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By J. J. Gilvarry, B. H. Bergstorm

The first part of this paper describes a new approach to the problem of energy relationships in fracture and comminution. The basic theoretical method used (as contrasted to previous empirical or semi-empirical approaches) is an attempt to verify or disprove the hypothesis of von Rittinger, In the second part of the paper, theoretical conclusions for the three headings of single fracture, plural fracture, and comminution are tested. Agreement between theory and experimental results is demonstrated. The classical hypotheses on the energy relationship in brittle fracture are those of von Rittinger-that the energy required is proportional to the new surface formed— and the alternative one of Kick.' Over the years, repeated attempts have been made to discriminate experimentally between the two assumptions, or to replace both of them, as has been done by Bond? However, no semblance of agreement seems to exist among different investigators. This paper is the first in a series essaying a new approach to the entire problem of energy relationships in fracture and comminution. The purpose of this program of research is to verify (or disprove) the hypothesis of von Rittinger. In contrast to previous empirical or semi-empirical approaches to the problem, the basic method adopted is theoretical. The prime aim is to determine the distribution function for fragment size theoretically and to verify the predicted distribution experimentally. This approach should make possible determination of the total surface of the particles by integration over the corresponding distribution function; one may hope that this direct procedure will yield more definitive results than use of gas adsorption or other methods. The surface area so obtained can then be compared with experimental energies, as obtained by Bergstrom, Sollenberger, and Mitchell,4 for example. It is convenient to distinguish between single fratture, plural fracture, and comminution. Single fracture of a brittle solid is defined as fracture by an external stress system which is removed instantly and permanently when fracture is initiated. Plural fracture consists of single fracture of the original specimen, followed by a sequence of only a few secondary fractures; comminution differs in that the sequence of secondary fractures consists of a large number of repetitive steps. For single fracture, Gilvarrys has given a rigorous derivation of the proper distribution function for fragment size, based on a closely defined physical model and deduced strictly by the laws of probability. Gilvarry and Bergstrome have compared the predictions of the theory with experiment, and have found excellent agreement. The prior work of Gilvarry and of Gilvarry and Bergstrom considers only single and plural fracture. The purpose of the present paper is to extend the discussion to the case of comminution. For prefatory purposes, the theory of Gilvarry for single fracture will be outlined. The considerations will then be applied to the cases of plural fracture and comminution. SINGLE FRACTURE The derivation is based on the Poisson law.7 For Points distributed with mean density y over a domain, this law states that the probability p(t)dt of one point lying in the range t to t + dt with none in the interval 0 to Ms p(t)dt = ye-yt dt. For this distribution to be valid, it is necessary and sufficient that the points be distributed at random, individually and collectively in the sense of ~~~.7 The former qualification requires that the position of One Point be independent of another, and the latter implies that the probability of a region containing a

Jan 1, 1961

By B. H. Bergstrom, J. J. Gilvarry

Previously the authors showed that the Gilvarry equation correctly describes the distribution of fragment size in single fracture, provided the exoclastic particles showing original surface of the specimen are removed from the distribution. In the present work this conclusion has been strengthened in two respects. First, it has been shown that the existence of two peaks in the differential distribution for the endoclastic particles, corresponding to the effect of surface and volume flaws in the interior of the specimen, can be demonstrated from the average of data for a large number of specimens. Second, it has been shown by means of a Coulter counter that data obtained from specimens fractured in gelatin display the behavior at fragment sizes approaching 1 p which is demanded by the limiting form of the Gilvarry equation in this case. In previous work, a rigorous derivation of the proper distribution function for fragment size in single fracture of a brittle solid has been given by Gilvarry,' based on a closely defined physical model and deduced strictly by the laws of probability. Experimental work has been reported by Gilvarry and Berg-str~m,~ confirming the theory in its broad outlines. Subsequently, Gilvarry and Bergstrom extended the theoretical discussion to include comminution and presented experimental data bearing on this point. They have described further experimental results for the case of single fracture recently. The experimental evidence regarding single fracture was rather convincing. However, it was based on only five specimens for which fragment sizes were determined by sieve analyses and three for which measurements in the finer sizes were made on a Coulter counter. The purpose of the present paper is to extend the experimental results as obtained by means of both sieve analyses and the Coulter technique. In the former case, sufficient specimens are included so that the average data should vield ac- curately the detailed behavior of the distribution function for the coarser fragment sizes. In the latter case, measurements for the finer sizes are reported which correspond to single fracture more closely than prior results and fill a lacuna left in the experimental verification. In addition, an extension of the theory given recently bGilvarry for the case of fracture of a two-dimensional solid will be outlined, and experimental results obtained by him will be discussed.6 THREE-DIMENSIONAL CASE Gilvarry' has shown that the Schuhmann size modulus is a measure, in a sense specified later, of the number of activated edge flaws in a specimen undergoing single fracture. Single fracture is defined as that resulting from a single instantaneous application of energy, such as the blow of a hammer whose motion is halted when fracture begins. Plural fracture consists of a short sequence of single fractures, as in a jaw or gyratory crusher; multiple fracture obtains when many applications of energy are made, as in a ball mill. The Griffith flaws associated with fracture of a specimen are flaws occurring within the volume of the body, on the new surfaces created within its interior and on the new edges (intersections of two surfaces) within the body. Fracture proceeds by activation of flaws in the volume of the specimen, in the fracture surfaces throughout its interior, and in the edges produced by intersection of fracture surfaces. If all the flaws and all groups of flaws are distributed at random, independently of other flaws or groups thereof, it has been shown by Gilvarry that the distribution function for fragment size has the form of a Poisson distribution. The fraction y, by weight or volume, passing a sieve of mesh size x is given by where k is an average spacing of activated edge flaws, j2 is an average amount of surface containing one activated surface flaw and i3 is an average volume corresponding to an activated volume flaw. The averages in question are referred to the mean dimension x of a fragment. On the basis of the theoretical

Jan 1, 1962

By D. S. Cahn, D. W. Fuerstenau

Specimens of glass, plexiglass, and naturally occurring minerals were irradiated by a 15-joule ruby laser pulse to determine criteria for the fracture of nonmetallic solids by internal thermal stresses. Fracture did not occur in those specimens with either a very high or very low internal absorptance or with a high spectral reflectance. If significant amounts of melting and vaporization do not occur at the surface of irradiation, the specimen will fracture if a derived 11fracture stress" exceeds the ultimate tensile strength of the specimen. The purpose of this investigation was to determine criteria for the fracture of nonmetallic solids by internal thermal stresses which are caused by absorption of laser irradiation. A laser produces a very short pulse of light which has a high energy density distributed over a very narrow part of the visible spectrum. The laser system used in this investigation had an energy output of 15 0 joules at a wavelength of 6940 A. Fracture in a non-metallic solid, then, would be caused by the absorption of light energy by that solid, and subsequent conversion of this energy to heat or mechanical energy. Fracture is caused by internal stresses that result from rapidly imposed nonequilibrium thermal gradients. ANALYSIS OF CONDITIONS FOR FRACTURE BY LASER IRRADIATION For this analysis, the laser beam is assumed to be circular and centered on the junction of the major and minor axes of a rectangular specimen. The laser beam forms a cylinder of radius rb extending through the specimen width. The beam cylinder may be then thought of as a heated mass constrained by a surrounding unheated mass. The instantaneously heated cylinder will at-

Jan 1, 1968

By W. David Leyshon

This paper compares the general characteristics of Florida phosphate rock for the manufacture of fertilizers and phosphoric acid. The subjects covered are relative reactivity for superphosphate manufacture and various properties of the rock for use in phosphoric acid production, including gypsum filterability, sulfuric acid consumption, foaming, relative cor-rosiveness, chemical composition, and others. Comprison will include rocks of North Africa, the western U.S., and other locations. This paper reviews some of the important properties of phosphate rock, giving particular attention to a comparison of Florida phosphate with some of the major phosphates produced in the world. It also discusses some of the techniques available for evaluatine Phosphate. rock for phosphoric. acid and fertilizer use. The use of Florida rock outside the U.S. has been increasing steadily, evidence in itself of certain desirable properties of this rock. Table I is a list of the major producers and exporters of phosphate rock as of 1965. Florida, meaning in this case Central Florida, was the largest producer of phosphate rock and the largest exporter of rock. Of a total world production estimated at 60 million tons, about 50 million tons were produced during 1965 by the areas listed in Table I. One major producing area omitted from Table I and from subsequent evaluations in this paper comprises Nauru, Ocean, and Christmas Islands, whose output totaled about three million tons in 1965. This phosphate is marketed entirely by the British Phosphate Commissioners, primarily in Australia and New Zealand, and does not get into the world market elsewhere to a significant degree at the present time. South Africa, Egypt, Israel, and North Vietnam also produce lesser quantities of rock for the international market, and these materials are also excluded from this review. The phosphate industry is now seeing the introduction of several new areas that will be significant factors in the rock market in the future. The North Carolina and North Florida areas each have capacity at present or in the near future for producing about three million tons per year. The Djebel Onk mine in Algeria has a potential of about 800,000 tons per year. In addition, announcements have been made of significant development of deposits in Peru and in the Spanish Sahara, and sizable expansions are planned in several of the existing areas that are now of minor importance. Competition for the rock market appears to be increasing so that more attention will probably be paid to evaluating phosphates available from different areas. It is not possible here to go into all the properties of even a few of the materials listed in Table I. However, it is practical to discuss some comparative properties important for making phosphoric acid and fertilizers, and to touch on the techniques used for the evaluation of various phosphate rocks. Table II gives typical analyses of the major phosphate rocks available in the world as of 1965. There is, of course, an inherent danger in labeling any analysis as typical, but it appears that variations between two samples of rock within one given area are generally less than between two samples from different areas. The analyses listed were compiled from several years of data so they may not be representative of the present output of any particular area. The range of grades represented by these analyses goes from 66 BPL for Gafsa rock to 85 BPL for Kola

Jan 1, 1968

By R. W. Coughlin, W. U. Ault

Experiments in laboratory and mine have clearly demonstrated the usefulness of radioactivity detection for directly measuring the potassium content of potash ore as mined. This paper presents the results of these tests and documents the design of a system for automatically grading and sorting potash ore on the basis of radiation detection. It is shown how the experimental results in the subject mine lead directly to the design of this kind of an automatic control system. The system is capable of grading the ore with an accuracy of approximately ±2% K20 content or better. Due to the usual stratification of the ore body, mining potash seldom presents an extreme problem of gangue separation. Nevertheless, it is frequently necessary to transport within a potash mine and even hoist to the surface considerable quantities of halite and other valueless materials. Up to recent times it has been customary to recognize and route these materials by simple visual and manual techniques. However, the work reported here clearly demonstrates that it is now practical and economical to improve upon these techniques by continuous measurement1 of the naturally occurring radioactive isotope, potassium-40. It was the purpose of the work reported here to investigate this technique and assess its applicability. LOCALE OF EXPERIMENTS The work described here was carried out in a potash mine in Carlsbad, N.M. At the working face of this mine the potash ore is freed by blasting; it is then transported a short distance by shuttle car to a moving-belt system which conveys it to the main shaft where it is hoisted to the surface. Quantities of gangue which are also produced in the mining operation are disposed of in mined-out sections of the mine where possible. Otherwise, this material which contains only a small percentage of potassium (as K2O) must also be transported by the conveyor-belt system to the main shaft and then hoisted to the surface and discarded. With this situation, any measuring device for automatic control purposes must be capable of clearly distinguishing between gangue containing insignificant amounts of K2O and ore containing larger percentages of the same mineral. The conveyor-belt system transports the material from the mine face to a rotary crusher and then to temporary storage pits at the main shaft. This is shown in Fig. 1. Upon leaving the conveyor belt the material enters either one of the two storage pits according to the position of a manually operated gate. It was decided that all experiments should be carried out at the conveyor belt between the crusher and gate with a view to the possibility of providing a signal which could operate the gate automatically. THEORETICAL BASIS AND LABORATORY EXPERIMENTS In principle it is possible to grade potassium-bearing ore on the basis of radioactivity. Natural potassium contains 0.012% K40, which is radioactive with a half-life of 1.3 x 10' years. Although about 89% of the K40 disintegrates to Ca40 by beta decay, the other 11% changes to Ar40 by electron capture and the emission of gamma rays. Unlike the ß-rays which are readily absorbed in the first few millimeters of ore surrounding the disintegrating nucleus, the y-ray photons will escape from an ore sample several inches in size. A detector rigidly mounted under a conveyor belt loaded with potash ore will intercept some constant fraction of all the photons emitted from the ore within its field of view. Under these

Jan 1, 1968

By F. J. Lecznar

An investigation of the problem of separating two minerals by a moving inclined plane has revealed that only minerals with considerable difference in sliding friction coefficients separate readily; that it is possible to separate two minerals when only one of them has a rolling friction coefficient; that difference in geometrical shape of the minerals is important to separation which improves as grain size increases; and that the thickness of the stream falling on the belt influences the grade of separation - the thinner the stream, the smaller the interference between the grains of the minerals to be separated and the higher the degree of separation. The problem of separating two minerals with different friction coefficients and shapes by a moving inclined plane has been investigated. Friction separators can be used to separate two materials that differ in shape and in friction coefficient. 1,3 Gladkov has presented some theoretical aspects of using a moving plane to clean corn and podded seeds; this method, however, did not result in practical application. When considering the separation of minerals, only those with considerable difference in sliding friction coefficients can be taken into account. Furthermore, it is possible to separate two minerals when only one of them has a rolling friction coefficient. The difference in geometrical shape of two minerals is also important to separation. It has been found that the separation of two minerals improves as the grain size increases. For example, coal and slate over 2 in. in grain size separate much better than materials of smaller grain size. As Table I indicates, the slate grains smaller than 25 mm gradually lose their flat shape with decreases in size and are more rounded than larger grains, which are almost flat. The smallest grains of slate in first and second fractions slide down the belt since their friction coefficients decrease and approach the value of the rolling coefficient. The percentage of separation of the first fraction is, therefore, low. Thickness of the coal stream falling on the belt also influences the grade of separation; the thinner the stream, the smaller the interference between the grains of the two minerals to be separated and the higher the degree of separation.

Jan 1, 1963

By Jonas Svensson

A new method is described for the production of cold-bonded pellets using a hydraulic binder, such as portland cement. Large-scale pilot-plant tests have proved that self-fluxing pellets of high reducibility and good handling strength can be made by the method. Blast-furnace trials have shown that the pellets are an acceptable burden material, comparable with self-fluxing sinter or heat-hardened pellets. Economic factors of commercial-scale production are discussed. The Grangcold Pellet Process—for which patents have been applied or already granted in a number of coun-tries—uses a hydraulic adhesive such as portland cement, slag cements, pozzolanic cements, etc., for the production of cold-bonded pellets. The idea of using a hydraulic binder for the agglomeration of iron-ore fines is not new. Portland cement was proposed as an adhesive for cold-bonded iron-ore briquettes in patents granted more than 50 years ago.' In a report on the briquetting of iron-ore fines, published in Stahl und Eisen in 1959; it is stated that briquettes bonded with portland cement are used on a small scale at an ironwork in Germany. According to the report, the briquettes showed excellent strength in the blast furnace although their general use was made impossible because they required a long hardening time, during which they are sticky, soft, and difficult to store and handle. The Grangcold Pellet Process has overcome this particular disadvantage by mixing the balls with a suitable amount of the balling concentrate before storing them. The pellets are embedded in the concentrated during storing in such a way that they are isolated from each other and thus prevented from sticking together to form clusters. Thanks to the embedding concentrate, the pellets are subjected to a more or less uniform pressure from all sides which does not deform them. Thus, the mixture can be stored in a stockpile or in a bin until the pellets have hardened sufficiently. The concentrate is separated from the pellets by means of screening. The concentrate is returned to the balling operation and the pellets are either shipped to the blast furnace or stored for final hardening. The binder preferred for the Grangcold Pellett Process is portland-cement clinker, ground without the admixture of gypsum in order to avoid sulfur in the pellets as far as possible. Usually a 10% binder content is used. Two-thirds of the portland-cement clinker consist of lime and the rest is silica, alumina, and ferric oxide. Thus, self-fluxing or overbasic pellets are produced with this binder if the amount of silica in the concentrate used does not exceed 4%. The Grangcold Pellet Process was developed by the mineral Processing Laboratory of the Granges Co. Work started in 1963 with batch-scale tests. In 1966, a small pilot plant was put into operation in which 1800 tons of pellets were produced using 10% of rapid-hardening portland cement as a binder. Favorable results from a blast-furnace test with this batch led to the decision to erect a larger pilot plant which went into production in the summer of 1967. Since then, approximately 100,000 tons of cold-bonded pellets have been produced, mostly with 10% gypsum-free portland cement as a binder. Several full-scale blast-furnace trials have been performed with the pellets. The results of the trials indicate that the Grangcold pellets constitute a satisfactory blast-furnace feed. An industrial plant for the production of Grangcold pellets with a rated capacity of 1.5 million tpy is now under construction at the Granges Co.'s mine at Grangesberg. The plant will come into operation in the summer of 1970. Results from Laboratory Work Pellets made from iron-ore concentrate bonded with portland cement harden slowly and their handling is very critical until they have hardened enough to loose their stickiness. It is therefore especially important to study the progress of the hardening action and the factors influencing it. This is best achieved by investigating the relationship between the compressive strength of the cement-bonded pellets and the curing time under varied conditions. The general course of this relation-

Jan 1, 1971

By R. J. Woody

Design of the most economic continuous counter-current decantation (CCD) circuit is based on selection of the number of stages and the wash volume that will give the minimum summation of the following items: 1) Capital and operating costs of the CCD circuit. 2) Capital and operating costs of the precipitation circuit. 3) Value of the dissolved product lost.

Jan 1, 1959

By F. C. Bond

SIZE of grinding media is one of the principal factors affecting efficiency and capacity of tumbling-type grinding mills. It is best determined for any particular installation by lengthy plant tests with carefully kept records. However, a method of calculating the proper sizes, based on correct theoretical principles and tested by experience, can be very helpful, both for new installations and for guiding existing operations. As a general principle, the proper size of the make-up grinding balls added to an operating mill is the size that will just break the largest feed particles. If the balls are too large the number of breaking contacts will be reduced and grinding capacity will suffer. Moreover, the amount of extreme fines produced by each contact will be increased, and size distribution of the ground product may be adversely affected. If the balls added are too small, grinding efficiency is decreased by wasted contacts that are too weak to break the particles nipped; these largest particles are gradually worn down in the mill by the progressive breakage of corners and edges. Ball rationing is the regular addition of make-up balls of more than one size. The largest balls added are aimed at the largest and hardest particles. However, the contacts are governed entirely by chance, and the probability of inefficient contacts of large balls with small particles, and of small balls with large particles, is as great as the desired contact of large balls with large particles. Ball rationing should be considered an adjunct or secondary modification of the principle of selecting the make-up ball size to break the largest particle present. Empirical Equation In 19521,2 the author presented the following emerical equation for the make-up ball size: B - ball, rod, or pebble diameter in inches. F = size in microns 80 pct of new feed passes. Wi - work index at the feed size F. Cs - percentage of mill critical speed. S — specific gravity of material being ground. D == mill diameter in feet inside liners. K - 200 for balls, 300 for rods, 100 for silica pebbles. Eq. 1 was derived by selecting the factors that apparently should influence make-up ball size selection and by considering plant experience with each factor. Even though Eq. 1 is completely empirical, it has been generally successful in selecting the proper size of make-up balls for specific operations. But an equation based on theoretical considerations should be used with more confidence and have wider application. The theoretical influence of each of the governing factors listed under Eq. 1 was accordingly considered in detail, as described below, and a theoretical equation for make-up ball sizes was derived. Derivation of Theoretical Equation Ball Size and Feed Size: The basis of this analysis is that the largest ball in a mill should be just sufficient to break the largest feed particle into several pieces, excluding occasional pieces of tramp oversize. In this article the size F which 80 pct passes is considered the criterion of the effective maximum particle feed size. The smallest dimension of the largest particles present controls their breaking strength. This dimension is approximately equal to F. As a starting point for the analysis it is assumed that a 1-in. steel ball will effectively grind material with 80 pct passing 1 mm, or with F- 1000µ or about 16 mesh. The breaking force exerted by a ball varies with its weight, or as the cube of its diameter R. The force in pounds per square inch required to break a particle varies as its cross-sectional area, or as its diameter squared. It follows that when a 1-in. ball breaks a 1-mm particle, a 2-in. ball will break a 4-mm particle, and a 3-in. ball a 9-mm particle. This is in accordance with practical experience, as well as being theoretically correct. Confirmation of this reasoning is supplied by the Third Theory of Comminution," which states that the work necessary to break a particle of diameter F varies as F. Since work equals force times distance, and the distance of deformation before breaka4e varies as F it follolvs that the breaking force should vary as F½ These relationships are expressed in Table I, with a 1-in. ball representing one unit of force and breaking a 1-mm particle. This establishes theoretically the general rule used in Eq. 1 that the ball size should vary as the square root of the particle size to be broken. Ball Size and Work Index: The work input W required per ton" varies as the work index Wi, and the

Jan 1, 1959

By B. Bernstrom

A 22-ft diam, 7-ft long, wet autogenous grinding mill was installed in the new Cretaceous plant of the Jones & Laughlin Steel Corp. to prepare crude iron ore for concentration in spirals and flotation cells. The paper describes the flowsheet and gives a description of the operating and metallurgical results that have been obtained. These results include capacity, horsepower, operating problems and a comparison of pilot plant results with those obtained in commercial operation. The Cretaceous plant at the Hill Annex mine of the Jones & Laughlin Steel Corp. was designed with a wet autogenous mill for grinding low grade iron ore so that it could be concentrated by spirals and flotation cells. The plant, on the west end of the Minnesota's Mesabi Range, went into operation in June 1961, and preliminary operating and metallurgical information describing the performance of this mill can now be given. The crude is lean ore from the Cretaceous iron foundation. This formation is a beach deposit of Cretaceous age that overlies the Precambrian direct shipping and concentrating ores. This rock in nature is not extremely hard but is quite resilient, resisting blasting in mining and size reduction in milling. It contains hematite as water-rounded particles and other fragments of the underlying iron formation and adjacent rocks cemented together by a high alumina-silica ferruginous matrix. It is interspersed with hard dense hematite agglomerates and also contains granite pebbles and boulders that were not cleaned off during removal of the surface. It has a 3.48 sp gr and in the stockpile the material consists of frag-

Jan 1, 1962

By E. M. Furness, A. S. Henderson

ORIGINALLY the Babbitt experimental plant grinding circuit consisted of one rod mill 101/2 ft diam by 12 ft long in open circuit followed by two ball mills 101/2 ft diam* 12 ft long in parallel cir- the slow-speed ball mills. The next step was to improve rod mill operation by changing the speed of this mill. Such an attack would affect only one mill and would allow slower speed on the ball mills. The rod mill, as originally purchased, operated at 15.45 rpm. Additional pinions were purchased to test this mill at 16.45, 17.45, and 18.45 rpm. When the mill speed was increased by the use of these pinions, rod breakage with this short mill increased out of all proportion to the improved grinding, until at 18.45 rpm breakage became as high as 7 lb per

Jan 1, 1958

By C. G. McLachlan

ONE phase of the treatment in sulphide flotation, covered generally in a review of pulp pretreat-ment by S. A. Falconer,' is the matter of grinding mills as conditioners, a subject on which further study could be advantageous. Flotation with few exceptions is preceded by grinding, and the grinding and flotation capacity required for the treatment of an ore is based generally on laboratory or pilot mill tests. The results of such tests, it then is assumed, will be duplicated in practice, and a flowsheet is prepared on that basis. This assumption may be too broad, because in several instances laboratory flotation tests made on massive sulphide copper ores could not be duplicated in plant operation without the introduction of an aeration step, either in closed circuit with the grind-inging mills, or between grinding and flotation.~' " The differences between laboratory and plant grinding practice are considerable, and may be summarized as follows: 1—In the laboratory a relatively fine charge is ground in a batch mill. In the plant, the feed, usually considerably coarser than the ore tested in the laboratory, enters the grinding mill continuously and more often than not, oversize material in the mill discharge is retained in the grinding circuit by a classifier until the required degree of reduction has been effected. 2—The rate of grinding in the laboratory is slower than in plant practice. 3—The percentage of volume occupied by the pulp in a laboratory mill rarely approaches that of a continuous overflow type mill. This does not apply if the comparison is made with grate type mills, but in that case an even wider discrepancy between the relative rates of laboratory and mill grinding almost invariably exists. The present trend in grinding is toward larger mills to reduce the number of operating units. This accentuates the foregoing differences. It is believed that the trend toward larger mills is sound, but that when large mills are used, metallurgists should check their flotation circuits carefully to determine whether laboratory tests are being reproduced in

Jan 1, 1952

By F. M. Lewis, J. E. Goodman

A larger, slow-speed, under-loaded ball mill and hydraulic classifier have almost doubled grinding efficiency at the lsabella mill. TENNESSEE Copper Co. operates two ore con-A centrators, the London and Isabella mills near Copperhill, Tenn. In 1948 and 1949 the small ball mills and rake classifiers in the London concentrator were replaced by one large ball mill and one hydraulic classifier. This new ball mill was designed oversize so that it could be operated at slower than normal speed. The proper operating conditions were established and the results published.' Later it was found that this mill operated more efficiently with a small ball charge." While this London grinding practice was being developed, the Isabella grinding circuit was not changed in any way. It remained a conventional two-stage rod mill—ball mill combination, the rod mill in open circuit and the ball mill in closed circuit with a rake classifier. Study of the data indicated that grinding cost could be lowered by converting to the London practice, but not enough would have been saved to warrant the expense of converting. No plan for improving this grinding operation could be developed. In 1953 this study was reopened when plans were made to increase the Isabella mill output from 1150 to 1500 tpd. At this time the grinding circuit consisted of a 6x12 rod mill followed by a 6x12 ball mill in closed circuit with a 6-ft rake classifier. Another 6x12 ball mill and 6-ft rake classifier would have been sufficient equipment and would have been the simplest and cheapest installation. However, this would have offered no improvement in grinding efficiency or operating cost. If, on the other hand, the grinding operation could be converted to the London practice, by installation of one large 101/2x9x9 tricone mill and a 10-ft hydroscillator,2 lower operating costs would justify the additional investment. Economics: Table I presents estimates for the two schemes for enlarging the Isabella mill. In the first scheme, column 1, one 6x12 ball mill and one 6-ft rake classifier were to be added. In the second scheme, column 2, the existing small ball mill and rake classifier were to be retired and one large mill and hydroscillator were to be installed. Estimated savings from the difference in operating costs made an attractive return on the additional investment for the large mill and hydroscillator. Actual costs of carrying out this plan are tabulated in column 3b. It will be seen that these costs follow the estimates in column 3a very Closely. The return on the additional investment is calculated from the difference in the operating cost in column 1 and the actual operating cost in column 3b. This too come very close to the estimate. Power and Steel: Power requirements and steel consumption for the rod mills and ball mills, both actual and estimated, are presented in Table 11.

Jan 1, 1958

By O. J. Malacarne, D. G. Millenbruch

The use of a series of thickeners for washing of the slime fraction of a leached ore is not uncommon in the uranium industry; however, few ore processing mills have used a thickener circuit for washing leached total ore. The common practice is to wash the sand fraction in classifiers and the slime fraction only in thickeners. At the Uravan mill of Union Carbide Nuclear Co., a total ore acid leach is employed to extract uranium values from sandstone ores. These leached ores are then washed at a nominal rate of from 1000 to 1300 tpd in a series of washing thickeners. A typical screen analysis of the ore being washed in these thickeners is: 25 pct, -200 mesh; 3 pct, +35 mesh; and about 1 pct, +14 mesh. The amount of slimes (-200 mesh) will vary from 20 to 30 pct as the mill receives ore from 150 different mines. The washing liquor is acidic, from a pH of about 2 in the first thickener to 0.6 pH in the last thickener and, of course, all acid duty equipment must be used. The Uravan mill of Union Carbide Nuclear Co. is located in Montrose County, Colo. and is physically located on a canyon wall along the San Miguel River. The thickeners referred to in this paper are located on top of the mesa above the canyon. For purposes of identification, these thickeners are numbered 1 through 8 from south to north. GENERAL DESCRIPTION OF CIRCUIT The circuit utilized consists of eight wooden fir tanks, 60-ft diam by 12-ft deep, set in a line and staggered in elevation by approximately 3 ft. The washing water flows through the thickeners by gravity, while the ore pulp is pumped uphill from thickener to thickener. Adequate piping and pumping capacity are available so that any of the thickeners can be bypassed in the circuit. All the controls for the

Jan 1, 1962

By Olav Mellgren

Part I. The interaction between potassium ethyl xanthate and lead salts has been studied thermo-chemically. It is shown that ethyl xanthate reacts with lead carbonate, basic carbonate, thiosulfate and sulfate according to principles of simple chemistry, whereas reactions with lead sulfite, metasulfite, basic thiosulfate and hydroxide at high pH are associated with secondary reactions. A new compound "xanthogen thiosulfate" is postulated. Part 11. Adsorption of ethyl xanthate on galena has been studied quantitatively and thennochemi-cally. It is shown that an ion-exchange mechanism governs the adsorption process and that the heat of adsorption of xanthate on galena is equivalent to the heat of reaction of ethyl xanthate with bulk lead salts. It is concluded that the entropy content of the surface "compound" on galena is different from that of a corresponding bulk lead salt. Alkali depression is in part shown to be due to the instability of the xanthate-galena system at high pH and to the formation of basic lead salts. Sulfate and carbonate ions are shown to have a stabilizing effect on the xanthate-galena system. PARTI. THERMOCHEMISTRY OF ETHYL XANTHATE AND OF ITS REACTION WITH LEAD SALTS In the past little attention has been paid to the chemistry of lead xanthate. It is known that xan-thates form slightly soluble compounds with lead and that suspensions of lead xanthates in alkaline solution decompose rather readily.' The solubility of lead ethyl xanthate is of the order of 2.6 x 10-6 moles per 1. At increasing pH the xanthate concentration of a saturated solution contains more xanthate. Thus at pH 11 this xanthate content is about 100-fold larger than at pH 7.1 Xanthates are known to be consumed when added to a suspension of slightly soluble lead salts such as lead sulfate and lead carbonate. It is unknown to what extent this xanthate consumption is due to a simple chemical reaction because, frequently, observations are made which reveal secondary reaction products in such reactions. Thermochemical study of the reaction between xanthate and lead salts offers a means to clarify the significance of these reactions as the enthalpy changes can be accurately predicted from heat of formation of lead salts in pure solution reactions. In the present work the heats of formation of lead ethyl xanthate from potassium ethyl xanthate and lead salts are determined and some thermochemical aspects of xanthate in solution are studied. The results of these investigations are presented in Part I. In Part II the heat of adsorption of ethyl xanthate on galena surfaces is presented. All enthalpy changes were measured directly in a microcalorimeter. Calorimeter: The calorimeter used for measuring enthalpy changes was similar to that used by Redinha and Kitchener2 which was based on a description by Derbyshire and Marshall.3 Other useful references4,5 are listed. A sketch of the calorimeter

Jan 1, 1967

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

Inasmuch as the higher grade potash deposits are rapidly being depleted, it is important that economic beneficiation techniques be developed for recovering potassium compounds from lower grade sylvite ores and deposits containing clay materials. The Federal Bureau of Mines has conducted research to develop processes for heavy liquid concentration of potash ores, This method of concentration is discussed, and results of tests made on potash ores from New Mexico are presented. The tests demonstrated the technical feasibility of producing near commercial-grade potash concentrates, with a high recovery of the contained sylvite. The concentration of minerals based on their differences in specific gravity dates back to the very beginning of mineral dressing, and the use of heavy liquids in the laboratory to achieve optimum mineral separations has been utilized for many years. With the advent of froth flotation and its success, the incentive to develop new processes has remained relatively dormant. However, flotation refinements in the nonmetallic ores have not progressed as rapidly as in metallic (sulfide) ores because of the complexity of the deposits and the chemical similarities of the minerals.

Jan 1, 1967

By E. C. Tveter, R. B. Tippin

The separation of minerals by heavy liquids is a standard laboratory technique which goes back at least 50 years, but commercially economic application of this principal to ore concenfration has been prevented by problems of liquid recovery. Recent investigations which demonstrate the different types of liquid-solid systems on various ores using the two major heavy liquids - methylene bromide and tetra-bromoelhane — are discussed in this article. As a result of these studies, over-all liquid recovery, and especially desorption of heavy liquid from the mineral surface, was improved by the use of surfactants as well as both thermal and mechanical energy. The examples given in the article indicate that heavy liquids which should find major commercial use in ore concentration are now available. The sink-float process using a heavy liquid medium has long been recognized as a nearly perfect method of mineral separation based on the differences in density of the minerals. The use of heavy liquid separation (HLS) offers the potential of an ore dressing method that provides the maximum possible natural concentrate grade and high recovery. In addi- tion the process requires no water, causes no waste disposal problems, and is extremely simple — the crushed ore is merely introduced into a liquid whose density is between that of the minerals to be separated. The minerals lighter than the liquid float while those heavier than the liquid sink. As such, HLS has become a standard laboratory tool but major commercial application has been prevented by the requirement for complete recovery of the heavy liquid, imposed by economics and materials handling hazards. Investigations, both in the United States and abroad, have shown that improved liquid stability as well as improved chemical engineering techniques brighten the future for HLS in ore dressing. The available commercial heavy liquids suitable for mineral separation cost from $2.00 to $20.00 per gallon1 and on this basis alone their complete recovery is economically desirable. Possibly of even greater importance is the fact that any residual heavy liquid remaining on the products could create a serious toxicity hazard. Consequently, to make the process commercially attractive requires (1) essentially complete removal of the liquid from the ore products and (2) optimum recovery of the liquid for reuse. Until recently these two stipulations have remained as obstacles but laboratory and pilot plant research now in progress suggests that these may soon be overcome. Various heavy liquid recovery systems have been proposed to meet these requirements and might be classified into one or more of the following three general classes: Solvent Wash (Fig. 1): The dissolution of the heavy liquid from the ore with a less expensive solvent,

Jan 1, 1969

By W. B. Dancy, A. Adams

Laboratory test work on heavy liquid separation of sylvite from halite is reported. Numerous tests were run on sylvite ore sized in the ranges of 4x20 mesh, 10x65 mesh, 8x100 mesh, -8 mesh and -10 mesh with heavy liquids in the range of 2.05 to 2.15 sp gr. From the test results, it was concluded that, with the type of ore under study and a size in the range of -8 mesh, a recovery as high as 90% could be achieved with a product grade of 70% KCl. However, a final product at an acceptable recovery cannot be made with one pass, and the float must either be further processed with heavy liquids or dried and sent to a conventional froth flotation circuit. Potash ores occurring in this country consist essentially of sylvite and halite plus minor amounts of magnesium sulfate salts and montmoril-lonite-type clays. Recovery of potash minerals from evaporite ores in the North American potash fields is accomplished almost exclusively by use of amine flotation. European practice involves froth flotation as well as solution-crystallization processes. Laboratory and pilot plant test work has been reported in Europe and the U. S. on the application of heavy media separation to potash ore beneficiation. Work was probably discontinued because of lack of ore with the required very coarse liberation characteristics (1/8 to 1/2 in. liberation size). Sylvite, with a gravity of 1.99, and halite, with a gravity of 2.17, appear to be ideal for separation by heavy liquids, which are now available in gravities from 1.59 to 2.95. This paper reviews preliminary results obtained from laboratory test work on heavy liquid separation of sylvite from halite. TEST WORK The heavy liquids used in the tests under discussion were chlorobromethane, with a specific gravity of 1.923, and dibromethane, with a gravity of 2.490. These liquids, completely miscible, were combined in the proportions needed to give a mixture having the desired specific gravity. Feed for the laboratory tests was mine-run ore screened to the desired mesh sizes. In conducting the tests, the sample was fed at a constant rate into a stream of heavy liquid and the mixture directed into a small separatory vessel. The float overflowed into a collecting pan while the sink collected in the bottom of the separatory vessel and was removed at the end of the test. Approximately 500 g of feed constituted a charge. Pulp density of the feed was kept low to prevent particle to particle interference in separation. With feed in the range of 8x100 mesh, a pulp density of under 10% solids by weight was found advisable. With coarser feed the pulp density could be carried as high as 15% solids. Time of separation was very rapid. In the case of 4x20-mesh material, separation was effected in 15 to 30 sec; with -10-mesh feed, separation required about 1 to 2 min. SPECIAL EQUIPMENT Since heavy liquids are toxic to varying degrees, all separatory work was carried out in a standard laboratory fume hood. It was noted that complete removal of fumes was not being effected; therefore the hood construction was modified, resulting in a completely satisfactory arrangement for heavy liquid test work. In the interest of safety, details of this fume hood are reported here. Unlike most fumes, heavy liquid fumes tend to settle and flow like water, rather than to rise like a gas. Working on this assumption, a standard water drain was installed in the hood. Across the front of the hood a 1-in. barrier was constructed. In the rear of the hood a false back was installed, with an adjustable sliding door on both the bottom and top of this panel. As shown in Fig. 1, the exhaust fan pulled a vacuum behind the barrier, sucking the heavy fumes from the bottom of the hood. Another addition was the drying box, shown to the right of the hood. This is simply a box covered on top with hardware cloth and connected by a 6-in. inlet to the hood. Sample trays made of fine mesh wire filter screens were found ideal for drying samples. With this arrangement, air flowed completely through the sample and all fumes were drawn into the hood. In use, it was found effective to cover with a

Jan 1, 1963

By R. Wild, F. A. Wright

When a blast furnace has a certain defined burden and is operated under fixed conditions of blast temperature, etc., the fuel efficiency is determined by the extent to which the reducing gases can remove oxygen from the burden in the furnace stack. This is determined by two distinct factors: 1) The uniformity of gas-solid contact, and 2) The ease with which oxygen can be removed from individual pieces of burden. This latter is often called burden reducibility. When burdens were poorly prepared the first factor was by far the most important and a study of the reducibility of individual lumps was of rather academic interest. In recent years good burden preparation with emphasis on uniformly sized material has led to greatly improved gas distribution in the stack, and thus the second factor has become much more important and there has been a marked increase in interest in methods of measuring reducibility. This paper explores the Linder method of measuring such reducibility. The measurement of reducibility of burden materials must be carried out under conditions duplicating, as nearly as possible, those of the blast furnace stack. This is very difficult since the blast furnace process is a counter-current one, and thus the initial conditions encountered by the solid (gas temperature, composition, etc.) are the result of heat and mass transfer occurring lower down the stack. Any method of burden testing which does not take this into account is, at least to some extent, based on arbitrary assumptions. In an attempt to study blast furnace reactions under non-arbitrary conditions BISRA adopted the SCICE technique as a method of investigation. This technique has been used with a measure of success.' The SCICE technique, however, was found to be too slow for use as a routine test for burden materials and it was decided to construct additional equipment for burden testing. A test was required which would: 1) Be as realistic as possible. 2) Be quick and easy to operate. 3) Give some indication of the breakdown likely to take place during reduction in addition to a reducibility index. After a critical assessment of the reducibility tests which have been proposed it was decided to adopt the Linder test equipment and procedure as a basis for burden testing. THE LINDER TEST APPARATUS AND PROCEDURE The apparatus which was constructed (Fig. 1) was the same as that described by Linder2 except for minor changes in design. Linder also laid down a test procedure which he had derived from the results of investigations on Swedish blast furnaces. The variations of temperature and gas composition during the test were defined; these are shown diagram-matically in Fig. 2. BISRA's intention was to use the standard temperature and gas composition programmes for testing a variety of burden materials and also to investigate the influence of different programmes on standard burden materials, making use of information from the SCICE apparatus wherever this is possible. Up to the present, effort has been concentrated on the first part of the programme, and work on the second part has only just commenced. For each test 200 g of coke and 500 g of burden material were used. Linder had recommended that the coke and burden material should be between 1 and 1 1/2 in. and this was adhered to in early experiments on ores and sinters. Since the eventual aim of this work was to relate the test results to blast furnace operation, it was decided to carry out subsequent experiments using burden material in the size range used in the blast furnace, as far as this was possible. If the main interest was, for example, a comparison of the products resulting from different methods of agglomeration, then there would be advantages in using burden materials as close as possible to a standard size. After the charge had been placed in the reaction tube and this had been connected to the gas supply, rotation of the reaction tube at 30 rpm was started and the reduction programme was commenced, the gas temperature and composition being manually controlled according to the programme shown in Fig. 2. After the reduction test the charge was cooled in a nitrogen atmosphere. It was then removed from the reaction tube, the coke and burden separated, and the extent of burden breakdown assessed by screening it at 10 and 30 mesh. The extent of reduction was then

Jan 1, 1964

By R. J. Charles

PREVIOUS study' of simple impact systems indicated that energy required for fracture and size reduction of brittle materials is greatly dependent on the type of loading that is employed. In this regard it was postulated that fracture and size reduction of a brittle specimen might be accomplished by a high velocity impact at an energy level at which a low velocity impact would leave the specimen relatively undamaged. The finite time required to propagate energy away from the point of loading of an impacted object to other parts of the object and to its supports may permit a concentration of energy at the point of impact if the rate of loading is high. Localizing of energy at the point of impact could quickly initiate fractures that would propagate through the impacted material along with the traveling stress waves. As the cracks form, the initially absorbed strain energy would be released and portions of it travel ahead of the wack tips as stress waves. These secondary stress waves, combined with the initial stress waves derived from the impact, would produce high localized strain conditions in front of the fracture zone. The fracture zone would propagate and the specimen shatter. High rates of loading are brought about by short impact times, in turn obtained by employing high velocity projectiles that are of small mass compared to the impacted object. In the case of a low velocity impact, at an energy level equal to that of a high velocity impact that causes fracture, loading rate would be relatively

Jan 1, 1957