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By Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951

By Jack White

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the l l-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the ll-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 1% in. and a further gain of work accomplished of 2.1 pct when we went from 1 % in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By Jack White

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the l l-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the ll-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 1% in. and a further gain of work accomplished of 2.1 pct when we went from 1 % in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

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 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 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 E. B. Johnson

Cleveland-Cliffs lron Co., which operates Marquette lron Mining Co.'s Republic mine, conducted a test whereby 35,000 tons of self-fluxed pellets were produced in the grate-kiln system from specular hematite flotation concentrates. A brief description of the pelletizing plant is presented along with pertinent operating details. Date is included which compare operating conditions and the chemical and physical properties of the green and fired pellets for the self-fluxing pelletizing test with the normal acid pelletizing operation. The success of self-fluxed sinter in enhancing the operation of the blast furnace during the past few years has reasonably led to the assumption that the use of self-fluxed pellets might result in even greater furnace efficiencies. A great share of development work in agglomeration has been directed towards techniques related to the production of self-fluxed pellets, but only a limited tonnage of self-fluxed pellets had been produced on a commercial scale until this past year when The Cleveland-Cliffs Iron Co. produced 35,000 tons of such pellets by the grate kiln system at the Republic mine during the period May 21-June 5, 1962. REPUBLIC'S GRATE-KILN PLANT The grate-kiln pelletizing process consists of four stages, namely: feed or balling circuit, traveling grate, rotary kiln and annular cooler (Fig. 1). The balling section is conventional and utilizes four 9x30-ft balling drums in closed circuit with 5x14-ft vibrating screens having 3/8-in. rod decks. The traveling grate is 12 ft wide by 126 ft long and has 14 windboxes — six in the drying section and eight in the preheat section. The rotary kiln is 15 ft ID shell and 114 ft long. The annular cooler is 40 ft in diameter and has 12 windboxes. The cooler is divided into two zones by a moveable curtain wall. The high temperature air from the first zone is used as secondary combustion air in the kiln. Normally the only heat supplied is through the burning of No. 6 oil in the single kiln burner. Hot gases pass counter to the pellet flow through the system. The exit gases from the kiln pass down draft through the preheat section through cyclones and are drawn through the drying bed prior to exiting via the stack. The temperatures are normally held within the following ranges: Kiln 2400-2445 °F Preheat 1850-1950°F Drying 480-520°F Exit gas 200°F Pellets 120°F The design capacity of the Republic mine grate-kiln system was 2500 Itpd. Sustained production in excess of 35% over design capacity has been achieved. No screening of the cooler product is practiced. PELLETIZING TEST CONDITIONS When a decision was made to proceed with a test at the Republic mine pelletizing plant to produce 35,000 tons of self-fluxed pellets, several potential failure factors had to be carefully considered. Research work on a batch and pilot plant scale indicated that the grate-kiln system at Republic would require some major changes in order to obtain optimum conditions. In many cases compromises were agreed upon as major equipment changes and revisions were not feasible. The pelletizing conditions which laboratory development tests revealed were required in order to insure successful production of self-fluxed pellets were: 1) Good mixing of limestone and concentrate and proper moisture control; 2) Higher gas flow rates in the preheat section than are normally used; 3) Higher preheat temperatures than normal; 4) Maximum kiln temperature 2350°F. There was just cause for concern about balling problems that might occur. Approximately 10% of dry pulverized limestone was to be added to the specular hematite concentrate just ahead of the Pekay mixers on the belt feeding the balling drum. Large scale

Jan 1, 1963

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 D. F. Kaufman, H. R. Spedden, A. M. Gaudin

MANY studies of comminution have been made to ascertain the size distribution of the product and to evaluate the work of comminution in the light of the size distributions of the feed and product. Up to now, these studies have been essentially statistical in character, that is, a certain lot of feed was subjected to comminution in some specified way, and the aggregate product was fractionated into sizes, thereby losing all knowledge of individual relationship of feed to product pieces. Radioactive tracers offer a means to do something in this respect which could not be done before, namely, to follow the rupturing of some particular piece in its normal environment of other pieces. That is, it permits going beyond the usual statistical limitations of size distribution studies to what may be termed a personalized or individualized study. The purpose of this paper is to present some preliminary experiments conducted with this tool. The method employed was to mark radioactively some constituent of a feed. It is possible, of course, to consider the preparation of two lots of material of which one is radioactive and the other is not, and to blend the two ahead of the comminuting step; but to do so is open to the objection that the two preparations may not be identical. Therefore a technique has been chosen that removes this objection by merely taking out a size fraction of a comminution feed, rendering that fraction radioactive by exposure to a neutron flux, and then by returning it to Table I. Size Distribution of Offspring Albite Particles Originally 28/35 Mesh and in Admixture with Other Sizes After Grinding 2 min in a Steel Ball Mill Specific Activity ' Cumu- Corrected Distrl- latlve Size for Back- butlon In Distri- Fractlon ground, Weight, Product, button, of Product, cpm/gm g Pctb Pct Mesh (A). (W) (P) (ZP) + 28 0 56.0 0 100.1 28/35 62.6 54.0 24.8 75.3 35/48 62.8 59.4 27.7 47.6 48/65 41.1 53.0 16.2 31.4 65/100 29.6 45.7 10.2 21.2 100/150 23.7 37.0 6.6 14.6 150/200 23.3 25.1 4.4 10.2 200/270 20.1 19.0 2.9 7.3 270/400 17.8 21.2 2.9 4.4 -400 22.9 25.2 4.4 — 100.1 a These activity determinations were made in rapid succession in the order given. The specific activity (Ao) of the active 28/35 mesh fraction of the feed was measured at the beginning, after the measurement on the 65/100 mesh size fraction of the product, and; The end. The decay-corrected activities at those times were 246.7, 241.0. and 236.9 cpm per gm. The weight (W0) of the active 28/35 mesh fraction in the feed was 55.0. b Example of calculation for P in the 65/100 mesh oroduct frac- A W tion; A = 29.6, W = 45.7, Ao = 242.7, Wo = 55.0: P = — x — Ao Wo = 0.102 = 10.2 pet. the remainder of the charge for the comminution experiment. A relatively simple procedure was developed by which albite, containing sodium, was activated in the M.I.T. cyclotron. The cyclotron makes highspeed deuterons which impinge on a beryllium target, thereby producing a concentrated neutron flux. The mineral was exposed to this flux for 2 hr. This treatment changed enough of the sodium to sodium 24 (14.8 hr half-life, 1.4 mev ß) as to make detection and measurement easy. The nuclear reactions taking place were: 11Na23 (n,?) 11Na24 (irradiation) 11Na24 ß,?,? 12Mg24 (decay) The detailed technique of the experimentation was as follows: 40 kg of hand-sorted, lump albite were crushed to pass 10 mesh. After careful mixing of the lot, a screen analysis was made. The whole lot of material was fractionated on standard Tyler screens from 14 down to 200 mesh. Samples for experiments were compounded from these fractions in accordance with the screen analysis. When it was desired to make an experiment in which, for example, the 28/35 mesh size fraction was to be studied, the blend of size fractions was made as indicated above, except that the 28/35 mesh size fraction was added only after irradiation in the cyclotron. The blended charge containing the activated albite was ground for 2 min in a laboratory ball mill with a steel ball charge of controlled size distribution. The ground product was carefully sized on a set of Tyler screens in a Ro-tap. Each size was analyzed for radioactivity by the use of an end-window Geiger-Mueller counter and standard scaling circuit. This analysis was carried out in detail as follows: a 20-g sample was placed in a Petri dish, packed carefully to obtain reproducible geometric distribution with reference to the Geiger-Mueller tube, and the activity was counted for a 2-min period. Several determinations of the activity of the active size fraction in the feed were made at various times to establish the decay in activity with time. Linear interpolation was used to evaluate the activity that the active size fraction in the feed would have had at any given instant. The ratio of the observed activity in a size fraction of the product to the activity that the active size fraction in the feed would have had at the same time gives the fraction in the product size that came from the irradiated size in the feed. The general formula for finding the distribution, P, of a specific individual size fraction in the feed

Jan 1, 1952

By F. M. Lewis, J. F. Myers

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the 11-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the 11-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 11/2 in. and a further gain of work accomplished of 2.1 pct when we went from 11/2 in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By J. F. Myers, F. M. Lewis

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the 11-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the 11-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 11/2 in. and a further gain of work accomplished of 2.1 pct when we went from 11/2 in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By W. F. Sullivan, J. R. Coleman

The use of radioisotopes in obtaining basic information on large scale industrial processing is illustrated with two examples. The first is concerned with the operating characteristics of a kiln scrubbing tower and the second with those of a Hardinge ore mill. The examples show the advantages associated with the industrial use of tracers as well as the limitations. The Atomic Energy Commission regulations governing the use of radioisotopes in this manner are also discussed. The nuclear properties of radioactive isotopes have provided a new technique for obtaining basic information on large scale industrial processing. A small portion of the process stream is made radioactive by mixing with a suitable radioisotope or by irradiation in a nuclear reactor. The "tagged" portion is then part of the normal process stream and can be followed by detection and measurement of the associated radioactivity. The advantages of radioactive tracing over other forms of tracers are well known, and include such factors as chemical similarity, capability of precise determination of very small quantities due to the high sensitivity of detection apparatus, and the absence of foreign element contamination of the final product. This report will illustrate two applications of the use of radioactive tracers in large scale industrial processing which have provided information that would have been hard to obtain by other means. The first is concerned with the operating characteristics of a kiln scrubbing tower and the second with those of a Hardinge ore mill. ATOMIC ENERGY COMMISSION REGULATIONS Atomic Energy Commission approval can be obtained by qualified personnel if the proposed quantity of radioactivity does not exceed AEC's standards for

Jan 1, 1965

By J. E. Campbell, G. D. Calkins, N. M. Ewbank, M. Pobereskin, A. Wesner

GRINDING for size reduction affects the economics of many processes and products. It is essential as the first step in many industrial processes and is also a finishing step for materials with properties depending on particle size, such as talc, cement, and silica sand. Intermediate and fine grinding are vital operations in the U. S. cement industry, which is producing more than 250 million bbl of cement per year.' Wear of the grinding media is a large part of the grinding operation cost. Problems encountered in grinding cement are so complex that evaluation of efficiency and economy of grinding media is difficult.2 It has been especially difficult to evaluate the relative effectiveness of different types of balls because there are no good testing techniques. Many other industrial operations can be evaluated on a laboratory scale with reasonable accuracy. This does not hold true for evaluation of grinding balls. The consistent results obtained in a laboratory test under a given set of conditions are not always borne out in field application. Rough evaluations of the effectiveness of various compositions and types of grinding balls have been made in the field by using a full charge of one type in a mill and comparing the production record with another run using another type of ball. This method is time-consuming and not very precise, as the second run may not have been carried out under identical conditions. Laboratory-scale tests, on the other hand, have yielded inconclusive results, and many investigators have turned their attention to the development of a field testing technique. Field testing small sample lots of grinding balls has been impractical because it is difficult to identify and recover the test specimens from the grinding mill, and individual groups of balls that have undergone different heat treatments can not be separated.".4 To overcome these difficulties, previous investigators have identified the balls by distinctive marks, notches, and drilled holes, but this procedure has three serious drawbacks: 1) Grinding characteristics and quality of the steel balls may be affected. 2) Physical markings may be worn away in the grinding process, especially during a prolonged run. 3) Recovery from the bulk of the charge will be extremely difficult because the markings are hard to see and may be masked by a coating of the product. To circumvent these difficulties, a radioactive-tracer technique was proposed for recovery and separation of steel grinding balls and subsequent evaluation of the various compositions of the balls. The proposed technique involved five basic operations: 1) Thermal-neutron irradiation activation5 of each group of test grinding balls to a different level of specific radioactivity. 2) Addition of groups of radioactive steel-ball specimens into a ball tube mill. 3) Recovery of radioactive steel-ball specimens from the bulk of the mill charge. 4) Separation of the various groups by their specific radioactivity. 5) Evaluation of actual grinding ball wear. Before any physical tests were performed, required neutron irradiation intensity and time were calculated. Probable composition of the steels to be used was ascertained. An examination was made of the radioactive nuclides8 to be formed which would contribute measurably to the radiation level immediately after irradiation and during the test operation. The radioisotopes formed, their types of radiation, and their half lives are listed in Table I. Of these radioisotopes only iron-59 and chromium-51 were significant for the actual wear test. The intensity of radiation that could be detected by a Geiger counter when the test was completed was the basis for the minimum activation level established. The intensity of radiaton that could be safely handled at the beginning of the test was the basis for the maximum activation level, although this was not considered a major problem. Ten groups of grinding balls of various composition and/or surface or heat treatment were to be tested. One group was designated for the minimum irradiation time. The remaining groups were designated for irradiation periods that increased by increments of 33 pct from that of each preceding group. This difference was considered enough for separation and identification of the groups by comparison of specific activity. Potential Hazards: Possible radiation hazards that might be encountered during this experiment were evaluated for the three important phases: 1) the radiation hazard of placing balls and removing them from the mill, 2) contamination of the product cement by radioactive material worn from the balls, and 3) contamination of the steel by the radioactive balls left in the mill. The radiation intensity expected from the whole group of radioactive balls was calculated to be 250 milliroentgen per hr at 1 ft. This meant the balls would require special shielded packaging and warning labels on the shipping containers. In a radiation field of 250 mr per hr a man can work for 1 hr without exceeding maximum permissible weekly exposure. Since the balls could be dumped into the mill in a matter of seconds, relatively little radiation exposure was anticipated at this stage of the operation. If the weight loss in the balls was 7.7 pct per month and the cement feed through the mill was

Jan 1, 1958

By G. W. Sheary

At Bethlehem's Grace mine's 6000-tpd pelletizing plant a cascade control system is utilized to approach a constant magnetite filter cake moisture, a critical factor in obtaining quality green pellets. Variables being controlled by instrumentation are filter feed solids, filter tub level, first and second stage pump sump levels and thickener sludge level, all at varying feed rates. Startup difficulties, modifications and contemplated additions to obtain closer control of the moisture are discussed. Situated 45 miles west of Philadelphia and 50 miles southwest of Bethlehem Steel Co. 's Beth-lehem plant, the integrated Grace mine facilities produce 4000 tpd of high-quality iron ore pellets. Startup of the Grace mill was in November 1958, and iron concentrates were shipped until March 1960, when the pellet plant went into operation. The underground mine, when fully developed, will hoist about 9000 tpd, including development rock. The crude ore, which has an iron content of approximately 42%, is magnetically cobbed and crushed in closed circuit with %-in. slotted vibrating screens. The ore is washed and screened in the mill and the 1/4-in. screen product is again cobbed prior to feeding to the rod mills. Grinding is done in rod mills, and in ball mills that are in closed circuit with 30-in. cyclones, to 1% on 100 mesh and 65% -325 mesh. The ground ore is magnetically concentrated by three-drum permanent-type finisher magnets to 67% Fe. The magnetite concentrate is pumped from the mill to a 50-ft thickener in the pellet plant. The thickened magnetite concentrate is relayed to the filters through a two-stage pump arrangement. The filter cake is fed by belt from filter concentrate bins to mixer-muller belts where it is intimately mixed with bentonite and a slight amount of water for moisture control. The fluffed concentrate then drops into balling cones. The green pellets from the balling cones are fed in a prescribed feed pattern to shaft furnaces through the use of shuttle car mechanisms. The green pellets are dried, heated to 2500°F and cooled to 1200°F as they descend through the furnace and into pellet coolers where the pellets are further reduced in temperature to 250°F. Heat is recovered in the cooling process through the use of air-to-air heat exchangers. The cooled pellets discharge on a rubber conveyor belt and are transferred to 6000-ton pellet railroad loading bin. The final pellet product consists of 90% +'1/4 in. material (after tumble testing1) and contains 66% Fe plus 3% silica. It has been the usual practice to maintain constant filter tank level by pumping an excess constant volume and using a moveable fishtail splitter to divert the desired amount of slurry to the filter tank. The excess from the fishtail and from the filter overflow pipe, which maintains a constant level, is returned to the source. At the Grace mill, tremolite builds up in the pipelines when pumping heavy magnetite concentrate slurries, making it necessary to use polyvinyl

Jan 1, 1962

By G. L. Simard, D. J. Salley, J. Chupak

DITHIOPHOSPHATES and xanthates are the principal collectors for sulphide minerals, and consequently any knowledge of mineral-collector systems of this type is of value. In the present investigation an attempt was made to obtain information on the interaction of a typical dithiophosphate with galena. In carrying out this study, radioiso-topes, so much discussed in the past few years,', ' were extensively employed. By using radioactive dithiophosphate synthesized from radioactive phosphorus, a rapid and sensitive analytical procedure G. L. SIMARD is in the Research Division, Stamford Research Labs., American Cyanamid Co., Stamford, Conn.; J. CHUPAK, formerly in the Research Division, Stamford Research Labs., is now at Camp Detrick, Frederick, Md.; and D. J. SALLEY is in the Research Division, Stamford Research Labs. AlME New York Meeting, Feb. 1950. TP 2815 B. Discussion (2 copies) may be sent to Transactions AlME before April 30, 1950. Manuscript received Oct. 17, 1949. was at hand. This permitted determination of such important quantities as the rate of uptake of dithiophosphate by the mineral, the amount existing at equilibrium on the mineral surface and in the solution, and the desorption of the agent from the surface. Such measurements were possible even on single crystals of galena. In addition the exchange of dithiophosphate between the solution and the sorbed phase was examined, a matter which could be accomplished only by the use of isotopes. It is desired to point out at this time that throughout this paper the term "sorption" has been employed to designate the uptake of agent by mineral, without implication as to the nature of the process by which the uptake was accomplished. Experimental Materials: Galena: The galena was from the Tri-State district. For studies on ground mineral four preparations were used during the course of the investigation. These were prepared from selected large crystals by wet grinding in order to reduce surface oxidation. Alcohol was chosen as a convenient medium for this purpose. The ground mineral was then fractionated by sedimentation in alcohol, dried by evacuation, and stored in a nitrogen-filled desiccator. Data on the preparations are tabulated in table I. The size analysis indicated that the areas of the preparations were of comparable magnitude, even though the absolute values may be somewhat incorrect. Dithiophosphates: Nonradioactive dithiophbsphate (di-isopropyl or di-secondary butyl) was obtained by purification of a commercial product. An aqueous acid solution of the agent was extracted with petroleum ether, the ether layer dried, and the dithiophosphate precipitated as ammonium salt with anhydrous ammonia. Several repetitions of this process resulted in a nearly colorless, flaky product of good purity. (General formula (RO,)PSSNH,). Radioactive dithiophosphate was synthesized by heating radioactive elementary red phosphorus* • For early experiments (1943 to 1944), the elementary red phosphorus was obtained from the cyclotron group at the Crocker Radiation Laboratory of the University of California. Berkeley, through the courtesy of Dr. Joseph W. Hamilton. For more recent work (1946 to 1949). the elementary radioactive Phosphorus has been supplied by the Oak Ridge National Labarotories on allocation from the U. S. Atomic Energy Commission. with sulphur at 270" to 300°C to produce radioactive P,S,. The latter was then treated with the desired purified alcohol (isopropyl or secondary butyl) at 60" to 80°C to form the dithiophosphoric acid derivative. Purification was effected in the same manner as for the nonradioactive material. The effectiveness of the purification method was established by the isotopic dilution method." The same technique was used to show that only negligible decomposition of neutral or of carbonate solutions of the dithiophosphate took place over a period of a day; this was true whether or not galena was suspended in the solutions. Procedures: Radioactivity: The activity of a solution was obtained by counting with a small glass-jacketed, silvered Geiger counter, using a conventional scaling circuit.' Crystals and other solids were counted under a bell-shaped, mica-window counter tube. Sufficient counts were made so that the probable error of counting was of the order of +I-3 pct. The specific activity in terms of counts per minute

Jan 1, 1951

By J. Chupak, D. J. Salley, G. L. Simard

DITHIOPHOSPHATES and xanthates are the principal collectors for sulphide minerals, and consequently any knowledge of mineral-collector systems of this type is of value. In the present investigation an attempt was made to obtain information on the interaction of a typical dithiophosphate with galena. In carrying out this study, radioiso-topes, so much discussed in the past few years,', ' were extensively employed. By using radioactive dithiophosphate synthesized from radioactive phosphorus, a rapid and sensitive analytical procedure G. L. SIMARD is in the Research Division, Stamford Research Labs., American Cyanamid Co., Stamford, Conn.; J. CHUPAK, formerly in the Research Division, Stamford Research Labs., is now at Camp Detrick, Frederick, Md.; and D. J. SALLEY is in the Research Division, Stamford Research Labs. AlME New York Meeting, Feb. 1950. TP 2815 B. Discussion (2 copies) may be sent to Transactions AlME before April 30, 1950. Manuscript received Oct. 17, 1949. was at hand. This permitted determination of such important quantities as the rate of uptake of dithiophosphate by the mineral, the amount existing at equilibrium on the mineral surface and in the solution, and the desorption of the agent from the surface. Such measurements were possible even on single crystals of galena. In addition the exchange of dithiophosphate between the solution and the sorbed phase was examined, a matter which could be accomplished only by the use of isotopes. It is desired to point out at this time that throughout this paper the term "sorption" has been employed to designate the uptake of agent by mineral, without implication as to the nature of the process by which the uptake was accomplished. Experimental Materials: Galena: The galena was from the Tri-State district. For studies on ground mineral four preparations were used during the course of the investigation. These were prepared from selected large crystals by wet grinding in order to reduce surface oxidation. Alcohol was chosen as a convenient medium for this purpose. The ground mineral was then fractionated by sedimentation in alcohol, dried by evacuation, and stored in a nitrogen-filled desiccator. Data on the preparations are tabulated in table I. The size analysis indicated that the areas of the preparations were of comparable magnitude, even though the absolute values may be somewhat incorrect. Dithiophosphates: Nonradioactive dithiophbsphate (di-isopropyl or di-secondary butyl) was obtained by purification of a commercial product. An aqueous acid solution of the agent was extracted with petroleum ether, the ether layer dried, and the dithiophosphate precipitated as ammonium salt with anhydrous ammonia. Several repetitions of this process resulted in a nearly colorless, flaky product of good purity. (General formula (RO,)PSSNH,). Radioactive dithiophosphate was synthesized by heating radioactive elementary red phosphorus* • For early experiments (1943 to 1944), the elementary red phosphorus was obtained from the cyclotron group at the Crocker Radiation Laboratory of the University of California. Berkeley, through the courtesy of Dr. Joseph W. Hamilton. For more recent work (1946 to 1949). the elementary radioactive Phosphorus has been supplied by the Oak Ridge National Labarotories on allocation from the U. S. Atomic Energy Commission. with sulphur at 270" to 300°C to produce radioactive P,S,. The latter was then treated with the desired purified alcohol (isopropyl or secondary butyl) at 60" to 80°C to form the dithiophosphoric acid derivative. Purification was effected in the same manner as for the nonradioactive material. The effectiveness of the purification method was established by the isotopic dilution method." The same technique was used to show that only negligible decomposition of neutral or of carbonate solutions of the dithiophosphate took place over a period of a day; this was true whether or not galena was suspended in the solutions. Procedures: Radioactivity: The activity of a solution was obtained by counting with a small glass-jacketed, silvered Geiger counter, using a conventional scaling circuit.' Crystals and other solids were counted under a bell-shaped, mica-window counter tube. Sufficient counts were made so that the probable error of counting was of the order of +I-3 pct. The specific activity in terms of counts per minute

Jan 1, 1951

By H. M. Fisher, R. E. Snow, S. C. Sun

RAPID colorimetric estimation of the amount of transparent minerals has been applied successfully to instream mill products at the cryolite flotation plant of the Pennsylvania Salt Mfg. Co., Na-trona, Pa. Apparatus used was a Bausch & Lomb monochromatic colorimeter. The optical assembly of the colorimeter, shown in Fig. 1, passes light from a 6-v, 32-cp, prefocus incandescent lamp through two condensing lenses, a heat-absorbing filter, an aperture-control diaphragm, an interference filter, and a sample container. The light finally strikes a barrier-layer photocell and the output current of the photocell is measured by a galvanometer. The engraved index line of the collective lens, located between the lamp and the galvanometer, is imaged upon the transmission scale of the colorimeter by the galvanometer mirror. Reflected light from the galvanometer mirror, the position of which is determined by the output current of the photocell, enables reading of percent transmission throughout the mineral sample being tested. The colorimeter, Fig. 1, was standardized as follows: 1—The colorimeter was' connected to a Sola constant voltage transformer, A, of 6-v output operated from a 110-v ac supply. 2—Turning the toggle switch, C, illuminated the transmission scale, B, and the control knob, D, was adjusted with a blank in the optical system until the hairline on the scale read zero. 3—A No. 430 interference filter, E, was inserted into the filter channel, F. 4—A test tube, G, containing 7 cc of distilled water was pressed into the sample space, H, and the fine adjustment knob,

Jan 1, 1955

By A. P. Herring, S. F. Ravitz

Manganese dioxide reacts with sulfur dioxide solution in the absence of oxygen to form manganous ion and dithionate ion, together with a small amount of sulfate ion. The effects of SO, concentration, hydrogen ion concentration, stirring speed, and temperature on the rate of the reaction were studied. If it is assumed that undissociated SO2 molecules react rapidly with MnO, at active surface sites and that bisulfite ions react slowly, an equation can be derived that satisfactorily represents the observed reaction rates. For solutions in which the reaction with undissociated SO2 predominates, reaction rates and activation energies estimated from diffusion theory agree well with those found experimentally. Manganese dioxide dissolves readily in aqueous solutions of sulfur dioxide. The nature of the dissolution reaction was first investigated by Gay-Lussac and welter1 in 1819; they identified a new oxysulfur radical, the dithionate ion, S2O6, as one of the reaction products. Since then, the reaction has become a standard method of preparing dithionates in the laboratory and one of the more promising methods of leaching manganese from low-grade ores. In the "dithionate process" developed by the U.S. Bureau of Mines2,3, manganese-dioxide ore is suspended in a solution of calcium dithionate and leached with gas containing sulfur dioxide. The manganese is dissolved and the sulfur dioxide is oxidized according to the reactions:

Jan 1, 1965

By F. H. Bunge, W. H. Jr. Dailey

This paper covers an investigation of the magnetic roasting characteristics of several iron ores when treated in the form of small pebbles (1/8" to 1/2" in size), and in the form of balled fines. The roasting treatment was carried out with apparatus capable of very close control of time, temperature, face velocity, and per cent combustible in the reaction gas. The effect of rapid initial heating was not studied although it is believed that this has a major effect on the results when processing solid pebbles. The test equipment is a fixed bed type, with such a wide range of gas velocities, temperature, combustible content and etc., that the results obtained can be made to approximate those from various commercial roasting processes. The data presented in this report covers reaction rates, effect of processing variables on silica in the concentrate, flow resistance of beds, and to a limited extent, the effect of particle size. The work is far from complete and raises many more problems than it gives answers for. We believe, however, that much of the information will be useful to those engaged in investigation of magnetic roasting processes. We measured the time required for various degrees of completion of the reaction of hematite to magnetite, for finely pound material in balled form and for carefully sized crushed ore. Measurements were made for a variety of Minnesota and Labrador ores. This report covers an ore requiring fine ginding (-325 M) for liberation (Lower Kevin) and an ore requiring coarser grinding (-100 M) for liberation (Upper Olson). These ores represent approximately the range of characteristics for a group in the west end of the Mesabi Range. Our data shows that the reaction rate is very fast for the finely ground material and that tests are readily reproducible. The reaction rate of the larger size material, in the range of 1/8" to 1/2", is much slower and the results more erratic. We found that for the ores tested, the percentage of silica in the Davis tube magnetic concentrate was a function of iron recovery. Normally the concentrate had a low and relatively uniform percentage of silica up to a critical value of recovery. Above this critical recovery value the percentage of silica increased very rapidly. For example, for one particular ore roasted in the form of 3/8" diameter balls made from -28 mesh fines, the silica remained at about 7% to 8% until the recovery exceeded about 93%. Increasing the recovery beyond 93% resulted in over 25% silica with the additional recovery. The degree of conversion of hematite to magnetite (% completion of roast) required to obtain 93% recovery for balled material, varied inversely with the gas flow rate. Percentage silica in the concentrate resulting from roasting 3/16" and 1/2" solid material was relatively erratic. However, for varying degrees of iron recovery the lowest silica percentages corresponded to those obtainable when roasting fines. It is thought the probable reason for the erratic values was the variation in shattering or cracking of the pebbles during initial heating and roasting. There is some indication that very fast initial heating will greatly increase the permeability of the individual particles of ore to gas flow so that the reaction rate for coarse sizes will approach that obtainable with balled fine material.

Jan 1, 1961

By E. Martinez, M. L. Hollander

Sulfurization of copper oxides was investigated using differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). When mixtures of cuprous oxide and sulfur were heated, the surface of the cuprous oxide was converted to cupric sulfide at the melting point of sulfur (119°C); additional sulfuriza-tion took place when the sulfur vaporized above 200°C. Cupric oxide also reacted with sulfur to form cupric sulfide but the starting temperature was approximtely 225°C. When the cupric oxide and sulfur were physically separated, sulfurization still occurred indicating a reaction with the vapor phase of sulfur. Sulfurization was obtained in runs with cuprite, chrysocolla and basic copper carbonate. Further decomposition of cupric to cuprous sulfide was noted at somewhat higher temperatures. The DTA-TGA data made it possible to set up postulated reactions in each system. These techniques were found to be well suited for studies of this type. Because of the difficulty encountered in flotation of many metallic oxide minerals, efforts have been made to find suitable reagent combinations to sulfi-dize these minerals in the flotation circuit. Recently Bautista and Sollenberger,1 using a different approach, showed that the surfaces of many oxide minerals can be converted to sulfides by heating them with sulfur to temperatures between 150° and 225°C. This technique was effective even with chrysocolla, a mineral which is very difficult to sulfidize. The primary objective of the present study was to obtain information on the reactions between the oxide minerals and suIfur. A better understanding of the mechanisms involved might prove useful in pointing the way to improvements in the process. A secondary objective was to demonstrate the value of thermo-analysis in investigations of this type. Bollin and Kerr2 have reported the use of a modification of differential thermal analysis to study the thermal reactions which occur on heating elemental constituents in a closed system to synthesize copper sulfides and iron sulfides. Differential thermal analysis (DTA) was used to detect enthalpic changes as mixtures of minerals and sulfur were heated. The DTA indicated the temperatures at which reactions started and whether they were exothermic or endothermic. Weight change data were obtained by thermal gravimetric analysis (TGA). These data could often be used to support postulated reactions and to eliminate consideration of other possible reactions. EQUlPMENT Differential thermal analysis (DTA) involves measuring the temperature difference between the sample being tested and a thermally inert reference, generally aluminum oxide, as both are heated at a uniform rate. Series connected thermocouples in the sample and reference are used to detect the temperature differences that occur as the sample undergoes enthalpic changes.3 Crystalline transitions, fusions, vaporizations, dehydrations, decompositions and many solid state reactions can be detected. A DTA unit manufactured by the Robert L. Stone Co., Austin, Tex., was used in this investigation with a temperature rate increase of 10°C per min. The samples and reference were placed in Inconel cups with wells in the bottom into which the differential thermocouples fit, and run in a nitrogen atmosphere. Thermal gravimetric analysis (TGA) measures the changes in weight that a sample undergoes as it is heated. A Chevenard thermobalance converted electronically for graphic recording was used with a heating rate of 5°C per minute. An X-Y recorder plotted temperature on the X axis and weight changes on the Y axis. A flow of 250 ml per min of nitrogen was streamed through the furnace. Samples were placed in a porcelain crucible open at the top, unless otherwise noted. EXPERIMENTAL RESULTS Sulfur: Two sulfur samples from different sources were tested. The DTA of a high purity sulfur and a reagent grade sublimed sulfur are shown in Fig. 1. Nearly all the tests were run with the sublimed sulfur due to the difficulty in grinding and high electrostatic charge encountered with the high purity sample. Sulfur undergoes three endothermal reactions below 200°C. The first is a solid-solid crystalline transi-

Jan 1, 1964