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By Charles C. Hawes

DETERMINATION OF SILICA OXIDE iron ores usually are found in intimate association with silica and phosphorus compounds. Their quantitative separation and purification sometimes present difficult problems to the chemist. Silicic acids may represent various degrees of dehydration. They are exeedingly weak, but combine with metallic oxides and form some of the most stable oxides in nature, the silicates. When iron ores are treated with hydrochloric acid, they yield practically all their iron. However, apart from rare exceptions, varying amounts of insoluble residues of silica, alumina, lime, magnesia and compounds of phosphorus, sulphur, titanium, chromium and others will be left. If the oxides of silicon, aluminum and titanium are present in small quantities, the insoluble residue is generally very small, but the presence of these oxides is more usual in larger quantities. Therefore, only a special treatment can break the bond between these elements. Ore Residues Acids remove the soluble constituent of the ore, leaving the insoluble residue, which chemically is rock or gangue and must be treated accordingly. Alkali carbonates are used frequently to break up the residue, but other fluxes have also been recommended, such as a number of lead and boron salts. The introduction 0f these salts, which offers certain advantages, requires, however, their removal before the other constituents can be analyzed, and this elimination is sometimes difficult. Usually in quantitative analysis silicates are dissociated by fusing them in a platinum crucible with sodium carbonate, thereby forming sodium silicate, Na2SiO3, sodium aluminate, Na2Al204, sodium titanate, Na2TiO3, and carbonates of other elements. All the carbonates are soluble in hydrochloric acid, but also some silicic acid is formed that must be rendered insoluble. To this end the solution is evaporated several times, as the first evaporation still leaves as much as 5 per cent of the silica soluble. Even after a number of evaporations, minute quantities of silica may be found with the iron and alumina present. Perchloric Acid as an Agent for Dehydrating Silicic Acid Perchloric acid, HC104, (60 per cent), has been used for years as a solvent, but few chemists have used it to dehydrate silica. It is derived from anhydrous perchloric C1207, which is most unstable and dangerous. Perchloric acid is a strong oxidizing agent, which boils at 203°C. It is more effective in rendering silica insoluble than any other known agent, and most rapid in action. After perchloric acid is added to a mixture containing silicic acid, the resultant solution is

Jan 1, 1945

By R. P. Aikman, F. G. Miller, J. M. Podgursky

As part of a larger study of the mechanism of froth flotation of coal, the rates at which coal particles of various size and specific gravity react to froth flotation were investigated. Results of this study show that in flotation: I) recovery rate increases, but selectivity decreases with a decrease in particle size down to 100 mesh, and 2) recovery rate is constant for all sizes between 100M and 325M. n response to the growing demand for low-sulfur coking coal for iron making, the coal industry is continually striving to perfect methods for cleaning coal to acceptable ash and sulfur levels while maintaining maximum product yield. By present standards the industry has adequate processes for cleaning coarse coal, i.e., + 28 mesh, at almost any desired specific gravity. However, in the case of fine coal, i.e., —28 mesh, neither of the two existing processes - froth flotation or hydrocycloning — in its present state of development yields products that are comparable in ash and/or sulfur content to products normally obtained from coarse coal processes. The work outlined in this paper is part of a larger study directed toward maximum exploitation of froth flotation for fine coal processing. The concepts and data presented are now serving as a basis for designing fine-coal circuits in which improved efficiency is obtained by optimum use of froth flotation with other fine-coal processing methods. The paper includes: 1) A discussion of the value of a detailed wash-ability study as a method of assessing the problems associated with fine coal processing, 2) A general discussion of hydrocyclone principles, included to serve as a basis for understanding the mechanism of coal flotation, and 3) A detailed discussion of the mechanism of coal flotation, as well as the research leading to the optimum use of froth flotation in combination with other fine-coal cleaning methods to provide efficient fine-coal processing. Our study of the rates at which coal particles of various size and specific gravity react to flotation shows that in flotation: 1) recovery rate increases but selectivity decreases with a decrease in particle size down to 100 mesh, 2) recovery rate is constant for all sizes between 100 mesh and 325 mesh, and 3) the flotation process effectively rejects only those extremely fine particles which do not contain surface carbon. (Because of this feature, froth flotation is an excellent method for rejecting clays - the only extreme fines in - 325-mesh coal that are always carbon free.) Washability studies of flotation products are used to assess flotation efficiency. These studies show that: 1) the increase in percent sulfur in flotation products above the theoretical sulfur content as defined by the washability study is due almost entirely to flotation of fine high-specific-gravity particles that are part coal and part pyrite, and 2) in a given size range particle floatability decreases with an increase in particle specific gravity. Error curves are explained and are used to compare the mechanisms of coal flotation and hydrocycloning. These comparisons show that hydrocyclone performance is affected primarily by particle mass, i.e., particle specific gravity and size, whereas coal flotation performance is dependent on both the particle mass and the coal surface available for bubble attachment. Since the error curve is a basic tool for rating the efficiency of any coal processing system, it will be useful to quickly review the principles of error curve construction. RATING COAL-PROCESSING EFFICIENCY BY THE ERROR CURVE The organic component of coal is of lower specific gravity than its impurities, and the most effective principle for coal concentration is therefore separation on the basis of specific gravity. However, since there is no sharp dividing line between pure coal and pure ash components, specific gravity must be closely controlled if the product is to have acceptable quality at maximum yields. In turn, the basis for close control in any coal processing system is the reference standard provided by precise laboratory washability studies. In these studies, coal is carefully separated into a series of progressively higher specific gravity fractions to establish a relationship

Jan 1, 1968

Jan 1, 1938

An "international gold conference" was held at Spokane, Wash., on Sept. 5, 1918, under the auspices of the Northwest Mining Association, which was attended by a large number of engineers, mine operators, and bankers, representing the northwestern states, as well as Alaska and British Columbia. The principal subject under consideration was the doleful condition of the gold-mining industry, gripped between the diminished purchasing value of its product, and the increased cost of its operation. Two resolutions were adopted by the meeting. The first approved the resolution adopted Aug. 12 by the American Gold Conference, a permanent organization formed at Reno, Nev., in which it is stated that relief by the United States to the gold producers of this country should he made to correspond with the increased cost of production, the extent of the relief to be fixed from time to time to Meet changing conditions. The second resolution, adopted after serious and long-continued discussion, was introduced by F. A. Ross, mining engineer, of Spokane, and a member of the Institute, and was as follows: WHEREAS, it is now clearly understood and agreed by the allied nations that national gold reserves must be augmented without delay if rapidly expanding credits are to be secured; and WHEREAS, at this time, the production of gold is daily decreasing to an alarming extent because of well known conditions that, render the mining of gold-bearing ores unprofitable; and WHEREAS there is no way by which the mining of gold-bearing ores may be made profitable except by the direct' aid and influence of the government; and WHEREAS the government has appealed to the mining industry for practical information and practicable suggestions upon this subject; therefore, be it

Jan 10, 1918

By M. J. Pool, C. E. Lundin

THE relative partial molar enthalpies of aluminum, gallium, and indium in liquid tin have been measured at 750°K by liquid-metal solution calorimetry. The measured heat effects and the calculated relative partial molar heats of solutions at infinite dilution are listed in Table I along with the relative heat contents of the solutes calculated from the data of Kelley. All of the values are referred to 750°K even though the calorimeter temperature fluctuated by +1°K during the entire series of experiments. The effect of this small temperature variation on the experimental values should be less than the experimental error. The limits of error given in Table I represent the standard deviations of all values from the mean. In no case was the experimental deviation greater than the listed error limits. The reported error estimates only apply to the estimated precision of experimental values and not to their accuracy. Systematic errors could be present in the calorimeter which would affect the absolute values given. Moreover the heat-content data used could also be in error and this would affect the end results. The relative enthalpies used were listed in Table I so that future corrections could be made if necessary. The relative partial molar heats of solutions of aluminum in liquid tin and dilute A1-Sn liquid solutions are large and positive. The value at infinite dilution is given in Table I. In the dilute-solution region investigated the relative partial molar heat of solution was a linear function of composition. A least-squares analysis of the data gave the following expression at 750°K: Atf^ = 6195 - 12,160 A:A1(cal/g-atom) where XA1 is the mole fraction of aluminum in the alloy. This can be compared with the data of Cohen, Howlett, and ever' who obtained the expression: 1SM = 6075 - 10,750 XM at 573°K Apparently both the partial molar heat of solution at infinite dilution and the concentration dependence increase with increasing temperature. This same

Jan 1, 1964

By A. D. S. Gillies

The effectiveness of coal mine face ventilation and the ability of airflow to maintain satisfactory methane and respirable dust levels in the working face area has been studied through the use of scale modeling techniques and underground testing procedures. In particular, consideration has been given to examination of the effects of operation of a high airflow, machine mounted dust scrubber unit on face air patterns. Results obtained through a cooperative study with the Peabody Coal Co. demonstrate that new mining cycles can be considered that increase production efficiency while maintaining satisfactory ventilation conditions.

Jan 1, 1986

By R. V. Ramani, T. V. Falkie, T. E. Wilson

In order to increase the efficiency of underground coal operations, it is necessary to understand not only the operating system in detail, but the interrelationships between system components and the operating environment. With such an understanding comes more exact application of simulation and other operations-research techniques to plan and operate mining systems. One aspect of the coal-winning system in need of better definition, namely, the performance of continuous miners, is dealt with in this paper. Detailed discussion of this research will be found in a report by the authors. Mining Rate The performance of continuous miners-that is, the rate at which miners can deliver coal from an underground face to the outby transportation system-is dependent upon the geometric and geological conditions of mining, the characteristics of the mining machinery, and the skill of the machine operators. In order to predict mining performance, it is necessary first to identify these factors, secondly to devise a data-collection technique, and lastly, to select a method of data analysis for drawing meaningful conclusions. Factors Affecting Mining Rate: The factors that affect the mining rate can be conveniently grouped in the following four categories: Geological Factors-The natural conditions such as seam thickness, hardness, and cleavage. Geometrical Factors-The height and width of the entries and the angle of drivage with respect to the main entries. Machine Factors-The mechanical characteristics and physical dimensions of the machine, condition of the machine, condition of the bits, and related items. Other Factors-Operator efficiency, size consist of cutter product, etc. A classification such as given here is not rigid; some of the factors can be included in more than one of the categories and some can be combined as ratios. Data Collection The empirical nature of this investigation required measurements of variables under operating conditions in underground coal mines. A majority of the variables measured were continuous over a specific range of values; for example, the "width ratio" varied in practice from 0.25 to 1.00. "Dummy" values were assigned to certain operational variables in order that the variables would have numerical expression, i.e.: [ ] Data Analysis The method of analysis was to test a large model containing potentially significant variables and then reduce the model to one containing only the essential significant variables. The procedure of "scanning" a large proposed model and subsequently condensing the model consisted of two cycles of the same series of computational procedures. Correlation analysis, factor analysis, and regression analysis were the three statistical procedures that were used to derive the rate-generating equation. Correlation and Regression Analysis: Multiple linear correlation coefficients between all of the variables were calculated using the Pearson product moment formula along with the statistical significance of the correlation

Jan 1, 1975

By L. R. Plitt, M. K. Kim

The collecting properties of the fatty acid type collectors for barite were studied using zeta potential measurements, infrared spectroscopy, and Hallimond tube flotation tests. Based upon the experimental results, an adsorption model has been formulated which emphasizes the predominant role of the hydrocarbon chain in collector adsorption.

Jan 1, 1975

By Joseph L. Gillson

YOUR speaker has long sought an opportunity to review the many differences between the subject matter called economic geology and the duties of a practicing economic geologist. As the subject was taught in my student days, professors teaching courses in economic geology paid little attention to what we now call industrial minerals. Certainly they were not concerned with problems involving production and marketing, but only with the origin of such mineral deposits as the Stassfurt potash deposits; with other evaporites, such as salt, gypsum, and Chilean nitrate; with the possible modes of formation of sulfur in salt domes; and the causes of asbestiform mineralization. Of course, much time and thought was given to the mineral identification and sequence of deposition of the minerals in pegmatites. The fact that such deposits actually produced commercial feldspar, mica, beryl, and lithium minerals for sale in industry was of minor importance. These observations lead us to pause to inquire, perhaps facetiously, what is and what is not to be included in the science of economic geology?

Jan 5, 1957

By C. W. Wark, Fred Wise

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By Fred Wise, C. W. Wark

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By T. Egleston

circumstances, would prefer the steel with which they are now familiar, to a specimen that Mr. Sandberg has described as having broken into seventeen pieces under the wheels. After blowing such low manganese steel, it may be coaxed into a rail, but it is a wonder that it holds together so long as it does, with so great a nornber of minute flaws. I would not in any way depreciate chemistry, but I think it should be kept in its proper sphere. Let the chemist look after the quality of pig metal, and apply common sense in the avoiding of extremes, then the most fastidious railroad cannot find fault with the result. PROFESSOR EGLESTON, New YORK City : It is not my place as an engineer to apologize for the chemists, but as no one seems disposed to do so, and as they have had more than their share of criticism, 1 am glad to say that I believe there are chemists in this Institute whose work and word are just as reliable, arid perhaps even more so, than that of the average engineer. But we ought to make a distinction; there are chemists and chemists. With the ordinary commercial chemist, who looks upon the science as so much merohandise, I have not a particle of sympathy; but with the chemist who looks upon his profession as engineers do upon theirs, I have every sympathy. When matiufactorers and engineers go to a chemist and ask him to make an investigation, and scretv him dowu to the lowest point, turning the equivalent of his brains into cents and mills, they ustially get an exact equivalent in poor work for miserable pay, and no one has or should have any sympathy with them, and the under these conditions has no right to ask for any, and no reason to criticise any work that he may get under such circumstances. But I am disposed to think that the chemists who have been replwented and discussed at this meeting do their work conscientiously, and that it is as reliable as that of almost any profession. I believe, however, that this problem of steel rails is being investigated in a wrong direction. I said so at the Pittsburgh meeting, and I think the discussion of this meeting will prove it to all those who have heard it. I think the chemigt is incapable of solving this problem unlees he goes very far into the domain of physical chemistry, so far that it becomes physics and not chemistry; and that the physicist will be the one on whom we must rely in the future for the elucidation of the subject. The chemist may aid the physicist, but it is my decided opinion that we are looking in the wrong place to get the explanation of the phenomena.

Jan 1, 1881

By J. C. Williams

The operations described in this paper are those of the flotation plant of the Valley Forge Cement Co. at Conshohocken, Pa., and concern the use of a particular cationic reagent—namely, DP-243. The role of this reagent in this plant may be better understood by first considering the purpose of beneficiating cement rock in general and the specific problems at the Valley Forge plant. The ultimate purpose of beneficiating cement rock is to obtain a kiln feed of the desired chemical composition in a form suitable for the most efficient kiln operation. The chemical composition is governed by the type of cement to be made. Before government restrictions (the purpose of which is to simplify operations during wartime) limited the types of portland cement a manufacturer is permitted to produce, five principal types of cement were generally recognized by the industry. They are based upon chemical and physical specifications proposed by the American Society for Testing Materials, federal and other governmental agencies, state highway departments, water boards and other users of large quantities of cement. The physical as well as chemical specifications for any particular type of cement necessitate proper proportioning of the major components—that is, the silica, alumina, iron and lime—and it is necessary to establish and maintain definite limits for each in the raw mix. The manufacturer must then prepare a kiln feed of a composition that will yield a final product that will fulfill the desired specification. This step is further complicated for kilns using coal as fuel, by the amount and analysis of the ash introduced. The amount and nature of the dust lost during burning will also affect the composition of the kiln feed. The problem of preparing a kiln feed of the desired chemical composition is not a simple one, and it is the final product that governs the methods used to obtain such a feed. For example, in starting with a material deficient in lime for the mix sought, it may not be sufficient to raise it to the required lime content. The other constituents must also be present in the proper proportions. In some cases, it may be necessary to raise the lime content to a point greatly exceeding that of the final mix, then, by addition of one or more other materials lower in lime, obtain the mix. The rock used at the Valley Forge plant is a metamorphosed argillaceous limestone of the Jacksonburg formation, the same formation as that of the Lehigh Valley cement rock (which occurs approximately 50 miles north), and its oxide composition is similar to that of the latter. The only practical differences between the two are that, as a result of metamorphosis, crystallization of the Valley Forge rock is considerably coarser and it contains less carbon than the Lehigh Valley rock as well as slightly more magnesia in the form of magnesian mica. The calcium carbonate content of the various strata in the Valley Forge quarry ranges from about 65 per cent to over 85 per cent, the average for the entire quarry being somewhat under the requirement for manufacture of cement, which usually is from 75 to 76 per

Jan 1, 1947

By J. C. Williams

The operations described in this paper are those of the flotation plant of the Valley Forge Cement Co. at Conshohocken, Pa., and concern the use of a particular cationic reagent—namely, DP-243. The role of this reagent in this plant may be better understood by first considering the purpose of beneficiating cement rock in general and the specific problems at the Valley Forge plant. The ultimate purpose of beneficiating cement rock is to obtain a kiln feed of the desired chemical composition in a form suitable for the most efficient kiln operation. The chemical composition is governed by the type of cement to be made. Before government restrictions (the purpose of which is to simplify operations during wartime) limited the types of portland cement a manufacturer is permitted to produce, five principal types of cement were generally recognized by the industry. They are based upon chemical and physical specifications proposed by the American Society for Testing Materials, federal and other governmental agencies, state highway departments, water boards and other users of large quantities of cement. The physical as well as chemical specifications for any particular type of cement necessitate proper proportioning of the major components—that is, the silica, alumina, iron and lime—and it is necessary to establish and maintain definite limits for each in the raw mix. The manufacturer must then prepare a kiln feed of a composition that will yield a final product that will fulfill the desired specification. This step is further complicated for kilns using coal as fuel, by the amount and analysis of the ash introduced. The amount and nature of the dust lost during burning will also affect the composition of the kiln feed. The problem of preparing a kiln feed of the desired chemical composition is not a simple one, and it is the final product that governs the methods used to obtain such a feed. For example, in starting with a material deficient in lime for the mix sought, it may not be sufficient to raise it to the required lime content. The other constituents must also be present in the proper proportions. In some cases, it may be necessary to raise the lime content to a point greatly exceeding that of the final mix, then, by addition of one or more other materials lower in lime, obtain the mix. The rock used at the Valley Forge plant is a metamorphosed argillaceous limestone of the Jacksonburg formation, the same formation as that of the Lehigh Valley cement rock (which occurs approximately 50 miles north), and its oxide composition is similar to that of the latter. The only practical differences between the two are that, as a result of metamorphosis, crystallization of the Valley Forge rock is considerably coarser and it contains less carbon than the Lehigh Valley rock as well as slightly more magnesia in the form of magnesian mica. The calcium carbonate content of the various strata in the Valley Forge quarry ranges from about 65 per cent to over 85 per cent, the average for the entire quarry being somewhat under the requirement for manufacture of cement, which usually is from 75 to 76 per

Jan 1, 1947

By Joseph Kittredge

AT the time this matter first came up in 1912, the National Malleable and Steel Castings Co. had seven basic-bottom open-hearth furnaces in its plant at Sharon, Pa., using fuel oil, then costing about 2 ½ ¢ per gallon. The fuel was satisfactory but its cost fluctuated widely, therefore the use of pulverized coal, which had been successful in the cement industry and in some metallurgical furnaces, was suggested. After a thorough investigation, and an actual trial on one of the open-hearth furnaces, to be certain that the siliceous ash in the coal would not seriously affect the basic slags, the management of this company approved its adoption and a complete installation was started. The conclusion had been reached that the success of the operation depended on pulveriz-ing the coal to greater fineness than had thus far been accomplished, and all equipment was chosen with this in mind. A coal drier was deemed essential, and one of ample capacity and efficiency was selected, which has done all that was expected of it, working without failure and with very small maintenance cost, and is still in operation.

Jan 1, 1939

By John Mills

THREE case histories from professions other than engineering will serve to introduce ideas basic to this discussion. Case (1) Date, about 1900. A young man, B. D. from a three-year graduate course in a prominent divinity school, was for two years a successful pastor of a small church; then friction over theological matters arose with a fundamentalist deacon of the "Bible belt" type. He resigned, returned to his university for a four-year medical course, interned for two more, and thereafter was successful and apparently happy as a physician. We may assume that he was converted in early adolescence and vowed himself to service as a minister. Acceptance of his announced career by parents, friends, and teachers plus his own determination confirmed his choice until illiberal criticism threatened his usefulness and he awoke to the fact that his inner needs were not along the direction in which inertia and his conscience had constrained him.

Jan 1, 1947

By E. H. Rose

The present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post- for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: 1. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. Analogy with Grinding In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1947

By E. H. Rose

The present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post- for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: 1. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. Analogy with Grinding In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1947

By Fred E. Beasley

Metallurgists have, in the last ten years, overcome many difficulties of high zinc in lead blast-furnace slags. This problem was brought to the front at the close of the war, by the price obtained for lead and the high cost of labor, fuel, and flux. The Port Pirie plant of the Broken Hill Associated Smelters, Australia, early in 1919 had increased the zinc oxide in its blast-furnace slags from 13.5 to 20 per cent., at the same time reducing the iron oxide in the slag from 33.5 to 25.6 per cent. These results were obtained by the addition of 10 to 12 per cent. granulated slag to the roaster charge, improved conditions of the crushing, mixing, and roasting of the constituents of the sinter. By experiments and improvements, the practice at this plant has developed a normal slag that carries 22 per cent. zinc oxide, and at times runs as high as 33 per cent. zinc oxide, with only slight furnace trouble. The granulated slag is used not only as a diluent to the sulfur in the roaster charge, but on account of its value as a slag-forming material and the physical properties it imparts to the finished sinter. The ore is mixed with flux and given a double roast. All the first roasting is by sintering with the Dwight & Lloyd machines. The final roasting is performed partly by Dwight & Lloyd machines and partly by H. & H. pots. The charge to the blast furnace consists chiefly of the final sinter, slag shells, siliceous and non-bearing sulfur lead ores. High blast and ore columns are the practice. The greatest care is taken to obtain in the final sinter as low a sulfur content as possible. Attracted by the success of the Port Pirie plant, The Consolidated Mining & Smelting Co. of Canada, at Trail, B. C., in 1920 treated its crushed primary sinter with about 15 per cent. granulated slag. The Trail smelter at this time was short of ore and a large part of the charge to the roasters consisted of the zinc-plant residue having the following analysis: 12 to 16 per cent. lead, 12 to 16 per cent. zinc, and 25 to 30 per cent. iron. The lead in the final sinter was low and the physical condition was so bad that all the sinter had to be screened before it could be charged to the blast furnace. Also it disintegrated on being trammed and dumped

Jan 1, 1925

By C. S. Allen

With increased coal production and mechanization of coal mines many transportation problems arise. The main objective is to bring the coal to the tipple or dump it as quickly as possible. Larger and faster haulage locomotives or two locomotives connected in tandem have been adopted to solve the problem. Motor equipment, gears and control systems of locomotives to be placed on these hauls are carefully studied to meet these requirements, by an analysis of motor characteristic curves with the corresponding tractive effort curves, in order to determine the correct load for maximum efficiency and to determine the approximate schedule. Control equipment is usually considered to be the controller and resistance. Sometimes accelerating curves are calculated in order to obtain smooth starting in the least possible time without slipping the wheels. Train control is very important, but why limit this to acceleration ? Deceleratien-or-retardation of the locomotive is a part of the train control and should be given as much consideration as the control for acceleration. Types or Braking Common braking systems for mine locomotives utilize screw-type mechanical brakes, dynamic braking and the air brake. For gathering and haulage locomotives, the srrew-type mechanical brake is regarded as standard. A brake of this kind cannot be applied without much effort and lapse of time. Motormen, like other people, follow the path of least resistance, and therefore have their brakes ready for instant application—which means that the shoes are virtually dragging. Even then, a motorman bucks and plugs the motors to gain time and avoid as much physical effort as possible. "Bucking of motors" consists in turning the reverse cylinder of the controller to parallel position corresponding to the opposite direction of travel while the locomotive is in motion, without applying power from an external source. When this occurs, both motors tend to act as series generators, because the armatures are revolving and there is residual magnetism in the field circuit. Owing to slight differences in motor manufacture, one of these will generate a slightly higher potential and will overcome the generated voltage of the second and tend to drive it as a motor in the reverse Direction.-This action retards the locomotive but is very severe on the segments of the reverse cylinder and on the motor commutators. "Plugging motors" is the application of power to the motors corresponding to the opposite direction of travel. In order to "plug" the motors it is necessary first to "buck,, them (which reverses the connections to the motors) and then to apply power. Theoretically, this should be done only in an emergency, but locomotives in mine service are frequently thus treated., Dynamic braking is retardation of the locomotive by the motors acting as generators, which take no power from the line.

Jan 1, 1942