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By F. X. Tartaron

This paper deals with basic physical phenomena, which when combined and interpreted, lead to the same mathematical equations that describe comminution phenomena. Thus, a physical model is described that corresponds to the mathematical model presented in the writer's previous papers. '12 In the mathematical model, the energy consumed in breakage is related to the volume or weight of material broken and the size of particles broken. The equation E=2.303 Ck-a log x1/x2 was derived by multiplying the volume or weight of each size in an ideal Gates-Gaudin-Schumann size distribution by an energy factor. The product of these two factors gives the energy distribution among the different sizes in a single size distribution. The energy of breakage of a specific constant weight of one size distribution to another size distribution is given by the equation E = constant/kn-1. In this case, where the volume or weight is constant, the energy is proportional to the size factor 1/kn-1. In what follows, a physical theory will be presented showing that the energy consumed in comminution is proportional to the volume or weight of the material broken and to the reciprocal of the size of this material raised to a constant exponent. THE VOLUME FACTOR The atomic theory of matter reveals that in solids, atoms or ions are arranged so as to be in equilibrium at specific distances from one another. Although the atoms or ions are oscillating, there is a definite determinable mean distance between them and this distance is a balance between repulsive and attractive electrical forces. It therefore requires force to separate the atoms or ions and when an outside force is applied, it first produces strain in increasing the distance between the atoms or ions. This strain increases to the breakage limit on application of sufficient force. In brittle materials, there is negligible plasticity and when an elastic limit is exceeded, breakage takes place. The work done is the force applied per unit area times the cross sectional area of the ideal particle multiplied by the maximum strain per unit length at right angles to the area times the length of the particle. Thus the work done is proportional to the area times the length, which is equivalent to the volume of the ideal particle. If more than one feed particle is considered broken, each particle must be subjected to sufficient strain so that the breakage limit of its contained atoms or ions is reached in order for the particle to be broken. Thus, the energy of breakage is proportional to the total volume of the particles broken. If the particles are of different sizes, the size factor must be included to get a correct determination of energy of breakage. In the preceding, it has been assumed that there is a constant binding force between the atoms throughout the volume being strained. This, of course, is not true. It is known that there are many irregularities in the structure of matter and the binding force differs markedly in different portions. But the differences are only discernible by examining extremely small subdivisions of matter. In one order of magnitude of volume, cracks can be discerned separately from non-cracked neighboring material. In a smaller subdivision of volume, lattice dislocations can be isolated. When these situations are brought into focus, mechanisms of their behavior can be learned, leading to a fuller understanding of phenomena that occur in larger scale subdivisions of matter. Very often, however, the mechanisms that operate in small scale subdivisions have negligible effect in those of large scale, and there still is a place for deriving a mechanism for large scale conditions. The quantum theory is extremely valuable for use with photons and electrons, but is of negligible use with ordinary atoms and molecules. This paper deals with relatively large scale subdivisions of volume present in comminution phenomena. Hence, the effects of cracks, lattice dislocations, misplaced atoms, etc., are smoothed out in an average, constant for each relatively large subdivision of volume. This attitude is supported by experience. If two 10 cc samples of the same ore were ground identically, the same product would be obtained. However, if two samples, each a cubic micron, were conceived to be broken, then one sample might contain a crack and the other not, hence a different product would be obtained. Experience shows that ordinary samples used in comminution, behave as though no irregularity existed

Jan 1, 1964

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 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 Gordon E. Agar

Several correlations have been proposed to relate energy consumption and size reduction in comminution, and although these are arrived at from different starting points, it is postulated that they are all statements of the same basic concept. To test this postulate, the dry ball mill grindability of langbeinite was determined, and the results were analyzed according to the various methods. An examination ' of the energy-particle size relationships proposed in the recent past revealed that they had much in common. In fact, the conclusion was tentatively drawn that these expressions were essentially the same, even though they were arrived at from quite different basic premises. The equations attributed to Charles2 and Holmes3 were shown to re- duce to the same single equation, W - A K-3, and the simulated continuous test used by Maxson et al.4 was shown to give the best experimental value of the required empirical parameter, ß. Furthermore, the size distribution parameter a in the Schuhmann equation, y - (x/K)2, was shown to be equal to ß in many cases. The kinetic approach used by Arbiter and Bhrany5 should give the same results as the Maxson test procedure; however, the batch operation used by Arbiter and Bhrany could not produce data that would be as precise as the simulated continuous technique. To test the validity of these conclusions, the simulated continuous method was used to obtain data on the dry grinding of langbeinite. The results are discussed in terms of the various equations and approaches. APPARATUS AND MATERIAL A ball mill with internal dimensions of 17.5 cm (6 7/8 in.) in length by 19.4 cm (7 5/8 in.) in diam was loaded to 50% of its volume with balls 3.2 cm in

Jan 1, 1969

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 R. J. Stoehr, F. Howell

Since the ore mined in the Grants-Ambrosia Lake uranium area is taken from a water-saturated sandstone formation, part of the milling operation includes a drying process. The authors discuss the merits and disadvantages of two methods: natural drying and mechanical gas-fired drying; operating and cost data are included. In the Grants-Ambrosia Lake uranium area, approximately 7000 tons of ore are mined each day from a water-saturated sandstone formation. This ore is sampled and treated at four separate milling operations. Prior to the design and initial construction of the milling facilities, very little was known as to the physical properties of the ore which would be handled in the sampling and crushing facilities at the mills. All four mills initially were constructed without ore-drying facilities. As mining proceeded, it was determined that, even after drying the ore in place by use of drainage development underground, the moisture content of the ore as hoisted averaged approximately 18 pct. Actually, some of the mines produced ores which resembled cream-of-wheat in character and had a moisture content in excess of 24 pct. Each of the four mills had a differently designed sampling and crushing plant; therefore, the moisture content of the ores had varying adverse effects upon the operation of the plants. It was determined after some experience that most of the sampling and crushing facilities could not economically handle ore if the moisture content was greater than 10 pct—one operator considers 12 pct moisture as a maximum. In all of the plants, moisture content over 8 pct increases costs to some extent. In addition to the added sampling and crushing costs, some of the mills were located 15 to 20 miles from the mines and the high moisture content of the ores increased the haulage costs from the mine to the mill. It was generally conceded that it was necessary to employ some method to reduce the moisture content of the ore prior to sampling. Several methods were considered: 1) natural drying at the mine site, 2) mechanical gas-fired drying at the mill site, 3) mechanical gas-fired drying at the mine site, 4) radiant heating on concrete clabs, and 5) infrared shed drying. After preliminary investigation all possibilities but natural drying and mechanical gas-fired drying at the mill site were eliminated. NATURAL DRYING A cooperative feasibility study was conducted on the natural drying method. This study indicated the following: 1) During summer months May through September, ore piled 4-ft deep would dry from +20 to +10 pct HzO in 30 days with no mechanical turning. 2) In the first four days after piling a reduction of 5 to 7 pct total moisture content was accomplished by drainage and absorption of the water into the underlying ground on which the ore was piled, providing the ground surface absorption rates approached 10 gal per sq ft per day. After the first four days the decrease of moisture caused by drainage was practically nonexistent. 3) In the summer months drying by evaporation amounted to approximately 0.1 pct decrease in total moisture per day and mechanical turning of the ore could increase this as much as three times. 4) Climatic conditions of the area indicated that: a) Air temperatures averaged 64' May, June, July, August, September; 41" October, November, December, March, April; 31" January, February. b) Average rain fall was 10 in. with 40 pct of this occurring during July and August. c) Only on very rare occasions were there over five consecutive days with below freezing temperatures. d) The annual evaporation rate was approximately 75 in. On the basis of this study, two of the mill operations decided to attempt to meet the moisture specifications by natural drying supplemented by mechanical turning. The other two operations decided upon rotary gas-fired drying at the mill site. The natural drying method has been quite successful in practice. The ore running approximately 18 pct moisture is trucked from the head frame and piled in rows about 6 ft wide, 3 1/2 ft deep, 5 feet apart during the months of May through September. In approximately 30 days this ore will

Jan 1, 1961

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

By O. E. Palasvirta

Close examination of most so-called n.eu processes in mineral dressing reveals that they were conceived and developed a long time ago. High-intensity magnetic separation is no exception. Although its application in iron ore beneficiation seems new, in Germany the process has been commercially successful for almost two decades. High-intensity separators have long been standard equipment in the glass sand industry; they have also been used in concentrating some of the more unusual para-magnetic minerals. The basic features of an induced-rotor, high-intensity magnetic separator were described by Desire Korda in Paris in 1905,' and many patents have since been issued for improvements on the original design. Fig. 1 shows the basic design of an induced-rotor separator. Several U. S. manufacturers have been supplying induced-rotor separators for many years, but these are all devices of relatively low capacity, not suitable for treating such low-cost commodities as iron ore. On the other hand, two German manufacturers are offering induced-rotor separators of high capacity, both proven in the field. The present investigation was prompted by the apparent success of the German installations and by the discovery of iron ores on this continent which may be amenable to high-intensity separation. The objective was to gain better insight into the underlying principles of separation in a high-intensity field and to use this knowledge in selecting or designing an optimum separator. THEORETICAL CONSIDERATIONS he well known low-intensity magnetic separators are. without exception, of the the magnet type with more or less similar field shapes. It is therefore permissible to speak of their comparative strengths in terms of field intensity alone. However, proper understanding of the high-intensity magnetic separation process requires somewhat deeper analysis of the relationship between the magnetic field and the attractive force. Contrary to a widely held belief, a magnetic field of parallel lines of flux exerts no attractive force on a magnetically susceptible particle, but merely aligns its poles parallel to the lines of flux. On the other hand, a non-uniform field attracts a magnetically susceptible particle in the direction of the converging lines of flux. Thus the force of magnetic attraction varies as the product of the field intensity and the field gradient: dH ds As the induced-rotor magnetic separator is primarily intended for separating feebly magnetic materials, it is important that maximum attraction be attained. For the designer, then, the relationship between the magnetic field and the attractive force has the following implications: 1) Insofar as possible, the separation gap must be saturated with flux. 2) The rotor surface must be shaped to insure maximum convergence of lines of flux toward the rotor. 3) The primary pole face (opposite the rotor, on the other side of the separation gap) must be shaped so that the lines of flux do not diverge from it at any location in the effective separation zone. The actual attractive force of the rotor, and its critical speed, can be calculated as shown by Runo-linna, provided that the true flux density, the field gradient, and the magnetic susceptibility of the mineral can be measured with sufficient accuracy. The irregular shape of the rotor surface makes flux measurements difficult, but it can be done if a suitable gaussmeter is available (see Fig. 2). Calculations by the author, and comparisons with the centrifugal force exerted by the rotor, indicate that the force pinning a hematite particle to the rotor is of the same order of magnitude as the force holding a

Jan 1, 1960

By M. E. Wadsworth, T. D. Chatwin

The kinetics of sulfidization of hematite (Fe203) by H2S gas at various partial pressures have been determined over the temperature range of 170° to 500°C. The process was controlled by thin films of pyrite (FeS2) and troilite (FeS) which were protective and formed diffusional layers. At higher temperatures the thermodynamic drive to form FeS2 decreased, and consequently the rate of formation of FeS, decreased. Conditions for the development of thin layers of FeSz consisting of 10 to 100 atomic diameters are presented. Laboratory flotation results on synethetic ores indicate floatability identical to that of pyrite of normal xanthate flotation. The prospect of high-temperature sulfidization of metal oxides using H2S, S, or other sulfur bearing compounds is intriguing as a potential activation process for flotation since the thermodynamics are very favorable for a large number of systems. If the kinetic parameters are sufficiently well understood it may be possible to provide surface sulfide layers a few angstroms thick, thus imparting normal sulfide floatability characteristics to the oxide. For thin layers the total sulfur pick-up would be small except for very high surface area materials but in most cases this sulfur could be readily removed by roasting if necessary. The general reactions for the sul-fidization of metal oxide MxO, to a metal sulfide MpSq using H2S may be written for the three conditions py > xq; py < xq; and py = xq: A large thermodynamic advantage is obtained in many systems because of the relatively large negative free energy of formation of H,O compared to H,S. For the temperature range 2S° to 1500°C &apos; where AFOis the standard free energy and T is absolute temperature. As is evident from Eqs. 4 and 5, the free energy of formation of H ,0 is approximately 35 to 37 kilocalories per mole greater than that for H,S over broad variations of temperature. For oxide-sulfide systems as outlined above, sulfidization may be expected to occur for the conditions: Groups I and 111: that is, if the terms on the right are more negative than those on the left the reaction is expected to occur thermodynamically. The value hFg,(,) is included in Eq. 6 since hFLzs is based upon gaseous S2 rather than the normal reference state, orthorhombic sulfur. The

Jan 1, 1969