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By Howard R. Keil

Reactor at Carlton mill successfully calcines pyrite-telluride flotation concentrate for cyanidation. CARLTON mill in the Cripple Creek district, 45 miles southwest of Colorado Springs, Colo., has been in operation for approximately three years, treating the custom sulpho-telluride ores formerly handled by the Golden Cycle mill in Colorado Springs. At the new mill the pyrite-telluride flotation concentrates have been roasted, prior to cyani-dation, in a Dorrco FluoSolids reactor, which has proved capable of yielding calcine readily amenable to cyanidation. Incoming ores are crushed at Carlton to % in. and sampled in 30 to 75-ton lots before the individual lots are mixed in the fine ore bin. The composite ore then goes to the grinding circuit. Classifier overflow from the grinding circuit is treated by flotation to produce a concentrate containing approximately 85 pct of the values in 5 pct of the weight. The flotation concentrate goes to the FluoSolids reactor prior to cyanidation. Cyanide solution values are recovered by zinc dust precipitation. Flotation tailings are thickened and cyanided, the values from this part of the circuit being recovered from solution by activated carbon. Gold values in the flotation concentrates consist of a small amount of finely divided metallic gold; the tellurides, sylvanite and calaverite; and gold very intimately associated with the pyrite. The metallic gold and a part of the telluride values can be extracted from the concentrates by cyaniding without roasting. However, extraction on the tellurides by this method is not satisfactory, and the pyrite values are so intimately associated with the pyrite that very fine grinding and cyanidation for a prolonged period under ideal laboratory conditions does not yield good recovery. Upon roasting, the tellurides are oxidized to very finely divided metallic gold which is readily soluble in cyanide. The pyrite is oxidized mainly to a porous iron oxide which allows cyanide solution to penetrate to gold particles and take them into solution. Feeding the Reactor Flotation concentrates move by gravity to a thickener. Thickener underflow is discharged at approximately 60 pct solids, by means of a diaphragm pump, into a storage and blending agitator. From here the concentrates are pumped to a disc filter. The filter cake, at 84 to 88 pct solids, drops directly into a repulper where it is repulped into a slurry of 78 to 80 pct solids. The repulped filter cake flows into a 6x6-ft slurry agitator from which it is fed directly into the reactor by the Moyno pump. The Moyno pump has eliminated a large number of difficulties that were first encountered in feeding a wet concentrate to the reactor. On the original installation in Canada a screw feeder was used to introduce the filter cake directly into the reactor. Since the reactor is under a slight positive pressure it was necessary to have a suitable seal to prevent gas escaping from the reactor at this point. When the concentrates produced in flotation were coarse or granular they could be readily fed to the reactor but when fine-textured, clayey concentrates were produced a great deal of trouble was encountered.

Jan 1, 1955

By G. O. Prickett, P. Somasundaran

An experimental progmm based on statistical principles is used for optimizing a flotation operation for the separation of minerals. In this case, flotation of sylvite from sylvinite ore with amine and starch as the flotation reagents was studied to find the optimum reagent levels at various slime levels. The experiments were first done on a two-level three-factor design. Analysis of the results showed that the interactions between various factors were significant. Hence, a first-order linear model of the flotation response would not be accurate. The two-level three-factor design was expanded into a composite design in order to test the fit of a higher-order model. Data were treated to obtain an expression for flotation response as a function of the different variables, to predict the optimum reaaent levels at various slime levels, and also to determine the path of steepest ascent. Tests around the predicted optimum and along the path of the steepest ascent were done until a near-optimum region was obtained. Exploration of the nature of the response surface in that region established the actual location of the maximum response. The optimum amine level was found to be 0.20 lb of amine per ten of ore at all slime levels, while the optimum starch level varied as a function of the slime present. Computer access at the laboratory by means of a teletype terminal on a commercial time-sharing system enabled efficient and accurate treatment of experimental data. The efficiency of experimental investigations can be improved significantly by the use of experimental design and data analysis based on statistical principles. The common approach used in experimental research is to study a property of a system as a function of "one-variable-at-a-time". Several authors'-4 have pointed out the inadequacy and deceptiveness of this method. Results obtained for the effects of a variable on the measured response of the system using

Jan 1, 1970

By H. J. Modi, P. L. De Bruyn

IN recent years interest has been aroused in flotation rate studies both from a technical and a more practical aspect. With increasing fineness of grind becoming a necessity in treatment of low grade ores, difference:; in flotation rates of various size ranges will assume importance in the evaluation of overall recovery results. Theoretical analysis of flotation kinetics was initiated by Schuhmann,' who sought to define flotation rate of a given mineral and specified size range in terms of certain physical variables of the flotation system. Using as a basis for his analysis the bubble-particle encounter hypothesis previously advanced by Gaudin,' Schuhmann writes for the rate of flotation (R,.) of particles of average size x, Rx = PcPa(x) V F. [1] In this relation Pc and Pa represent, respectively, the probability of collision between particle and gas bubble in a given time interval and the probability of adhesion of particle to the bubble after collision. Rigorously speaking, the term PC does not have the significance of a probability factor, since it depends on the time interval and therefore can be numerically greater than unity. The term c(x) expresses the concentration of particles of an average size x in the flotation cell in weight per unit volume of pulp, and V denotes the total volume of pulp in the cell. The product, P,c(x)V, defines the frequency of particle-bubble collisl.on, and when it is multiplied by P,, the number of collisions per unit time resulting in successful adherence is obtained. The factor, F, measures the froth stability and is introduced because after fruitful bubble-particle contact some particles may still riot reach the froth launder. Coalescence of bubbles in the froth column or too short a bubble life could account for this effect. Schuhmann also introduced the concept of "specific flotation rate" (Q) to allow for the relative abundance of different mineral species in the flotation pulp. The specific flotation rate, having the dimension of reciprocal time, is defined as Eq. 2: Qx =Rx/c(x)v [2] and it also follows from Eq. 1 that Qx = PcPaF. [31 The theoretical aspects of flotation kinetics were also considered by Sutherland,3 who tried to improve Eq. 1 by expressing the probability factors in terms of such physical variables as particle radius (x), bubble radius (X), velocity of bubble relative to particle (v), the number of bubbles per unit volume of pulp (N) and the induction period (A) necessary for air-mineral adhesion. According to Sutherland's treatment, the rate equation, Eq. 1, transforms to R, = [3pX x v N] [sech2 (3 v ?/4 X) ] ?c(x) V [4] where the first term in brackets on the right-hand

Jan 1, 1957

By T. M. Morris, W. E. Horst

W. E. Horst—In regard to the flotation rate being described as "first orcler" for flotation of quartz particles below 65 p in size (or any size studied in this work) in this paper, it appears that the authors' conception of rate equations is not in agreement with cited references. A first order rate equation has as one of its forms the following: a In.=a/a-x=kt where a = initial concentration, a—x = concentration at time t, t = time, and k = constant. The constant, k, has the dimension of reciprocal time which is similar to the specific flotation rate, Q. described by Eq. 2 in the authors' article, as has been previously discussed by Schumann (Ref. 1 of original article). The plotted data presented in Fig. 4 of the article utilizes the specific flotation rate, Q (min.'); however, there is not adequate data given to indicate the order of the rate equation which describes the flotation behavior of the quartz system studied. Results from the experimental work indicate that the relationship between rate of flotation (grams per minute) and cell concentration (provided the percent solids in the flotation cell is less than 5.2 pct and the particle size is less than 65 p) is described by an equation of the first order (R, = k c+", n being equal to 1 in this size range) and the use of the first order rate equation does not apply. Similarly the relationship for other particle size ranges studied is expressed by equations of the second or third order depending on the magnitude of n. T. M. Morris—The authors are to be commended for the experiments which they performed. As they state in their discussion the concentration of collector ion In solution did change with change in concentration of solids in the flotation cell. Since for a given slze of particle, flotation rate increases with concentration of collector until a maximum is reached, the effect of concentration of particles in their experiments was to vary the concentration of collector ions. A collector concentration which insures maximum supporting angle for all particles eliminates the unequal effect of collector concentration on various sized particles and the effect of size of particles and concentration of particles upon flotation rate could be more clearly assessed. I believe that if the authors had increased the concentration of collector to an amount sufficient to attain a maximum supporting angle for all particles they would find that the specific flotation rate of particles coarser than 65 p would be constant with change in the concentration of solids in the flotation cell, and that a first order rate would apply to the + 65 as well as to the —65 p sizes. It might also be discovered when this change in collector concentration was made that the maximum specific rate constant would be shifted toward a coarser fraction than when starvation quantities of collector are used since this practice favors the fine particles and penalizes the coarse particles. P. L. de Bruyn and H. J. Modi (authors' reply)—The authors wish to thank Professor Morris for his kind remarks and for mentioning the influence of equilibrium collector concentration on flotation rate. With a collector concentration sufficient to insure maximum supporting angle for all particles, a first order rate equation may indeed be found to be generally applicable irrespective of size. Such a concentration would, however, lead to 100 pct recovery of the fine particles and consequently defeat the essential objective of the investigation to derive the maximum information on flotation kinetics. To establish absolutely the validity of any single rate equation for a given size range, the ideal method would be to work with a feed consisting solely of particles of that size range. Use of such a closely sized feed would also eliminate the possibility of the interfering effect of different sizes upon one another. The authors do not believe that increasing the collector concentration would shift the maximum specific flotation rate (Q) towards a coarser fraction. Experimentation showed Q to be independent of solids concentration for all particles up to 65 µ in size, whereas the maximum value of Q was obtained in the range 37 to 10 p. Professor Morris contends that the addition of starvation quantities of collector favors fine particles at the expense of coarse particles, but the reason for this is not entirely clear to the authors. The comments by Mr. W. E. Horst are concerned only with the concept of the term "first order rate equation." According to the usage of this term in chemical kinetics, time is an important variable, as is shown in the equation quoted by Mr. Horst. All the experimental results reported by the authors were obtained under steady state continuous operations when the rate of flotation is independent of time. To be consistent with the common usage of the "first order rate equation," it would be more satisfactory to state that under certain conditions the experimental results show that the relation between flotation rate and pulp density is an equation of the first order.

Jan 1, 1957

By U. N. Bhrany, J. H. Brown

The specifications of particle size and the size analyses of fine particulate materials are commonly presented without reference to the method of analysis. A review of the various sizing methods showed that not only numerous sources of error but also different definitions of size are inherent in the different methods of analysis. To demonstrate the magnitude of the differences, the size distribution of finely crushed quartz was determined by several methods and a critical comparison of the data was made. The experimental results show that appreciably different sizes may be indicated for one material depending on the method of analysis used. However, because the shape of the size-distribution curves determined by almost all methods was constant, a simple method of correlating the data can be adopted. Conversion factors, by which size measurements obtained by the various methods can be expressed in terms of screen sizes, are presented as one method of correlating the results. The size analysis and size specification for lump rock and particulate materials are commonly determined with sieves. With this technique, a readily reproducible measure of size can be obtained merely by duplicating the sieve and the method of sizing. For coarse materials, such as lumps of rock, this technique is simple and effective; and the effects of properties of the material such as shape can be readily understood. When very small particles are measured, however, problems arise. Small sieves are difficult to construct and maintain, and size separations become sensitive to the analytical procedure. For these difficult sieve sizes and for still smaller particles, other sizing techniques employing, for example, sedimentation, light extinction, and microscopy are available. Because these various techniques employ different size-dependent responses to establish size, it is appropriate to examine them and the results they yield to determine their relationships and their limitations. The 'size' of a lump of rock is not an absolute quantity, but rather a defined one. With sieves, for example, the size is defined by the screen aperture and no reference is made to the actual shape of the particle, yet only in the case of a regular body, such as a sphere, can the sieve size be related to the shape. With sieves only two dimensions of a particle can be suggested, for the sieve makes no allowance for the length of the particle passing through an aperture. Similarly, the size of a particle as measured by a microscopic technique need bear no relationship to the size as measured by sieves, although the two measurements may be related statistically. In fact, all size measurements are not only defined quantities, but the measurement actually used is an average that is controlled by the technique employed and, hopefully, reflects the property in question of the material of interest. Many papers have been presented that illustrate the accuracy of the various sizing methods, present the advantages of certain methods for certain applications, and describe the principles and technical backgrounds of the various methods. No attempt will be made in the present paper to review this work in detail. The purpose of the present paper is to examine the limitations of the various sizing methods, and to demonstrate experimentally the relationships between the sizes as determined by the various methods of analysis. It is hoped that by recognition of these limitations and relationships, errors in size specification and size interpretation will be avoided. REVIEW OF SIZING METHODS In specifying the size of a material, not only the data but also the method of measurement must be stated. For example, the size of the lump might be specified as its total volume, as its weight, as its maximum dimension, or as its minimum cross section; any one definition might be completely satisfactory for one problem, but not for another. In practice, size analyses follow four basic approaches and fortunately, the conditions that restrict each approach are known. These major methods include: 1) screen analysis, 2) direct measurement of particle dimensions, 3) determination of equivalent spherical size in response to fluid flow, and 4) specific surface determination. Each of these techniques has several variations and some new techniques are also available, but the limitations of the basic techniques are sufficient to permit at least a preliminary

Jan 1, 1962

By O. A. E. Jackson

Pebble milling has been practiced in the reduction works of South Africa gold mines for well over 50 years. Originally flint pebbles were imported from Denmark to grind stamp-mill amalgamation-process tailing, which contained a good deal of extractable gold, but local operators soon found that large pieces of ore could be used for the same purpose. The ore is a hard, tough conglomerate in which quartz pebbles are cemented together by a matrix of redeposited silica interspersed with pyrite crystals. The gold, rarely visible, occurs as fine particles mostly segregated at the interface of the pebble and matrix, although a small fraction occurs within the pyrite crystals. There is seldom any gold in the pebbles themselves. Following the usual South African practice in pebble milling, Union Corp. grinds the ore wet in two or three milling stages incorporating classification. The sized broken ore used as grinding media is separated from the main ore stream in the crushing section that prepares the ore for milling. Where the ore channel, or reef, is narrow there is a shortage of large pebbles. In this case primary grinding may be done in ball mills or, more recently, in rod mills, which cost less per ton to operate. The trend, however, is to prepare finer feed for the milling section. This makes it possible to use smaller primary pebbles and eliminates the need for steel. REDUCTION WORKS OF UNION CORP. LTD. Union Corp. Ltd. exercises financial and technical control over a group of seven gold mines in the Transvaal and Orange Free State. In the Transvaal, with one exception. the mines lie 20 to 40 miles east of Johannesburg, in flat or gently rolling countryside. Winkelhaak, the first of several new mines that will be developed by Union Corp., is located in similar terrain in an entirely new gold mining district about 80 miles east of Johannesburg. Table I gives details of milling units for six of the Union Corp. mines, together with the tonnage milled in 1957. Winkelhaak Mines Ltd. is not included, as it did not begin milling until 1958. This reduction plant has no crushing section; ore is ground directly from the mine (autogenously) in 12x16-ft mills. Because these operations are still in development, they are not described in this article. It will be noted that certain reduction works have mills of more than one size in the same milling stage. This came about when plant extensions in- corporated larger units. In the case of Geduld Propty. Mines Ltd., which began milling operations 50 years ago, the primary stage is stamp milling. The reduction works follow a uniform pattern and are usually joined to the main ore shaft. Ore from other shafts is brought by standard-gage railway and dumped into a common transfer bin. The trend is to increase surface storage capacity to enable the crushing and milling sections to operate at a steady rate, independent of fluctuating ore deliveries from mine. Milling and cyanide extraction divisions of new mines are always designed to allow for extensions as mine production increases. The conveying, washing, and screening system of the crushing section is usually laid out in final form, with additional space for more crushing equipment. The crushing sections operate on one shift during early years of mine production; a second shift is introduced when the mining rate warrants it. Ample surge capacity is provided. Crushing and milling is done only on weekdays, as the law does not allow these operations to take place on Sunday in any plants constructed since 1911. The cyanide extraction sections, however, operate continuously seven days a week, drawing on mill pulp gradually built up in the thickeners during the week. Construction and equipment of milling plants follow standard practice. Dilution water is drawn from a large, high-level tank to obtain constant pressure, but gland service water for the pulp pumps is reticulated from high-pressure, two-stage pumps. The mills are equipped with the most up-to-date machinery and are designed to save labor. They compare favorably with milling plants in countries where native labor does not exist, and automatic controls are being installed wherever feasible. Hydrocyclone classifiers have replaced mechanical classifiers in modern milling plants, chiefly because of the saving in capital outlay, maintenance, and building space. The hydroclones are fed from steady head boxes rather than directly from pumps, and dilution water is introduced into these boxes. Tests have shown that in steadiness of operation and separating efficiency cyclones are comparable to mechanical classifiers. but protective stationarv screens are needed to keep the spigots clear. Rubber-lined pumps are used for pulp of about 3 mesh or finer and metal-lined pumps for coarser material. None of the Union Corp. milling plants practices gravity concentration of coarse gold by amalgamation or the use of corduroy blankets. Studies have proved that no economic case can be made for these methods, which complicate the milling process and demand extra precautions against theft.

Jan 1, 1960

By S. G. Dixit, A. K. Biswas

The flotation and adsorption behavior of ilmenite, monazite and zircon have been studied with sodium oleate as the collector and pH as the principal variable. The maxima in the flotation and adsorption properties of any of the above minerals fall within similar pH ranges. The bell shaped pH versus adsorption magnitude (r) curve is explained on the basis of collector concentration (Co and the concentration of hydroxyl ion (COH-) acting as the depressant. An equation of the type: 1 =KiCor~K2 Coh-/Coi-is proposed to account for the experimental data. The beach sand deposits along the sea coast in Southern India contain ilmenite, monazite, zircon, rutile and sillimanite as valuable minerals while quartz, garnet, shells and traces of hornblende form the gangue. A typical composition of the deposits is reported to be as follows:' Ilmenite 75%, Zircon 4-6%, Rutile 3-4%, Monazite 2 - 4%, Garnet 2 - 5% and the rest 8 -10%. Selective flotation of beach sand constituents has been reported by many workers.2-4 The fatty acid soaps have been invariably used as collectors. It is seen that all the valuable constituents of the beach sand are efficiently floated by soap collectors such as sodium oleate. In order to selectively float the beach sand constituents, it is necessary to use some depressants such as sodium silicate.3 The physicochemical parameters that govern the flotation of these minerals are yet to be studied in detail. In the present article, the findings on the parallel studies on the flotation and adsorption of the systems — ilmenite-sodium oleate, monazite-sodium oleate and zircon-sodium oleate — are being reported. In these correlative studies on adsorption and flotation of beach sand minerals, ilmenite (FeTiO3), monazite (Ce, Dy, Nd) PO4 and zircon (ZrSiO,) have been selected because they represent three important

Jan 1, 1970

By R. B. Booth, W. F. Linke

The continuous interest of the American Cyanamid Company in producing superior polymeric flocculants and dispersants for the mining industry has resulted in a broad, general study of the physical chemistry of mineral-polymer systems. The aim of this program has been to establish a sound theoretical basis on which practical problems could be analyzed, and which would lead to the development of new and improved flocculating and dispersing agents. This study has required the cooperative effort of physical, analytical, and polymer chemists, as well as engineers familiar with mining problems. This report summarizes the results of some of our investigations over the past two years. Further studies along these and other lines are in progress and will be reported in the near future. MEASURING FLOCCULATION IN THE LABORATORY The foremost problem in studying flocculation in the laboratory has been to establish reproducible methods of comparing and evaluating the effectiveness of various reagents. Previous workers have measured settling -rates 1,8 filtration rates2 , and sediment volumes3 . Direct observations of the floc size, or the clarity of the supernatant liquid, can also be used. AIthough these parameters are all interrelated, they do not always measure the same physical properties of the suspensions, and hence are not equally useful or reliable. Whatever the method of comparison selected. the testing procedure must be carefully defined and standardized. The results must be satisfactorily reproducible, and for practical purposes must accurately reflect the results obtained in mill use. In order to compare the data obtained in these various types of test, the physical nature of the flocculated system must be examined in some detail. In an unflocculated system, only the larger or denser particles (the "sands") settle at a useful rate. The sediment can be observed to build up from the bottom of the cylinder, while the fines remain in suspension for a long period of time. If the system is not disturbed, even the microscopic fines can be seen to classify themselves as they fall over a period of days. When a high molecular weight polymer is added to the system, the colloidal particles are tied together randomly in flocs, and the sands and fines settle together. The line of demarkation between solid and liquid phases moves downward from the top of the cylinder. The settling rate of the flocs depends on their size and on their effective density. For a given mineral sample, the larger the flocs, the faster the settling. Since the large flocs cannot pack as closely as small ones, they produce a sediment with more pore spaces, and hence of larger volume, which is easier to pump and to agitate. Although it is nearly impossible to measure individual floc sizes, it is relatively easy to determine settling rates and sediment volumes. Of these two, the settling rate is simpIer to measure, and hence generally more practical. Reproducible sediment volumes can only be obtained by allowing the solids to pack for a rather long period of time (e.g., hours to days), and although the volumes thus observed are directly related to the degree of flocculation (average floc size), it is often impractical to wait so long for the results.

Jan 1, 1961

By V. Lessels, D. J. Buckwalter, D. H. Rose

A cyanide leaching process developed at Michigan Technological University was tested in a 12 to 24 tpd pilot plant. Flowsheet layout and equipment are described. Cyclic operation of leaching, copper precipitation by acidification, and regeneration of the cyanide leaching solution are discussed. Copper extraction and reagent consumption were as predicted by laboratory tests, and the process was proven to be technically feasible. The White Pine Copper Co., a wholly-owned subsidiary of the Copper Range Co., produces copper in the Upper Peninsula of Michigan. Production in 1965 amounted to 138 million pounds of metal. An expansion program now under way will increase this by approximately one half. The highly mechanized room and pillar mine produces 18,000 tons of ore per day, which is treated in a flotation plant. The concentrate is smelted in a reverberatory furnace, the matte is blown in converters to blister copper, and the blister copper is further refined to a uniformly high-quality Lake copper, which is 100% conductive and has a silver content of 25 to 40 oz per ton. The run of the mine ore averages 22.5 lbs of copper per ton with the ore mineral occurring in the form of chalcocite. The nature of the mineralization is such that the sand tailings average a little over 4 lbs of copper per ton. Intensive research has improved copper recovery in the flotation operation to some extent, but at the scale of present operation about 12,000 tons of sands containing over 48,000 lbs of copper go to the tailings pond daily. When the plant expansion is completed, the loss of copper will be somewhat greater. In 1959, the White Pine Copper CO, initiated a study of the recovery of copper in the sand tailings. Inasmuch as the flotation operation was considered to be optimized, a logical approach seemed to be leaching. For various reasons ammonia, ferric chloride, acid, and biological leaching were ruled out as impracticable. This led to a consideration of cyanide leach-

Jan 1, 1968

By S. F. Ravitz, S. R. Zimmerley

High-grade cobalt concentrates were produced from the complex Blackbird ore with very good recovery in continuous pilot-plant operations in which a low-grade bulk cobaltite-pyrite flotation concentrate was roasted in a multiple-hearth furnace to oxidize the pyrite selectively. The cobaltite was floated from the resulting calcine. THE existence of the cobalt-copper deposits in the Blackbird district, Lemhi County, Idaho, has been known for many years. However, these deposits have not yet been exploited successfully because of the low grade of the ore, absence of information on ore reserves, and lack of feasible methods for concentrating and recovering cobalt from the ore. In 1941, when it appeared that World War II might interfere seriously with imports of cobalt from overseas, the U. S. Bureau of Mines started an extensive mining and metallurgical investigation of the Blackbird deposits. Exploration by trenching and drilling1 indicated that ore reserves were adequate to warrant consideration of large-scale exploitation. A process for producing high-grade cobalt concentrates from the ore was developed at the Salt Lake City Station of the Bureau,2 and hydrometallurgical procedures for producing cobalt metal or oxide from the concentrates were developed at the Boulder City Station.3 The Howe Sound Co., through a subsidiary, the Calera Mining Co., carried out further exploratory work, which increased probable ore reserves to a tonnage commensurate with a large investment in mine, mill, transportation facilities, and refinery. Simultaneously, the metallurgical staff of the company, working in cooperation with the Bureau of Mines, further investigated the metallurgical problems. It was the opinion of both groups that pilot-plant testing was required to translate laboratory results into plant operation. In general, the Blackbird ore has a high content of pyrite and some pyrrhotite and contains 0.5 to 1 pet cobalt and 1 to 2 pet copper, which are present essentially as cobaltite and chalcopyrite. The copper can be recovered readily as a high-grade concentrate by selective notation. The cobalt then can be recovered as a bulk sulphide concentrate assaying 4 to 5 pct cobalt, but no satisfactory procedure for separating the cobalt from the pyrite and pyrrhotite directly by flotation had been found. The method developed by the Bureau for producing a high-grade cobalt concentrate is essentially an application of the old Horwood process of selective roasting.4 The bulk sulphide concentrate is roasted to form an oxide coating on the iron mineral particles, and the calcine is then treated by flotation. The oxide coating prevents the iron minerals from floating so that it is possible to obtain a high recovery of cobalt in concentrates assaying more than 20 pct cobalt. Laboratory tests by metallurgists of both the Bureau and the Howe Sound Co. had indicated the optimum roasting temperature to be 425° to 450°C. At lower temperatures the rate of oxidation of the iron minerals is too low, whereas at higher temperatures substantial quantities of cobalt are rendered water-soluble. Because of its high sulphur content (35 to 40 pct), the bulk concentrate tends to continue to burn after it has reached the optimum roasting temperature. Although excessive temperatures could be avoided in the small-scale laboratory furnaces, there was considerable doubt that the temperature could be controlled properly in industrial equipment. To provide material for the pilot-plant tests, the Calera Mining Co. installed a small, temporary flotation mill at the Blackbird property and treated

Jan 1, 1951

By S. R. Zimmerley, S. F. Ravitz

High-grade cobalt concentrates were produced from the complex Blackbird ore with very good recovery in continuous pilot-plant operations in which a low-grade bulk cobaltite-pyrite flotation concentrate was roasted in a multiple-hearth furnace to oxidize the pyrite selectively. The cobaltite was floated from the resulting calcine. THE existence of the cobalt-copper deposits in the Blackbird district, Lemhi County, Idaho, has been known for many years. However, these deposits have not yet been exploited successfully because of the low grade of the ore, absence of information on ore reserves, and lack of feasible methods for concentrating and recovering cobalt from the ore. In 1941, when it appeared that World War II might interfere seriously with imports of cobalt from overseas, the U. S. Bureau of Mines started an extensive mining and metallurgical investigation of the Blackbird deposits. Exploration by trenching and drilling1 indicated that ore reserves were adequate to warrant consideration of large-scale exploitation. A process for producing high-grade cobalt concentrates from the ore was developed at the Salt Lake City Station of the Bureau,2 and hydrometallurgical procedures for producing cobalt metal or oxide from the concentrates were developed at the Boulder City Station.3 The Howe Sound Co., through a subsidiary, the Calera Mining Co., carried out further exploratory work, which increased probable ore reserves to a tonnage commensurate with a large investment in mine, mill, transportation facilities, and refinery. Simultaneously, the metallurgical staff of the company, working in cooperation with the Bureau of Mines, further investigated the metallurgical problems. It was the opinion of both groups that pilot-plant testing was required to translate laboratory results into plant operation. In general, the Blackbird ore has a high content of pyrite and some pyrrhotite and contains 0.5 to 1 pet cobalt and 1 to 2 pet copper, which are present essentially as cobaltite and chalcopyrite. The copper can be recovered readily as a high-grade concentrate by selective notation. The cobalt then can be recovered as a bulk sulphide concentrate assaying 4 to 5 pct cobalt, but no satisfactory procedure for separating the cobalt from the pyrite and pyrrhotite directly by flotation had been found. The method developed by the Bureau for producing a high-grade cobalt concentrate is essentially an application of the old Horwood process of selective roasting.4 The bulk sulphide concentrate is roasted to form an oxide coating on the iron mineral particles, and the calcine is then treated by flotation. The oxide coating prevents the iron minerals from floating so that it is possible to obtain a high recovery of cobalt in concentrates assaying more than 20 pct cobalt. Laboratory tests by metallurgists of both the Bureau and the Howe Sound Co. had indicated the optimum roasting temperature to be 425° to 450°C. At lower temperatures the rate of oxidation of the iron minerals is too low, whereas at higher temperatures substantial quantities of cobalt are rendered water-soluble. Because of its high sulphur content (35 to 40 pct), the bulk concentrate tends to continue to burn after it has reached the optimum roasting temperature. Although excessive temperatures could be avoided in the small-scale laboratory furnaces, there was considerable doubt that the temperature could be controlled properly in industrial equipment. To provide material for the pilot-plant tests, the Calera Mining Co. installed a small, temporary flotation mill at the Blackbird property and treated

Jan 1, 1951

By M. F. McCarty, R. S. Olson

New organic flocculants of high-molecular weight are replacing lime, glue, and starch in liquid-solid separation. Of these synthetic materials, poly-acrylamides are outstanding. Polyacrylamides form a class of compounds varying widely in composition and activity. Separan 2610, one of the first of these flocculants' released to the mining industry in 1954, is described' as a high-molecular weight polymer of acrylamide with or without a copolymer material and with some replacement of amide by carboxylic groups. MECHANISM OF FLOCCULATION In practice, flocculation makes solids settle faster and filter better. But two opposing factors control it. The random motion that causes particles to collide favors Van der Waals' forces, which hold the particles together. The effect of zeta potential opposes this contact, and particles are mutually repelled. Flocculation can be achieved by reducing the magnitude of the zeta potential; this is the mechanism of flocculation by lime addition. The way in which polyacrylamides accomplish flocculation is not fully understood, but the action is undoubtedly comp1ex and probably varies considerably depending on the system. To begin with, the active groups of these compounds have an extremely high affinity for solid surfaces, probably by hydrogen bonding." Initially some active groups of the molecule attach themselves to the solids with a large portion of the molecule extending into the solution. Two particles can become tied together by long flocculant molecules in the following ways—by bonding between the active groups of two flocculant molecules on different particles, or by an extended molecule from one particle attaching itself directly to the surface of the other particles." These flocculant bridges prevent the particles from moving apart through random motion, but when such motion moves them together other active groups on the flocculant bridge bond to the solids, holding them in the closer position. This process is repeated until the particles are close together. Although bridging is probably a major flocculation mechanism, it is not the only phenomenon involved. Relatively neutral acrylamide flocculants reduce the magnitude of the zeta potential by increasing the viscosity of solution in the immediate vicinity of the solid surface which contains the mobile counter ions of the double layer. In other words, it moves the shear surface which bounds the zeta potential out into the solution containing counter ions which tend to neutralize the zeta potential. This effect is probably more pronounced during the initial adsorption period, before the flocculant molecules are pulled close to the solid. In some cases the slight charge of the polyacry-lamide molecule tends to neutralize a zeta potential. However, this is not a primary flocculation mechanism, as evidenced by the fact that the polyacrylamides will flocculate solids with either positive or negative zeta potential. Adsorption of polyacrylamide flocculants by a solid is substantially an irreversible reaction. Since the flocculant molecule is attached to the particle surface by many active groups of high bonding strength, desorption would require the simultaneous breaking of many bonds—an unlikely occurrence. UNIQUE PROPERTIES The irreversible adsorption of certain polyacrylamides, the flocculant's high affinity for solid surfaces and its rapid adsorption make it effective, but these unique properties require special techniques to achieve maximum flocculation. If flocculants of this type are added to a pulp as a relatively concentrated solution, the solids in the

Jan 1, 1960

By N. Arbiter, J. Steininger, C. C. Harris

Power consumption and operating variables in air-liquid systems for a wide range of operating conditions in the turbulent region have been correlated by two dimensionless groups involving the power ratio and the flow number. The form of the correlation was developed by considering flow balances related to the air-liquid composite density in the impeller zone. Data for two phase air-liquid systems, both from literature and from the present investigation, has been correlated with high statistical significance. Two phase solid-liquid and three phase gas-liquid-solid systems have also been correlated. The quantification of power requirements in the agitation of liquids is well established.1 Rushton, Costich and Everett have shown that, for mixers operating in the fully turbulent region with no vortex, the power number

Jan 1, 1964

By J. W. Snavely

THE high tension belt conveyor is introducing a new and tremendously expanded era of low cost bulk material handling. High tension belt conveyors are generally those installations involving very long centers, high lifts, or drops, in which the belts are stressed up to their maximum tension values, and further, where the belt construction provides tension capacity far beyond what is possible with conventional belt constructions. With these high tension installations, the magnitude of the forces involved demands careful refinement of accepted design practice in order to achieve optimum balance of all factors. No attempt will be made to evaluate the relative merits of belt conveyor haulage with other means of transportation. For present purposes, it is assumed this has already been done in favor of belt conveyor. Neither will any attempt be made to evaluate the various conveyor belt constructions now available or to balance the advantages of various types of mechanical equipment. It is also assumed that the basic haulage information on which the conveyor design is based is accurate and complete. A sustained maximum, uniform load on the belt at all times must be achieved through proper feed control and the use of adequate surge storage to level the peaks and valleys of any varying demand for the material being handled. General Belt Capacity Considerations The belt conveyor capacity tables published by various belting and conveyor equipment manufacturers vary to a considerable degree, and the ratings given are quite conservative. Of necessity, these published ratings are based on the handling of average materials under average conditions. In applying a high tension belt, all possible capacity from the belt must be obtained in order to hold its width to a minimum and thereby limit the initial cost. Two factors are involved, loading to maximum cross section area and traveling at a maximum practical speed. Belt Loading: Proper treatment of the loading of the belt will result in maximum cross section to the load, and published capacity ratings can be exceeded, sometimes by appreciable margins. On the 10-mile conveyor haul used in the construction of Shasta Dam, California, although the rated capacity of the belt line was 1100 tons per hr, at times the system handled peak loads of 1400 tons per hr, almost 25 pct better than the rated capacity. One of the large coal companies has been able to exceed rated capacity by as much as 50 pct. Loading conditions which must be controlled are: 1. Large lumps must be scalped off and rejected or the load must be primary crushed before being placed on the belt. 2. The material weight per cubic foot must be accurate, must be known for all the materials being handled, and must be known for the complete range of conditions of the individual material being handled. Long centers and high lifts magnify small differences into serious proportions. 3. Uniform feeding to the belt is most important. Various types of feeders are available, which can be used to place a constant predetermined volume of material on the belt, or, where an appreciable range of material weight exists, through electrical control actuated by current demand, to place a predetermined uniform tonnage on the belt. One long slope belt in a coal mine in Pennsylvania is being fed at three separate stations with the controls so arranged that whenever the maximum load is going onto the belt from the first station, the other two stations automatically cut out. Whenever the load from the first station drops back, the other two stations again automatically cut in. 4. Careful design of the chutes and skirts is necessary to get the load centered on the belt with a minimum of free margin along each edge. Some free margin at the edge of the belt is necessary to prevent spillage, but if the load can be kept accurately centered, this free margin area can be reduced, and more material can be carried on the belt. What can be accomplished in this respect will vary widely, depending on the nature of the material being hauled. The chute and skirt design must also protect the belt. 5. The design of chutes and skirts should also get the load traveling in the same direction and close to belt speed, so that the load comes to rest on the belt as quickly as possible. The design of the chutes and skirts is worthy of careful study, and after a system is put into operation it should be experimented with to get the best results. Belt Speed: High belt speeds should be used in high tension work. Obviously, high belt speeds enable haulage on a narrower belt, reducing initial cost. The major portion of belt wear takes place at the loading point and around the terminal pulleys. The

Jan 1, 1952

By F. A. Schaufelberger

METAL can be recovered from a leach solution either indirectly by precipitation as a compound that is later reduced or directly by electrolysis, cementation, or chemical reduction, for example, with hydrogen. The widely used electrowinning processes suffer from the inefficiency of converting fuel to low voltage dc and from low current efficiency when complete recovery is attempted. With the batch hydrogen reduction methods outlined here, a unit of fuel converted to hydrogen in modern gas reform plants will make two to three times as much metal as can be obtained with electrolysis. With continuous hydrogen reduction, four to six times as much metal is obtained. Precipitation by reducing a metal salt solution with H, has been known for almost 100 years.' Commercial use of the process, however, awaited the research and development program initiated by Chemical Construction Corp. Within the last few years this program has brought about construction of commercial plants, listed on page 545. Ideal hydrogen reduction will precipitate a pure metal from a solution obtained by commercial leaching methods at a rapid rate without excessive temperatures and pressures. The metal precipitate will be of desired size and density, less than a percent left in the vessel being deposited on the wetted parts. The present discussion will outline Chemical Construction Corp.'s early development program and will discuss the chemistry and mechanics of reducing copper, nickel, cobalt, and cadmium from solution by H2. Work on selective reduction of nickel from cobalt has been described earlier.' The article is not concerned with precipitation of copper by disproportionation of cuprous solutions where yield is limited3 to 50 pet.* It does not discuss use of gaseous reducing agents other than hydrogen, such as SO2,4 which may lead to contamination with sulfur, or C0,5 which is considerably more expensive than hydrogen and produces a gaseous reaction product, CO2.† The article does not include use of other nongaseous reducing agents,7 which are far more expensive than CO. Also, notes on the history, engineering design, and performance of commercial plants have been given elsewhere.' It will be shown that a pure metal can be precipitated by reduction with hydrogen from solutions obtained by commercial leaching methods when the solution composition is controlled and proper acidity, proper metal ion concentration through controlled complex formation and hydrolysis, and adequate agitation for suspension of metal and for transfer of reducing gas are maintained. Reasonable temperatures and pressures are used which are less excessive than those used by previous workers, although they are maintained at above equilibrium values because of process kinetics.' It has also been found necessary to induce nuclea-tion and to control growth and agglomeration in order to get a powder product of suitable physical properties. Review of the Literature Muller, Schlecht, and Schubardt of I. G. Farben10 claimed a successive reduction of silver, copper, nickel, cobalt, and zinc from ammoniacal solution by applying progressively higher temperatures and hydrogen pressures. Metals produced by this technique were not pure, since no attempt was made to adjust the solution composition between the different reductions. It is doubtful that the zinc product contained any metallic zinc. The major contribution to the literature" was made by the Ipatiews,12 who prepared platinum, iridium, copper, nickel, cobalt, lead, tin, arsenic, antimony, and bismuth at often extreme conditions, up to the critical temperature of water (373ºC), at 1500 to 7500 psi, for as long as several days. Even cadmium and zinc were claimed to have been observed in trace quantities precipitated together with basic salts. In the early experiments solutions were not stirred. Extensive formation of oxides and basic salts occurred, probably long before sufficient hydrogen could be supplied in the unstirred liquid to decrease the metal concentration in solution by reduction. The importance of hydrogen pressure was

Jan 1, 1957

By B. C. Mariacher

EXTRACTION of uranium from ores is being ac-complished by processes which. for the most part, subject the entire ore to acid or carbonate leaching. Ore deposits with a U 3 O 8 content below 0.10 pct U 3 O 8 are seldom considered suitable for treatment by leaching. A preliminary concentration that would enrich the uranium content of an ore by a simple, low cost process based on physical properties of the ore might result in some low grade deposits becoming commercial ores. In addition, the process might be employed in existing operations to reduce transportation and leaching costs and to increase capacity of existing leaching plants. A study to attempt the development of a preliminary concentration process for primary uranium ores was undertaken by the Colorado School of Mines Research Foundation under sponsorship of the U.S. Atomic Energy Commission. The objective of this work was to produce concentrates containing 0.25 pct U3O8 from the low grade ores tested. Ores Tested: The main effort was devoted to the low grade primary uranium ores from northwestern Saskatchewan. Samples were obtained from the Beaverlodge operation of the Eldorado Mining & Refining Ltd. Additional primary ores, obtained from deposits in Gilpin County, Colo., contained from 0.07 to 0.10 pct U3O8. Summary of Concentration Tests: The Beaverlodge ore was tested to determine amenability of the ore to concentration by magnetic, electrostatic, gravity, and scrubbing processes. None of these produced satisfactory results. Both gravity and magnetic processes produced fairly good concentrates when closely sized fractions of the ore were treated, but on the basis of treating the total ore, recovery was poor. Preparation of sized fractions and the low capacity of equipment for suitable concentration made these methods impractical. As flotation offered the advantage of treating the total ore without intermediate sizing, the main effort was in this direction. A flotation process was developed that fulfilled the concentration objectives as set by the AEC. Pilot plant testing was used to verify results obtained from laboratory batch testing. Mineralogy: A petrographic examination of the Beaverlodge ore included a study of polished sur- faces and identification of the radioactive mineral by autoradiograph and X-ray diffraction. Approximate quantitative mineral identification was as follows: quartz, 60 pct; orthoclase feldspar, 20 pct; chlorite, 10 pct; carbonates, 5 pct; and miscellaneous minerals, 5 pct. Included in this last group were plagioclase feldspar, pyrite, mica, chalcopyrite, pyroxene, sericite, magnetite, galena, and uraninite. The most general occurrence of uraninite was in the form of crusts and thin coatings on limonite-stained grains of pyrite, quartz, and pyrite-quartz intergrowth. At least 90 pct of the uraninite was still attached to other minerals in a 100 by 200-mesh size fraction. The uraninite crusts were as small as 10 to 20 µ diam, and 5 to 10 µ thick. The Flotation Process Petrographic examinations of the Beaverlodge ore had indicated the impracticability of attempting to concentrate the uranium by floating individual grains of uraninite. Liberation of the uraninite required grinding to sizes below those suitable for flotation. However, there was preferential association of the uraninite with some minerals while others were free of uraninite attachment. The approach to the development of a flotation process was, therefore, based upon an attempt to concentrate the uraninite by floating carrier minerals. The following paragraphs discuss the various stages of the process with regard to the factors tested and the conditions under which best results were obtained. Grinding: The most effective size range for flotation was —150 mesh + 13 µ. The —13 µ material in the final concentrate had a higher U3O8 content than the total ore, but not as high as the average concentrate; however, rejection of slimes before flotation was prohibitive because of the loss in uranium carried in the —13 µ fraction. Grinding techniques which contributed to a minimum production of fines, such as stage grinding, were then employed. Quartz and Silicate Depression: These minerals represented approximately 80 pct of the ore and were free to a large degree of uraninite attachment. Significant improvement in the grade of the concentrate was obtained by depression of these minerals with hydrofluoric acid or sodium fluoride. Promoter: Selective stage flotation of uraninite carrier minerals was simplified by development of a single promoter mixture. The mixture consisted of an emulsion of a fatty acid, fuel oil, and a petroleum sulfonate and was selected after a comprehensive series of tests. It contained three parts by weight of an oleic and linoleic acid such as Emersol 300,

Jan 1, 1957

By E. J. Roberts, E. B. Fitch

THE mechanism of classification by settling pools is most simply shown in the case of batch sedimentation such as was analyzed by Oden.1 The batch model will be considered, therefore, and it will be shown that the same mechanism applies to continuous operation. Classification in Batch Settling Pools: Consider a beaker or tank of slurry containing particles so dispersed that they can settle independently of one another under the influence of gravity. Originally all particle sizes are uniformly distributed throughout the suspension. All particles start immediately to settle towards the bottom at rates determined by their size, shape, and density—among other factors. Particles reaching the bottom settle out of suspension. After some given time the slurry above the deposited solids is siphoned off or otherwise removed, leaving the deposited solids behind. The deposited solids, together with entrained and any unremoved supernatant slurry, represent the underflow of a classifier. The slurry removed represents the overflow. There is a certain separation, size of particle which during the sedimentation has just time to settle from the top surface of the slurry to the level at which underflow and overflow are parted. All particles with a faster settling rate will have dropped to below this level, no matter where in the beaker they started from originally, and therefore will be absent from the overflow. Particles with a slower settling rate will not reach this level if they started from the top of the slurry, but a portion of them,

Jan 1, 1957

By B. Ball, D. W. Fuerstenau, P. C. Kapur

A two-phase distributed-parameter model of the flotation process has been extended to allow consideration of the important process parameter, concentrate grade. A simple expression is presented for the grade of concentrate in terms of the zeroth and first moments of the concentrate distribution. An example is given for a binary ore containing two pure minerals each having nondistributed rate constants. For extend-ing the model to the distributed rate constant case, a cubic polynomial zoas utilized to simulate the feed distribution. One particular example is considered and grade-recovery curves are generated to show the dependence of such curves on the initial feed distribution. Operating flotation plant engineers have traditionally evaluated plant performance in terms of the recovery of valuable mineral and the grade of the concentrate. In both operation and research, grade-recovery relationships have been used as a guide for selecting flotation conditions. Weissl illustrated the use of grade-recovery curves in plant operation and considered the grade-recovery relationship to be a metallurgical one. The extensive experimental work of Chi" using standard batch flotation tests on iron ores shows that, even with extensive statistical correlation of data, little insight can be gained into the basic fundamentals of these relations. Such relations must depend on the kinetic behavior of the various mineral species in the flotation circuit. In his flotation text book, Gaudin3 considered the relationship of recovery and of product grade to flotation kinetic indices, but he did not extend his calculations to yield grade-recovery relations. Although Gaudin used a simplistic model for his analysis, he did recognize that variations in flotation rates must result from variations in locking, surface heterogeneity, particle size and shape, among other things. In a recent paper,' we presented a model of the flotation process that could be readily analyzed with an analog computer. This model combines the concept of a distributed set of rate constants"" with that of the two-phase flotation model,8" resulting in a model more powerful and flexible than previous ones. The purpose of the present paper is to show how this model can be used to generate grade-recovery curves and to determine the dependence of such curves on the various kinetic parameters. In this paper, the model will be limited to the semi-batch case, but extension to the continuous case is a straightforward step.

Jan 1, 1971

By P. Somasundaran, D. W. Fuerstenau

This article presents the results of an analysis of grinding of 1:1 mixtures of 4x8-mesh quartz and limestone in laboratory ball and rod mills. From these experiments, the amount of energy consumed in grinding each component of the mixture was determined. In the ball mill, limestone consumes slightly more energy than quartz in the initial stages of grinding, but as the grinding proceeds quartz begins to consume a slightly greater fraction of the energy. The postulate that the fraction of energy consumed by any material in a ball mill is proportional to its volume fraction has been shown to be approximately true. On the other hand, in a rod mill energy consumption by quartz and limestone is equal only in the initial stages of grinding, with energy consumption by quartz greatly exceeding that by limestone as rod milling proceeds. Preferential energy consumption in tumbling mills has been discussed in a number of papers. For example, from an analysis of the kinetics of comminu- tion in ball mills, Arbiter and Bhrany' obtained rate equations based on the postulate that only coarse particles are ground, indicating that fine particles are protected from the action of the balls by coarser particles. In rod mills this protective action is much more pronounced than in ball mills. Recent studies on the rod milling of limestone-quartz mixtures have shown that this protective action causes the mineral with the lowest grindability (quartz in this instance) to consume the greater fraction of energy. 2 However, in the ball milling of mixtures it has been postulated that the consumption of energy by one of the constituents is determined by its volume fraction within the mill. 3 Possibly this hypothesis may be a simplification of the ball milling of mixtures since the protective action suggested by Arbiter and Bhrany may result in some preferential energy consumption by one of the constituents of a mixture. The objective of this study has been to determine preferential energy consumption in laboratory ball and rod mills by analyzing size distributions of each material in the products obtained from grinding 1:l mixtures of quartz and limestone. MATERIALS AND METHOD The procedure for carrying out the tests on wet grinding quartz-limestone mixtures has already been described.'r3 In the present investigation, 1:l mix-

Jan 1, 1963

By W. J. Tait, A. E. Craig, E. P. McCurdy

The Circuitron herein described applies current from the classifier motor circuit and energy from the sound of grinding media to move an oscillating disc. The disc through a photoelectric cell controls the speed of the ore feeder and consequently the amount of ore delivered to the mill. THE Telluride Mines, Inc. at Telluride, Colo., is operating a flotation concentrator recovering lead, zinc, and copper with some gold and silver. The nominal capacity is 500 tons per 24 hr. The ore, which comes from several different mines, is characteristic of the district, changing from extremely tough rhodonite to soft gouge. The ore is crushed and washed, recrushed, and delivered to the mill bin at about % in. to serve as ball mill feed. Details of Grinding Circuit The material from the fine-ore bin is carried over a belt feeder driven by a Reeves vari-speed device. The belt discharges to a No. 86 Marcy ball mill, which operates in closed circuit, with jigs and a 6-ft Dorr classifier. A Massco density controller holds the density of the classifier overflow at a predetermined point. The overflow, which contains about 5 pct + 48 mesh, goes to the flotation section. Manual Operation Before the adoption of automatic control for the grinding circuit, this work was all done manually by the operator. The bin gate was set to deliver 40 to 50 lb of ore per lineal foot of belt, and the speed of the belt was adjusted by the operator to deliver at the rate of about 500 tons per 24 hr. As the ore changed in hardness, specific gravity, or in the amount of moisture, or, as the size of the feed changed due to bin segregation, crusher setting, or other causes, the operator made corresponding changes in the rate of feed by manipulation of the Reeves control. Sometimes these changes occurred so rapidly that they caused sharp, pronounced surges, which created very considerable unbalance from normal conditions in the grinding circuit and, more critically, in the flotation section. Since it is not possible for the operator to anticipate or follow such changes closely, a very considerable unbalance from normal conditions may occur, which will require a drastic change in the rate of feed. These surges upset the fineness of grind and interfere radically with the balance between tonnage and reagent control. All of this tends to reduce the grade of concentrates, increase the loss in the tailing, and thus decrease the recovery of values. The Circuitron Some years ago, The Mine and Smelter Supply Co., in conjunction with the operators at the Telluride Mill, undertook to build a device that would give close control of the rate of feed from the ore bin and to coordinate with this a related device that would regulate the amount of sand load on the classifier rakes. Emphasis was strongly laid on the fact that such a control must apply to the entire grinding circuit and must include the feed to the ball mill, the classifier sand return, and the density of the classifier overflow. If this could be done successfully, we felt that the circuit could be operated under best possible conditions, with compensations

Jan 1, 1951