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By K. E. Merklin, F. D. DeVaney

Students of the modern blast furnace seem unanimously agreed that they are observing a major revolution in practice. Rather than changing construction and operation of the furnaces, most of the great advances now under way deal with the raw materials feed. Prepared burdens have gained increasing significance during recent years. Sizing, concentration, and agglomeration have all become bywords to greater production and reduced operating costs. The remarkable results with test furnaces in which self-fluxing sinter makes up part or all of the ore charge foretell even greater improvements. Other investigators have reported many of these advances, and it is enough to predict here that more and more operators will turn to the use of self-fluxing sinter. This change in the operator's requirements comes at a time when one major section of the iron ore regions, the Mesabi, is finding it difficult to maintain, let alone improve, the tonnage and grade of its shipping product. To meet the new demands with respect to quality and structure, the iron ore producers of this area turned first to screening, crushing, and concentrating intermediate ores. More recently, declining higher-grade reserves have led to multimillion-dollar investments in plants to treat the low-grade magnetic taconites.1,2 To yield a suitable concentrate, the Minnesota magnetic taconites must be ground to about 80 pct through 325 mesh, too fine to be charged directly to the blast furnace or to make acceptable feed to the sinter strand. The pelletizing process was developed to meet the structure requirements laid down by the blast furnace operators. Taconite producers are succeeding in their efforts just when the attention of the blast furnace man is being drawn to the great advantages of self-fluxing sinter.3 The taconite operators are ready with an alternative. Why not produce sclf-flus ing pellets? At the Hibbing laboratory of Pickands Mather & Co., where much of the early work was done that led to Erie Mining's 7.5-million ton taconite plant at Hoyt Lakes, Minn., preliminary investigations have been made. Present Taconite Operations: Details of this work will be clarified by a brief review of agglomeration practice in commercial taconite plants. After the taconite has been ground to about 80 pct through 325 mesh to liberate the magnetite grains, the ore is concentrated on magnetic separators, thickened, and filtered to about 10 pct moisture. This structure and moisture fit in closely with the conditions required to produce good balls. It has been found with taconite concentrates that at least 60 pct of the feed to the balling drum must be —325 mesh to obtain good ball compaction within reasonable time. The filter cake moisture usually permits a slight water addition at the balling drum, and further balling control is obtained by adding about 12 Ib of bentonite and 1 1/2 lb of soda ash per dry short ton of concentrate. These additives, and occasionally small quantities of fine coal, are thoroughly mixed with the filter cake ahead of balling. The swelling property of bentonite improves the green and dry strength of the balls so they can be handled without breakage until they are fired. Soda ash is added for pH control to insure maximum performance of the bentonite, and coal is added for more heat in the firing process when required. The balls, normally 3/8 to l-in. diam, are brought to a temperature of 2350" to 2500°F in the indurating furnace. Heat for this process comes from three sources: 1) fuel burned in the combustion chamber, 2) oxidation of the magnetite concentrates to hematite, and 3) coal added to the balls. Pellet strength is developed by grain growth and bridging and, if enough slag-forming constituents are present, by slag bonding. Conversion to hematite is usually close to 100 pct in the fired pellets. Batch Test Procedures: Initial studies of self-fluxing pellets were made by producing balls on a batch basis under carefully controlled laboratory conditions. The l-in. balls, uniform in size, were fired in an oxidizing atmosphere in a standard laboratory muffle furnace. The firing procedure consisted of

Jan 1, 1961

By W. L. Connell, O&apos

The properties of some heavy liquids which can be used in sink-float separations are discussed with major emphasis being given to the halogenated hydrocarbons. Data are presented on density, viscosity, thermal characteristics and the effects of moisture and temperature on the corrosiveness of the liquids towards various materials of construction. The toxicity of these materials, the hazards involved in handling them, and the procedures for handling them safely are also discussed. True liquids have certain advantages over colloidal suspensions for sink-float separation of minerals which makes their use especially advantageous when close separations, fine particle sizes or water-soluble minerals are to be treated. For best separating results, the medium should have an easily controlled, uniform, stable density. Here, true liquids have the advantages of allowing closer control of density and of not requiring agitation to maintain a uniform den- sity in the equipment. Low viscosities are desired in a separating liquid so that rapid separations of fine material can be made. Many liquids have lower viscosities than the colloidal suspensions at operating gravities of 1.4 to 5.0. For economic reasons and to prevent contamination of the product, the separating medium must be easily and completely removed from the ore. Normally, true liquids must be more completely removed than the cheaper suspensions and thus cause more problems in this area. The fact that the medium should not react with the minerals, air, water or the common materials of construction limits the choice of liquids which can be used, but many are available which are sufficiently stable. A more serious objection to the use of true liquids is that many of them present a toxic hazard to operating personnel and users of the product. Most true liquids are toxic to some extent, but modem knowledge of toxicity allows us to recognize the hazards inherent in handling them and enables us to design around them. Three types of true liquids are available for this use: 1) Halogenated hydrocarbons cover the gravity range from 1.2 to 3.3. They are generally more volatile and less viscous than the other liquids.

Jan 1, 1963

By K. W. John, L. Mueller

The rock mechanics approach to stability problems of steep rock slopes dealt with in this article has particular reference to the concept and methods of analysis developed by the Salzburg, Austria, school of rock mechanics. These theories are based on the principal postulate of the anistropy of the technological properties of in-situ rock masses. The present approach, which is more in line with geological engineering concepts, differs distinctively from the soil mechanics and engineering geological concepts commonly practiced in the United States. PRINCIPAL ENGINEERING APPROACH TO GEOTECHNICAL PROBLEMS AVAILABLE ENGINEERING CONCEPTS At the present time two different concepts are available for engineering analysis of underground structures and foundations in different earth material: soil mechanics and rock mechanics. Of these two branches, soil mechanics is, after 40 years of initially revolutionary and more recently evolutionary development, well established. Its principal concept of replacing in-situ soils or rock masses by perfectly homogeneous substitues has resulted in many practicable methods of analysis which, if applied with judgement, have resulted in successful applications throughout the world. Conversely, rock mechanics is still in the early stages of revolutionary development. Few, if any, standard concepts or methods of analysis have been established. Opinions vary widely. At present the professional background of an investigator very often determines the engineering concept to be applied to a particular problem, rather than its actual requirements if considered objectively. This paper stresses the rock mechanics or geological engineering point of view advocated by the Salzburg group. A review of the soil mechanics concepts is considered beyond the scope of this paper. THE "SALZBURG APPROACH" TO ROCK MECHANICS The rock mechanics concept as outlined here was initiated by Stini, based upon work by sander.1 Since

Jan 1, 1963

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By S. M. M. Safvi, A. Jowett

Small-scale continuous flotation tests are described in which the influence of 1) pulp density and 2) feed rate on the rate of flotation are investigated. The results provide further evidence of first-order kinetics for flotation and indicate that tests at different feed rates provide a particularly simple method of determining rate constants. Flotation kinetics are discussed in general terms. Over a period of many years numerous research workers have studied the rate of recovery of valuable materials from flotation pulps with a view to deter-mining the mechanism of the froth flotation process. Froth flo ation is fundamentally a collision process involviug at some stage collision and then mutual attachment between air bubbles and mineral particles. It is possible to draw analogies therefore between the flotation process involving mainly macroscopic particles and the ordinary processes of chemical reaction kinetics involving collisions between atoms, molecules and ions. By adopting this analogy we may postulate a general equation representing the kinetics of the flotation process: -dC/dt =K.CC ,Aa (1) The rate constant, K, will be a complex function involving reagent concentrations, particle and bubble sizes, induction times and no doubt flotation cell design. However, these are factors which can be kept constant in any series of experiments, permitting investigation of the exponents a and c, which determine the order of the reaction occurring in flotation. The various research investigations carried out so far have all aimed at evaluating a and c and in the majority of cases only since the concentration of air has usually been kept constant in a series of tests, so that a relationship of the following type has been investigated: -dC/dt = K1 Cc (2) The techniques of both batch testing and continuous testing have been used to investigate this relationship. The technique of continuous testing, in which steady-state conditions are established, permits direct investigation of the above equation. In the case of batch testing, C is decreasing continuously

Jan 1, 1961

By David R. Mitchell, Ernest M. Spokes

There is evidence that study of minerals now considered to have susceptibilities too low for magnetic separation should be continued. Present concepts may be false. INFORMATION on magnetic properties of minerals other than magnetite is scanty and sometimes erroneous. Often there is no information at all. W. R. Crane's table of tractive forces published in 1901 is still the best guide in the field, but Crane's data can be used as a guide only, since he ignored such important factors as impurities in the minerals, size, and packing effect. During the 1930's work by the U. S. Bureau of Mines added greatly to knowledge of the behavior of certain minerals in a magnetic field. The USBM bulletin gave for the first time complete magnetic characteristics for magnetite and fragmentary data for a few other minerals. The work reported here covers a study of the magnetic-susceptibility relationships of the wolframite series of minerals and sphalerites as related to their chemical compositions. Theories are proposed to explain divergences found from commonly accepted beliefs. Gouy Method for Determining Magnetic Susceptibility: The Gouy principle has been adequately described in the literature1-3 and applied in various ways.'- It may be summarized by stating that a body of material (gas, liquid, solid, or powder mixed with gas or liquid) of cylindrical shape changes in apparent weight if suspended with one end in a zone of higher magnetic field strength than is found in that portion of the field surrounding the other end of the cylinder. This may be expressed as K2 - K1 A (H2 - H2) - mg. [1] In determining the susceptibility of mineral specimens by the Gouy method, the mineral is reduced to a fine powder and packed in a glass weighing tube. Precise measurement of the bore and depth of the tube is made unnecessary by a calibration procedure. Using the basic expression of Eq. 1, Hoare"

Jan 1, 1959

By L. W. Pommier, F. B. Brien, A. K. Bhasin

In the field of mineral processing, although many operations are applied to suspensions of ore particles in fluids, there is a lack of fundamental knowledge relative to the flow resistance encountered by immersed bodies moving through solid-liquid suspensions. Previously proposed equations, based on the Ostwald-de Waele (power-law) model for non-Newtonian fluids, are exam ined in the turbulent flow region relative to a number of pertinent variables and their effects on the movement of immersed bodies. As more knowledge is acquired in regard to the effects of suspension variables on the rheological properties, useful relationships may be proposed to allow predictions of conditions to achieve optimum results for engineering purposes. In the field of minerals processing, many operations are applied to suspensions of ore particles in fluids. The differential movement of solids of various sizes and densities in suspensions is of utmost importance to gravity concentration, wet classification, flotation, hydrometallurgy, and pipeline transport of solids. The subject is also of importance to other branches of engineering; for example, those involved with soil mechanics, ceramic slips, and oil-well drilling fluids. There is a lack of fundamental knowledge relative to the flow resistance encountered by immersed bodies moving through solid-liquid suspensions. The problems encountered in the direct application of the well-known Stokes and Newton's Laws to non-Newtonian systems have been examined in a previous paper.' Equations were proposed based on the Ostwald-de Waele (power-law) model for non-Newtonian fluids. This present paper examines these equations in the turbulent flow region relative to a number of pertinent variables and their effects on the movement of immersed bodies. It is thought that, as more knowledge is acquired as to the effects of suspension variables on the rheological properties, useful relationships may be proposed to allow prediction of conditions to achieve optimum results for engineering purposes. The purpose of this paper is to make an initial assessment of the nature and magnitude of pertinent suspension variables. In order to avoid too voluminous a document, the examination of kinetics and mechanisms will be omitted at this time. THEORY A Newtonian fluid is defined as a fluid in which the stress is proportional to the rate of strain, 2 i.e.: H = viscosity; see Table I. for list of symbols The following equations are conventionally used to assess the motion of spherical particles moving under laminar and turbulent flow conditions through Newtonian fluids: When these relationships are applied in practice to suspensions, the liquid density is substituted by a composite suspension density, and viscosity is usually taken as that of the suspending liquid or an approximated "apparent viscosity." However, suspensions are non-Newtonian in character3 and thus, more than one viscosity parameter is required to describe their behavior. A number of models have been proposed for non-Newtonian fluids;2,4 among these the simplest is the Ostwaldde Waele or power-law model for incompressible fluids which for one dimensional flow may be reduced to:

Jan 1, 1970

By K. C. Ko, D. A. Pifer

The theory of probability with respect to the failure of structures is discussed. It is shown that the probability of safety, probability of failure, and factor of safety are directly related to each other and that as standard deviation approaches zero, the probability curves approach a unit step function which can be interpreted as the conventional factor of safety. The factor of safety is numerically the ratio of the load to the strength of the material at a particular point or plane in an engineering structure. Mathematically, it can be illustrated as a unit step function undefined at the point where the factor of safety is one (Fig. 1). The uncertainty of variation of soil and rock strengths require factors of safety much higher than those used in other engineering structures. The factors of safety used are generally determined from past experiences with similar structures. However, the factors thus determined do not have any theoretical meaning or do not indicate any relationship to actual safety.' It seems, therefore, logical to develop a system with a sound theoretical basis to obtain a more suitable factor of safety for structures. An application of elementary theory of probability indicates that it is possible to relate this number to theoretical and practical engineering facts.

Jan 1, 1971

By C. F. Cornell, R. C. Emmett, D. A. Dahlstrom

FOR many years the disk-type and cloth-covered drum filters have found widest application in liquid-solids separation, which uses continuous filters. The disk type is less expensive, occupies less floor space per square foot of filtration area, and requires negligible down time for filter media changing. However, because filter cloth is billowed out by the blowback in the cake discharge, it is very short-lived if the scraper blades are set too close. Consequently cake weight is very important in assisting cake discharge, which normally necessitates a minimum cake thickness of 3/8 to ½ in. Also, the vertical position of the disks eliminates the possibility of efficient cake washing for removal of soluble values. The disk-type filter is generally limited, therefore, to applications that require no cake washing and form relatively thick cakes within the normal range of filter cycle times. The conventional cloth-covered drum filter is a more flexible unit, permitting wider variation in the percentage of cycle time devoted to cake formation, dewatering, or washing, as required by the application. Thinner filter cakes can be discharged because a horizontal line across the filter medium at the location of the scraper blade is equidistant from the edge of the blade at all points. The scraper blade therefore may be set a short distance from the filter medium or, where wire winding is not used, a floating scraper can be utilized so that the contour of the cloth can be followed. The drum filter is generally employed when there are difficult filtering slurries with resultant thin cakes or when cake washing to recover soluble values must be practiced. Difficulties sometimes encountered in continuous filter operation can be summarized in the following four categories: 1) Low filtration and/or washing rates due to blinding of the filter medium. 2) Poor cake discharge resulting from sticky cakes, filter medium blinding, or too thin a cake for the particular type of filter. 3) High filter media costs caused by short life and frequent down times for cloth changing. This results from blinding or from abrasion by the scraper blade when cakes are thin. 4) Blowback of filtrate at cake discharge because not enough air passes through the cake to purge the drainage area and filtrate lines during dewatering. To reduce or eliminate blinding tendencies several modifications have been tried, for example, cover-cleaning showers, blowback under the feed slurry, and snap-blow cake discharge. In many cases, however, this simply retards the rate of blinding, since the filter medium is kept clean at its outer surface only by these methods. As blinding usually occurs within or on the back of the medium, the operator is finally forced either to remove the cloth and scrub the medium intensively with a cleaning solution or acid-rinse it on the filter—all resulting in down time, short cloth life, and additional costs.

Jan 1, 1959

By Norman F. Schulz

The technical excellence of separation achieved in a mineral concentration process, or any other process where two constituents of any kind are physically separated from each other, is expressed uniquely and quantitatively by the separation efficiency (E,): E. = (R - R,), where R is the percentage of the valuable constituent and R, the percentage of the waste appearing in the concentrate. The separation efficiency meets all the usual requirements of a true efficiency and is independent of the nature of the process used. It can be visualized as the percentage of the feed material that, in effect, actually undergoes complete separation while the rest of the feed is distributed randomly and unchanged into the two separation products. The problem of quantitatively assessing the results of a mineral concentration process recurs whenever one tries to decide which of two similar test results is more desirable. Although economic factors may greatly influence the eventual choice of process, they usually have very little to do with the technical excellence of a separation and are, therefore, not a consideration at this point. Concentrate grade and recovery, used simultaneously are the most widely accepted measures of performance, and if both the grade and the recovery are greater for one case than for another, the choice is simple. However, when the results of closely competitive operations are compared, it often occurs that one set of test results shows a higher grade and a lower recovery than another and the choice is no longer obvious. Laboratory investigations of a mineral concentration process applied to a given feed material can often be designed to yield successive paired values of concentrate grade and recovery over the practical range of the course of a separation process.' These results may be plotted as some combination of the cumulative values of grade, weight recovery, or unit recovery in the concentrate. The resulting correlations for two slightly different sets of conditions can then be compared qualitatively, or even quantitatively in particular cases.2 Such comparisons are not applicable, however, to individual concentration test results. The reduction of a separation test result to a single efficiency figure, preferably as a percentage of ideality, would afford the simplest and most precise means for comparing the results of competitive tests. The effective proportion of feed material that undergoes complete separation in any real concentration process, hereafter referred to as "separation efficiency," is such a figure. It is essily calculated from the usual feed and product data and can even be evaluated, in many cases, from data in publications describing various concentration processes. Mineral Concentration Processes and Their Evaluation Mineral Concentration Processes: The overall concentration process may be subdivided into a liberation process which severs the bonds between unlike mineral species, and a separation process, which sorts the resulting liberated particles into two products—a concentrate rich in mineral A and a tailing lean in mineral A. A perfect mineral concentration process is represented by the idealized scheme shown in Fig. 1 in which liberation and separation are not differentiated. The feed mixture consists of two mineral types which are completely liberated early in the process and then sorted by type. All the mineral particles of type A which respond uniquely to the selection forces acting in the ideal separator, report to the concentrate, while all the mineral particles of type non-A report to the tail-

Jan 1, 1971

By R. C. Ferguson, Harlowe Hardinge

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By F. W. Bowdish

A theoretical and experimental laboratory study has shown that the rate of breaking of oversized particles in a ball mill is proportional to 1) the concentration of such particles, and 2) the area of the balls, provided that the balls are larger than a certain size relative to the particles. If all undersized material were removed as formed, the rate of breaking would be almost independent of the quantity in the mill. Rate equations are given which are similar to those for zero, first, and second order reactions. Deviations of small balls from the equation for a second order phenomena are quantified by a factor called "nipping efficiency." Some data are correlated by a dimension-less equation, which is used to estimate nipping efficiencies from data not in the correlation. The investigation described here was undertaken after a literature survey revealed that little was known concerning the mechanism of breaking of particles within a ball mill and the effect of this mechanism upon grinding rates. One paper, by Roberts1 , offered a theory concerning the mechanism of breaking, and a rate equation for the breaking of particles larger than some size to particles smaller than that size. That paper provided the starting point for a theoretical and experimental investigation of grinding under carefully controlled conditions. THEORETICAL CONSIDERATIONS In his paper on the probability theory of ball mill grinding, Roberts theorized that "... a given expenditure of power leads either to a certain number of point to point blows per bp-hr or to a certain distance of line contact per hp-hr ..." He assumed that the balls touch each other during each blow or during moving contact. Then he said that a volume of pulp, dependent upon the size being considered, would be covered per minute within which all oversized particles would be crushed to particles smaller than that size. Naturally, the higher the concentration of oversized particles in the pulp, the higher the rate of which they would be broken. From this reasoning Roberts obtained the differential equation -dC/dt = k hp/ton C (1) where k is a constant for any size of particle, density of particle, and moisture content of pulp. C is the percentage of oversize in the material to be ground, t is the time, and hp/ton is the power input per ton of material in the mill. He then integrated the equation to log Co/Ct = k/2.3 hp/ton t , where Co and Ct are the percentage of oversize at time zero and t respectively. By always using the same quantity of material in a mill and assuming that it required constant power to run the mill, he was able to combine the horsepower per ton factor with the other constants to give an equation for a first order rate.

Jan 1, 1961

By P. K. Guerrero, N. Arbiter

The correlation of power consumed by a tumbling mill with the dimensions, speed, and load has been attempted by three principal methods. One of these, the torque formula, has been reviewed critically elsewhere.' This approach, while analytical, disregards the individual motion and speed of the tumbling bodies and assumes that the gross geometry of the entire load is sufficient to establish the power requirement. A more general approach, using dimensional analysis, has been applied in this laboratory by the present authors1'1, and more recently by Rose.3 This disregards the internal conditions within the mill entirely and in effect is an empirical procedure for correlating data. Finally, the complete analysis of tumbling mill dynamics as first made by Davis' attempts to develop the power requirement from individual ball and rod paths and velocities. Although Davis' result is applicable only at a single speed for a given size and load it could be useful, since his optimum speed is at the upper limits of conventional tumbling mill practice. Where power values predicted by the Davis equation were compared with actual mill requirements, they were found too low for small mills and much too high for the largest mills. This suggested that the agreement with actual intermediate range requirements might be fortuitous, and the development of the equation was examined critically. The probable source of error was found to lie in an erroneous assumption made by Davis in converting the energy of ball or rod motion to power. With the same approach, but with different assumptions, a modified equation has been obtained which predicts results that are usually consistently high. This is to be expected, since it is assumed that there is no slip between load and shell and no interference among tumbling bodies. Both these effects should account for the difference. Development of Equation: When a cylindrical shell charged with a substantial weight of rods or balls is slowly rotated about a horizontal axis, the load surface first assumes an inclination that gradually increases to a maximum. Continued rotation results in failure within the load, characterized by a sliding of the top layers down the slope and a reduction in the angle of inclination. Further rotation again increases the angle, and the cycle repeats. If the speed is increased, there is continuous load failure and build-up and a continuous stream of rods or balls descends over the upcoming mass, which moves in circular arcs. At higher speeds the up-

Jan 1, 1961

By Rodney L. Stone, David S. Cahn

As part of an over-all laboratory and plant test program to determine mixing ad other techniques for a more economic use of bentonite as an iron ore pellet binder, a settling test for measuring the bentonite content within a pellet or a batch of pellets has been developed. Through this test, it has been shown that dry fracture load is directly proportional to bentonite content in batches of pellets, but inversely propor tional to bentonite content in individual pellets of the same batch. This apparent paradox is explained by a negative correlation between a pellet's size and its bentonite content and the fact that large pellets can withstand a higher fracture load than small pellets. In past attempts to characterize green and dry iron ore pellets, the authors found large variations in the results of strength tests on closely sized pellets from the same batch.' Reasons advanced for such variations in the quality of pellets in which the binder is bentonite were the presence of internal flaws and the nonuniform distribution of the binder. Even when resulting pellet strength is good, there is still the question of whether the same strength could have been attained through a more effective use of a smaller amount of bentonite. The authors' own and other studies 2-3 have pointed out the strong dependence of pellet strength on the degree of "mixedness" of the iron ore concentrate and the bentonite binder. However, before such a relationship could serve as a guide for the development of better mixing techniques and the associated economy of

Jan 1, 1969

By W. V. Long, D. H. Reynolds, R. B. Bhappu

Equilibrium distribution coefficients were determined for some monovalent and polyvalent ions for several absorbing and exchanging media as functions of pH and initial ion concentration. Among ions studied were NH4 (I), K(I), Ni(II), Ce(III), Th(IV), and Mo(VI). Absorbents used were: ion exchange resins; a tertiary, and a quaternary amine in an inert solvent; and some activated carbons. The ions studied were in aqueous HCI, HNO3, H2SO4, or HC104 solutions. The classical mass action expression proved of little value in describing the variations observed. The data indicate that the determining factors in absorption and ion exchange may be the ionic species in the aqueous solution, which in turn depend upon pH, dilution, and perhaps other conditions. Three sorption processes have become prominent in recovering values from leach solutions. These are: 1) ion exchange, 2) solvent extraction (liquid ion exchange), and 3) physical adsorption by materials of high specific surface. In the first two of these processes the exchange reaction is considered chemical in nature, and is discussed in terms of the mass action principle. The third is considered a surface phenomenon and is discussed in terms of surface areas and various forms (Gouy, Stern, Grahame, etc.) of the Helmholtz double-layer theory. Generally speaking, we have found in chemical analysis and hydrometallurgical studies that there are close similarities between the three sorption processes, and also differences which are difficult to reconcile with any of the usual theoretical treatments. That others have experienced the same troubles is shown by the fact that usable processes, in chemical analysis as well as in hydrometallurgy, generally are discovered by empirical procedures. EXPERIMENTAL CONSlDERATlONS The gaps between theory and practice have allowed many conflicting statements to appear in the litera- ture. To learn, in particular, the behavior of some metallic ions on a cation exchange resin at a constant pH, equilibration studies were made over a wide range of concentrations.' Results are shown in Fig. 1 for four metals in hydrochloric acid solution on a strongly acidic cation exchange resin, AG 50W-X8 (Bio-Rad). The distribution coefficient, defined as D = amount of metal extracted per g of resin/amount of metal unextracted per ml of solution is seen to vary in a way to support all manner of reports about its behavior with changes in solution concentration. The metals used were chosen from those least likely to form stable chloro-complexes. The changes of slope in the various isotherms may reflect formation of unsuspected chloro-complexes of low stability (ion-pairing for example) whose concen-

Jan 1, 1967

In 1962, minerals beneficiation showed definite signs of coming out of the doldrums into which it has drifted after the uranium boom. Demand by steelmakers for high-grade iron pellets has created the need for concentrating plants to upgrade specular hematite by spiral concentration or flotation and magnetites by magnetic means and flotation. According to a paper given at the AMC annual meeting by C. S. Arms of Pickands Mather, "the shipment of pellets, started in the mid-1950's, has increased to 17.6 million tons in 1961 from North American operations. Many plants already underway will raise this production to about 42 million tons by 1965. This will be a very sizeable percentage of the total iron ore production contemplated by that date." Iron ore concentration methods are gradually raising the assay of pellets into the 65 to 69% Fe range and lowering silica to the 4 to 7% range. Quebec Cartier's Lac Jeannine concentrator went into operation in late 1961 and all grinding units were running by 1962. The Iron Ore Co. of Canada's new Carol Lake plant was started in mid-1962.

Jan 2, 1963

By J. N. Murphy

The use of the National Science Teachers Association (NSTA) regional meetings as a forum to educate teachers on the role of minerals in our society has been immeasurably successful. Each year, on average about 3,500 to 4,000 teachers participate in each of the three regional NSTA meetings. At each of these meetings, nearly 2,000 teachers visit the “Minerals Coalition Booth” and receive free mineral samples and other educational support materials. Some rules of thumb for a successful booth at the NSTA are provided. Considering the level of effort for such a booth, are there alternatives? The NSTA also organizes a national annual meeting with attendance of about 20,000. All who have participated in a regional NSTA agree that the traditional approach of providing a mineral set and teacher materials for a substantial fraction of the 20,000 attendees is not feasible. Conceptual alternatives for minerals education at the national NSTA meeting are introduced and evaluated.

Jan 1, 2001

By Robert A. Clifton

It seems probable that more and more minerals will be used in plastics as time goes on. This must occur because the minerals will; aid in processability, enhance the engineering properties, reduce the cost, and above all conserve large amounts of petroleum based chemicals. This paper examines the advantages and disadvantages of using the minerals and the ability of the domestic minerals industry to support this possibly very large undertaking. In the author's opinion, it can. This talk is excerpted from a forthcoming Bureau of Mines Information Circular.

Jan 1, 1985

By Elmer W. Pehrson

Minerals have played a prominent role in the affairs of man and nations since time immemorial but with the advent of the industrial revolution, expanding use of the products of the mineral kingdom became synonymous with the material progress of mankind. Individuals and business groups have recognized this fact so have been aggressive in searching for and utilizing mineral resources, with notable success. Indeed, their achievements are primarily responsible for the present Machine Age. On the other hand, until recently the body politic has failed to grasp fully the essentiality of minerals although, at times, access to mineral supplies has been an important factor in the political maneuvers of nations. This chapter is concerned primarily with the broader political and economic relationships of mineral resources to the world economy, and with the internal and external actions of nations in the field of minerals. Many factors that condition the mineral economy of regions and the world as a whole, influence the internal and external affairs of individual nations. The more important of these are: 1. The nature of man's dependence on minerals. 2. The problems resulting from the unique characteristics of mineral occurrence. 3. Geopolitical concepts. 4. Minerals and world power. Actions of individual nations affecting their mineral economy seek two basic objectives : 1. Attainment and maintenance of prosperity. 2. Military security. Other objectives may be cited but on careful analysis they become integral parts or extreme phases of the two points noted above. For example, the twentieth century aggressions of the USSR, Germany, Italy, and Japan might be considered as direct bids for world power but they also can be interpreted as exaggerated or mistaken efforts toward military security and peacetime prosperity. Similarly, trade barriers, nationalization of industries, foreign exchange manipulations, taxation, monopolistic prac-

Jan 1, 1959