, corresponding approximately to 20% of the world iron ore reserve.1 It is readily apparent, therefore, why the iron laterites have attracted much attention of the steel industry as a potential source of iron ores. Numerous attempts to remove nickel and chromium as byproducts, have been reported in literature.&apos; Worldwide discoveries of high-grade iron ore deposits in recent years have inevitably lessened the interests in iron laterites. However, in view of the tremendous potential for their gross value and supplies of nickel, cobalt, chromium, and alumina along with iron, the problem remains as a challenge. Much of the chromium in laterites is present in the form of relatively coarse particles of chromite, and various classification methods have successfully been applied for its removal.2,3 The nickel, however, is intimately associated with limonite, and ordinary concentration techniques are totally ineffective. Much attention has, therefore, been given to hydro-metallurgical or pyrometallurgical means. Yet, all the known processes appear to have limitations either economically or for the use of products as iron raw materials due to a high residual content of nickel. Some attempts have been made to pre-roast laterites or oxidized nickel ores, particularly with sulfidizing agents, and to remove nickel either by magnetic separation,&apos; or by flotation.4 It is apparent that any such concentration process would require grain growth of nickel during roasting for effective separation. Oxidized copper ores are reported to be concentrated by the segregation roasting-flotation process,5 in which the ores are roasted with a chloridizing agent and a carbonaceous reductant in order to promote the growth of metallic copper through repeated cycles of copper chloride volatilization and its reduction on the solid reductant surfaces. Examination of a vapor pressure-temperature diagram of various metal chlorides indicates that a separation of nickel from iron may be effected by a similar process. In fact, the segregation roasting of laterites was reported to increase the extraction of nickel by the ammonia leach process, as compared to the ordinary reduction roasting.6 In the present article some of the preliminary experimental results of the segregation roasting, followed by either flotation or magnetic separation, on laterites and oxidized nickel ores are reported. The effects of various parameters in the roasting as well as in the concentration steps are discussed. MATERIALS AND METHODS An oxidized nickel ore from Thio, New Caledonia, and three laterites from Homonhon, the Philippines, Wakasa and Miyagawa, Japan, were selected for the

Jan 1, 1967