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By R. S. Silo

A new process for leaching oxidized copper ore (CuO) is presented. Its principle advantage is that it requires low cost raw materials. For large operations the only materials used are sodium chloride (NaCl) and calcium carbonate (CaCO3). In those operations of medium size, sulfur, which is abundant in the North of Chile, is used. However the process may also be applied to small copper plants provided low cost sulfuric acid is available as a substitute for sulfur. That is, it may be applied to those small capacity plants where large additional installations for the regeneration of the leaching solution are not needed. The method is comprised of six stages, namely: Crushing, Leaching, Washing of the Solution, Precipitation, Regeneration of the Leaching Solution, and Fusion. A flow diagram for this process in a large or medium-sized plant is given. LEACHING This process requires hydrochloric acid as the leaching solution for treating copper ores. This is easily produced from sulfuric acid and sodium chloride as follows: A small excess of NaCl is advisable to avoid the presence of free sulfuric acid. The experiments reported here were performed with 36 grams of HC1 per liter of solution to obtain a free acidity equivalent to 50 gpl of H2SO4 (the usual leaching solution concentration). Hydrochloric acid leaching may be done by either of the usual methods; i.e. percolation or agitation. The method selected depends on the characteristics of the ore. No matter which method is chosen, low pressure air can be used either as an air lift for circulation of the solution, or for agitation of the pulp with "Pachuca" tanks. In either case, there are no problems with respect to oxidation. Leaching by agitation produces good recovery; but at times, the problem of high consumption of leaching solution or the overproduction of grinding slimes which interfere with contact between phases negates this advantage. The percolation method was used in this study. Percolation took place in four 15-liter containers, 11 1/2 in. diam. by 12 in. height, having hemispheric bottoms. Each container was fitted with a wooden bottom having 1/4 in- holes. The air lift was a 3/4 in. plastic tube 10 in. high. Air at 6 psi was supplied from a rotating laboratory compressor. 10 kg of ore were charged to the container and counter-current leaching was used. The leaching solutions were fed through the center pipe in order to use the ascending solutions to reject the air in the charge. The solution circulated through the air lift in order to facilitate the movement according to the increase in density produced by the reaction between the phases present. Specific observations made in this report cannot be generalized because the leaching characteristics are dependent on the nature of the ore used, the method of leaching, and the production of primary or secondary slimes while crushing or leaching.

Jan 1, 1967

By A. W. Dana, F. J. Shortsleeve, A. R. Troiano

The phenomenon of flake formation which may occur during cooling or room temperature aging of large steel sections is caused by a combination of hydrogen and stress. As such, the transformation characteristics of austenite in a particular section play a major role in the occurrence of flakes. Isothermal and continuous cooling studies demonstrated that flake formation is particularly sensitive to nature, distribution, and relative proportions of the microconstituents in the cooled sections. In the absence of transformation stresses, abnormally high hydrogen contents could be tolerated without flaking. Flakes were not found in the absence of transformation stresses. No correlation existed between average hydrogen content and flake formation where cooling stresses were low. Hydrogen in the steel was a necessary but not sufficient condition for flake formation. DESPITE a long history of laboratory and commercial investigations, the mechanics of hairline crack or flake formation in large steel sections remains unresolved. Recent publications1-5 emphasize that flaking is still a serious problem.* immediate cause of flaking is the pressure exerted by entrapped hydrogen gas that collects in small internal cavities or voids during cooling. When the temperature of steel is low enough to preclude extensive plastic flow, the cavity pressures cause local brittle fractures (flakes). 2—The microsegregation and transformation stress theory. According to this theory, flaking results from stresses arising from the volume changes accompanying the formation of martensite and/or low temperature bainite after a large portion of the section has transformed at elevated temperatures. The proponents of the hydrogen theory suggest that stresses from other sources may be additive to the stresses produced by hydrogen and may, in some cases, provide the increment of stress necessary to produce flakes. However, the critical aspect of this theory is that no stress, except that produced by hydrogen pressure, is believed to be sufficient in itself to produce flakes. For practical considerations, this means that the only safe thermal cycle is one that reduces the hydrogen content to some low critical level below which flaking will not occur regardless of the state of transformation. On the other hand, the transformation stress theory does not rule out the embrittling effect of hydrogen. This theory simply proposes that flaking does not occur in the absence of transformation stresses. With the transformation stress theory as a basis, commercial antiflaking cycles would be designed to complete the transformation of austenite prior to cooling to ambient temperatures. Both of these theories appear compatible with industrial experience. Antiflaking cycles, such as cooling at extremely conservative rates or soaking for periods of time in the upper subcritical temperature ranges prior to cooling to ambient temperatures, may allow the following: 1—diffusion of hydrogen out of the section and, as a consequence, reduction of the hydrogen content to some safe minimum value (hydrogen pressure theory) and 2—complete decomposition of the austenite at rela-

Jan 1, 1956

By Yves Dardel

THE literature concerning the remelting of aluminum and its alloys is very rich. Unfortunately the greater part of the papers about this subject have no scientific value, for the theories or explanations are too often in opposition to the laws of physical chemistry. Furthermore only a few of the methods of purification1,2 are considered in the literature. Hence it has seemed necessary to the author to sketch a general theory of the purification of secondary aluminum. The purpose of the purification is to eliminate the impurities that the charges and the atmosphere of the furnaces bring into the bath during the successive remeltings. Indeed scrap as well as metal can be used when soiled with oil, greases, water, .. , if it has been sufficiently dried and contains oxides. At the melting point of the metal, the oils and greases are decomposed into their constituent elements H2, 02, H2O, C02, CO, CH4, and others. These gases are added to the corresponding gases in the atmosphere of the furnace. The impurities introduced into the metal during the remelting will accordingly be the products resulting from the action of these gases on the liquid bath. The quantity of hydrogen that is soluble in aluminum and its alloys is given by the relation of Sieverts3 as a function of its partial pressure PH2 in the atmosphere [m = K _%/PH, [I]] where K is a constant, which depends only upon the temperature and increases with it. The experiments of Röntgen and Braun4, Röntgen and Möller5, Bircumshaw6 and Winterhager7 have rendered it possible to determine the value of K, when the temperature is inferior to 1.000°K. One obtains: [log K = - 4,702.5 + 0.238 [2]] Besides its usual signification, the term PH2 represents also the internal pressure of dissolved hydrogen, since in the equilibrium, that is, when no gas exchange occurs between the two phases, the pressures of hydrogen in the metal and in the atmosphere must be equal. Eq I shows that when the partial pressure of H2 increases the amount of dissolved hydrogen in equilibrium increases, and that, if the contrary occurs, the quantity of H2 should diminish. However, if for any reason the diffusion of hydrogen is hindered and the total amount of dissolved gas remains constant while the temperature lowers, the internal pressure of H2 increases until it reaches the value of the hydrostatic pressure. At this point, bubbles of hydrogen appear in the bath and escape out of it. The amount of hydrogen in the metal is then limited by a maximum depending on the hydrostatic pressure, which is practically equal to one atmosphere. Indeed hydrogen does not accumulate in the form of a hydride appearing as a separate solid phase. However, even the mechanism of the solubility is not yet elucidated. One does not know exactly as yet, whether the

Jan 1, 1947

By Hjalmar W. Johnson

It has long been accepted that blastfurnace practice varies to some degree with the coke used. While the qualities desirablc in iron have been known for some time, the qualities in coke that produce such iron have not been established. There are no objective standards for blast-furnace coke, which is usually described in general terms, as good or bad. Many experienced operators believe that coke cannot be measured objectively except by using it in the furnace, noting the result, and labeling the coke according to furnace operations. This is admittedly unscientific, and besides is useless for the two main purposes for which a measure of coke quality is needed: (I) to produce the best possible coke from the coals and equipment on hand; (2) to test new coals and determine their value for blast-furnace coke. In seeking such an objective measure, much work has been done on tumbler tests, shatter tests, sizes of coke, porosity, specific gravity, reactivity or combustibility and other tests, but as yet very little has been published in which conclusive correlations have been established between such objective measures and furnace operation. In our study of the problem during the past 8 years we have developed no new measures of coke quality, but we have collected data that are presented here in the hope that they may assist in further work, so that at some future time there may be developed such positive measures. The purpose of this paper is to show that there are variations in coke and that such variations do influence the operation of blast furnaces. At the same time we must remember that there are variations in blastfurnace operation and it may be the blast furnace that is at fault. This correlation of coke properties with blast-furnace operation will be worked out in the following steps: . Discussion of a period when a change in the coke resulted in an increase in efficiency on one furnace and a decrease in efficiency on another. 2. Adaptation of the inefficient furnace to the available coke so that normal efficiency was restored. 3. The use of radial distribution: a discussion of the effect of size of ore layer in adapting blast-furnace operation to the coke. 4. An example of a change in size of coke and the detrimental influence on the pro duction of one furnace. 5. An example of a change in size of coke, together with manufacturing data, to show why it changed, the conclusion being that coke is a product with an optimum value. 6. Illustration of the fact that coke has different optimum values on two furnaces, with discussion on adaptation of operation so the optimum values will be the same. 7. Limitations of the use of screen tests as an objective measure of coke quality. 8. Importance of uniformity in coke. Functions of Coke and Qualities Required to Perform These Functions Coke has three important functions in a blast furnace: (I) to furnish most of the

Jan 1, 1942

By C. C. Wright, R. J. Day

ALTHOUGH data on the exact tonnage of recoverable coal silt are not known, the quantity produced in 1943 was estimated to be over five million tons for the anthracite region of Pennsylvania alone. Since that time the recovery of the No. 5 buck- wheat size previously included in silt estimates has increased greatly thus reducing the total silt tonnage. As a compensating factor, however, the methods of recovering silt previously escaping into the rivers and streams have been markedly improved so that the net figure may not have changed materially. Comparable estimates for bituminous coal are lacking, but undoubtedly the silt represents a significant tonnage. With the trend toward increased wet cleaning of the smaller sizes of bituminous coal, it is to be expected that even greater quantities of silt will be available in the future. The raw silt from anthracite collieries usually contains a high mineral matter content, but technological developments in fine coal cleaning have been making rapid advances and it appears probable that more extensive use of these methods will be employed in the near future. Bituminous coal silt in many instances contains a similar high mineral matter content. As with anthracite silt, however, the cleaning of this material, where necessary, appears both technologically and economically feasible. With a view to upgrading the size and value of silt, considerable laboratory and pilot plant work has been conducted by a number of organizations on what is commonly known as pelletization. Actually the process is one of extrusion in which the moist silt is extruded from a die in the form of rods, which usually break into relatively short lengths-a few inches long-as they fall from the die. Fig I is a photograph of some extruded 3/8-in. diam pellets and of the anthracite silt from which they were produced. This technique of producing larger sized material from fines is by no means new, having been used extensively by the industrial carbon and clay industries for years.' The application of the process to coal silt appears, however, to be relatively new. The first major program of investigation dates back only to July 1945, when a group of engineers representing the Anthracite Institute, The Pennsylvania State College and the J. F. Pritchard and Co. conducted preliminary tests on the extrusion of anthracite silt at the Floridin Co. plant at Warren, Pa. Since that time these three organizations-individually, together, and in co- operation with other interested groups- have actively pursued work on various phases of the problem. Much of the early history and details of the laboratory and pilot plant work by the Anthracite Institute, has been reported by Mulcey and Eckerd.2 It is the purpose of this paper to report additional information on the process, especially with respect to the factors influencing the quality and use properties of the product, and the current develop

Jan 1, 1948

By Hjalmar W. Johnson

It has long been accepted that blastfurnace practice varies to some degree with the coke used. While the qualities desirablc in iron have been known for some time, the qualities in coke that produce such iron have not been established. There are no objective standards for blast-furnace coke, which is usually described in general terms, as good or bad. Many experienced operators believe that coke cannot be measured objectively except by using it in the furnace, noting the result, and labeling the coke according to furnace operations. This is admittedly unscientific, and besides is useless for the two main purposes for which a measure of coke quality is needed: (I) to produce the best possible coke from the coals and equipment on hand; (2) to test new coals and determine their value for blast-furnace coke. In seeking such an objective measure, much work has been done on tumbler tests, shatter tests, sizes of coke, porosity, specific gravity, reactivity or combustibility and other tests, but as yet very little has been published in which conclusive correlations have been established between such objective measures and furnace operation. In our study of the problem during the past 8 years we have developed no new measures of coke quality, but we have collected data that are presented here in the hope that they may assist in further work, so that at some future time there may be developed such positive measures. The purpose of this paper is to show that there are variations in coke and that such variations do influence the operation of blast furnaces. At the same time we must remember that there are variations in blastfurnace operation and it may be the blast furnace that is at fault. This correlation of coke properties with blast-furnace operation will be worked out in the following steps: . Discussion of a period when a change in the coke resulted in an increase in efficiency on one furnace and a decrease in efficiency on another. 2. Adaptation of the inefficient furnace to the available coke so that normal efficiency was restored. 3. The use of radial distribution: a discussion of the effect of size of ore layer in adapting blast-furnace operation to the coke. 4. An example of a change in size of coke and the detrimental influence on the pro duction of one furnace. 5. An example of a change in size of coke, together with manufacturing data, to show why it changed, the conclusion being that coke is a product with an optimum value. 6. Illustration of the fact that coke has different optimum values on two furnaces, with discussion on adaptation of operation so the optimum values will be the same. 7. Limitations of the use of screen tests as an objective measure of coke quality. 8. Importance of uniformity in coke. Functions of Coke and Qualities Required to Perform These Functions Coke has three important functions in a blast furnace: (I) to furnish most of the

Jan 1, 1942

By G. K. Manning, E. R. Stein, E. E. Underwood

Columbium, carbon. and nickel additions were made to iron-base alloys with 20 pct CY. The effects on microstructure, precipitation-hardening characteristics, and High-temperature properties were investigated by means of metallographic, hot-hardness, and magnetic measurements. An austenitic alloy consistilzg of Fe-20Cr-14Ni-0.8C-2.5Cb had higher hot-lzaciness at 1700°F than the other alloys investigated. Its favorable standing is attributed to an enhanced interaction of Cb and Cr4C particles dispersed within a stable austenitic matrix. FOR the past 20 years, cobalt- and nickel-base alloys have been used for the more stringent high-temperature applications because of their excellent creep resistance. Although their performance is quite satisfactory, other base materials must be considered for large tonnage, low-cost applications such as for the automotive gas turbine. Iron-base alloys are also used extensively, but at somewhat lower temperatures, primarily because of the instability of particles or because precipitation occurs at relatively low temperatures. Since unusual stability has been reported for Fe-Cb and Cb-C compounds.1-3 it was hoped that effective amounts of these compounds could be taken into solution at temperatures just below the melting range, then precipitated upon reheating. Actually, strengthening was found to occur by both precipitation hardening in some alloys and by phase transformations in others. Among the alloys investigated, that with the best high-temperature properties owes its strengthening to a combination of carbide particle interaction and a relatively stable austenitic matrix. EXPERIMENTAL PROCEDURES Small preliminary melts of about 100 g were made in a vertical Glotube furnace, using alundum crucibles and an argon atmosphere. Alloys weighing 35 lb were melted in a basic-lined induction furnace, also under argon. The compositions of the alloys are given in Table I. The nickel-free alloys were usually ferritic, while those containing nickel were generally austenitic. Depending upon the particular heat treatments, or the combination of alloying elements, two-phase alloys of ferrite and austenite could be obtained. In the following text, however, the nickel-free or nickel-bearing alloys are referred to simply as either ferritic or austenitic alloys respectively. The small ingots were swaged at 2000°F, in steps, to rods of approximately 3/16-in. diam. The large ingots were forged at 2100°F to 1-in. sq bars. No difficulties were encountered in swaging or forging any of the alloys to as much as 90 pct reduction in area. For heat treating at temperatures above 2000°F, specimens were sealed in quartz tubes under a partial pressure of helium. No special precautions were found to be necessary at lower temperatures. Intermetallic compounds in the preliminary alloys were identified by means of X-ray diffraction measurements. Correlation of these results with micro-structural characteristics (such as color, shape, size, and location) of the compounds aided in their identification. Magnetic susceptibilities were also obtained at room temperature with a Magne Gage instrument after the various heat treatments. To

Jan 1, 1962

By R. J. Hafsten, E. S. Davenport, R. A. Grange

THE influence of austenite grain size upon the hardenability of steel is now fairly well understood; for a given austenite, increasing the grain size increases the depth of hardening, and, since the latter is related to the fundamental transformation behavior of austenite, it follows that austenite decomposition must, in general, be retarded by increasing grain size. To date, very little quantitative information bearing directly on the effect of austenite grain size upon the progress and result of isothermal transformation has been published. Davenport and Bain,1 and Zimmerman, Aborn and Bain2 have shown that the transformation is definitely retarded in hypereutectoid plain carbon steels at temperatures near and above the upper region of rapid transformation (Ar') when the austenite grains are increased in size. More recently, Griffiths, Pfeil and Allen3 investigated the effect of a series of different heating (austenitizing) temperatures upon the progress of isothermal transformation at 1245°F. (675°C.) and 745°F. (395°C.) in a medium carbon, low-alloy steel; while they reported no measurements of the austenite grain size, their results suggest that increase in grain size retards austenite transformation at 1245°F. (675°C.), but has no appreciable effect at 745°F (395°C.). The present paper deals with the results of a study of the isothermal transformation of austenite in a single commercial alloy steel in which a fine-grained or a coarse-grained state was induced by appropriate selection of austenitizing temperature. The data are presented in the form of isothermal transformation diagrams, which clearly show that coarsening the grains lessens the rate of transformation of austenite to proeutectoid ferrite and pearlite-that is, at temperature levels above about 1000°F. (540°C.)-but is without significant effect at lower temperatures where the transformation product is bainite or martensite. It is to be presumed that the foregoing statement would be valid, at least in principle, for many other grades of steel; it confirms again that the mechanism of transformation to a pearlitic structure is fundamentally different from that to an acicular structure. A number of related observations were made, which seem to throw some light on the transformation process as it takes place in alloy steels, particularly at temperature levels in the vicinity of 1000°F. (540°C.), where the transition from lamellar to acicular structures occurs. These data are discussed briefly in the belief that they may lead to a clearer understanding of this aspect of the subject. MATERIAL AND EXPERIMENTAL PROCEDURE The material investigated was a commercial grade of S.A.E. 4140 steel, the composition of which was as follows: 0.37 per cent C, 0.77 per cent Mn, 0.98 per cent Cr and 0.21 per cent M0. It was hot-rolled to a 1 1/8-in. round bar and normalized by air-cooling from 1800°F. (980°C.).

Jan 1, 1941

[PAGE PREFACE. A. B. PARSONS 3 OFFICERS AND DIRECTORS 7 STANDING AND SPECIAL COMMITTEES 8 PROCEEDINGS OF 1935 MEETINGS 10 NEW YORK, FEBRUARY 10 CHICAGO, OCTOBER 14 SAN FRANCISCO, OCTOBER 14 HOUSTON, OCTOBER 15 ST. LOUIS, OCTOBER 15 ABSTRACTS OF PAPERS PUBLISHED IN 1935 17 TECHNICAL PUBLICATIONS AND CONTRIBUTIONS, 1935 64 INDEX OF 1935 PUBLICATIONS 71]

Jan 1, 1936

By KING HAMILTON GRAYSON

IRON and steel were the only basic industries in the Soviet Republic in 1928 that lagged behind the pre-war production on a comparative basis. This was due to the almost complete obliteration of all industrial units manufacturing iron and steel products during the Revolution, which resulted in the production of 1921 being only 4 per cent of that of 1913. By the end of 1927 the production of pig iron had reached a point equal to 70.5, per cent of 1913; steel was 84 per cent ; and rolled iron was 78 per cent. According to the. planned program of the Supreme Economic Council, the pre-war level of the iron and steel industry was to have been reached and exceeded during 1929-30. The production of the entire period from 1917 to 1928 in comparison with 1913, is shown in the following table

Jan 1, 1929

By E. R. Gilbert

Klopp et al. have reported data on tensile and creep properties of are-melted and electron-beam-melted tungsten. We would like to point out some similarities between their creep results and ours on are-melted and powder-metallurgy tungstenand offer some alternate interpretations of the data which lead to different rate-controlling mechanisms. Klopp et al. observed extensive transient creep in the 2250" to 3500°F temperature range where the acthetivation energy is 106,400 ± 3300 cal mole-1. They suggest cross slip as the possible rate-controlling mechanism since "cross slip may occur only by the movement of screw dislocations, and not edge dislocations, thus allowing only partial recovery of the strain hardening and resulting in a continuously decreasing creep rate, characteristic of transient creep.'' Our tests on are-melted tungsten and powder-metallurgy tungsten displayed an extensive transient creep stage in the same temperature range but they also displayed a steady-state creep stage. The stress levels and strain rates of the two investigations, 1.4 to 7 kg mm-2 and 10-7 to 10-4 sec-1 for that of Klopp et al. and 4.5 to 10 kg mm-2 and 10-8 to 10-4 sec-1 for those of Gilbert et al., are overlapping. The activation energy for creep of our powder-metallurgy tungsten is 105,000 cal mole-1. This close agreement with the value found by Klopp et al. suggests that the same mechanism may be common to both types of materials. It may be possible to gain some additional insight into the rate-controlling mechanism by applying the analysis made by conrad22 in terms of the activation volume. The value for the activation volume depends upon the dislocation mechanism and may be used to help differentiate between various dislocation mechanisms. The activation volume for cross slip is 10 to 100 b3. Values determined from our data were 130 to 220 b3 which are in better agreement with the range of 102 to 104 b3 for nonconservative motion of jogs. Values determined from data of Klopp et al. are 240 to 1040 b3. We suggest that the rate-controlling mechanism for creep in this temperature range may be nonconservative motion of jogs assisted by enhanced diffusion along grain boundaries or dislocations. This model requires an activation energy lower than that for self-diffusion as is observed. If this is the controlling mechanism, a higher activation energy which is closer to that for self-diffusion might be expected for tests conducted on specimens with reduced grain boundary area, i.e., larger grain size. In support of this premise we have conducted tests on speci- mens having one to three grains per cross section for which a higher activation energy of 130,000 cal mole-1 was observed. Also, in agreement with results of Klopp et al., a faster creep rate resulted from the increased grain size and probable decreased dislocation density. Further argument against the cross slip mechanism is based on the constant activation energy we obsserved over the range of 5.0 to 10 kg mm-2,21 whereas the activation energy for cross slip displays a significant stress dependence.23 In addition to this, the powe on the stress term for the strain-rate dependence appears too high for cross slip. The values for tungsten fall in the range of 4.5 to 7 (Klopp et al. and Ref. 21) whereas values of approximately 2 are observed in creep by cross slip for hep metals.24 At temperatures from 3300° to 4000°F, Klopp et al. observed an activation energy of 141,000 ± 4000 cal mole-' which is less than that for self-diffusion, 153,000 cal mole-1, 25 and less than that reported for high-temperature creep of powder-metallurgy tungsten by Green:' 160,000 cal mole-1. Klopp et al. suggested that recovery of strain hardening by dislocation climb is the rate-controlling mechanism in this temperature region. Actually an activation energy larger than that for self-diffusion is predicted by this model if the temperature dependence of the elastic modulus27 is taken into consideration. Klopp et al. attributed this lack of correlation to the compositional differences between powder-metallurgy tungsten used by Green and their own arc- and EB-melted tungsten. It is our position that the value of 141,000 ± 4000 cal mole-' represents a transitional value between the higher and lower activation energies and does not represent a discrete mechanism of deformation. If the tests had been extended to higher temperatures the higher value of activation energy would have been observed as has been demonstrated by both our analy sis'' of isothermal test data by GE-NMPO28 and differential temperature tests28 on are-melted tungsten. The activation energies determined (170,000 and 160,000 cal mole-1) were greater than that for self-diffusion and are in better agreement with the value found by Green. The model of dislocation climb is associated with ; activation volume of 1 b3 according to Conrad.22 The activation volume calculated for the creep data of fou independent investigators in the temperature range where the activation energy for creep is near that for self-diffusion (Klopp et al. and Refs. 26, 28, and 30) is 110 to 2800 b3 and is in the range suggested for nonconservative motion of jogs. In the case of tungsten this mechanism may be preferred over that of dislocation climb. Sincere appreciation is expressed to Messrs. J. E. Flinn and F. L. Yaggee for the helpful discussions and suggestions in the preparation of this communication.

Jan 1, 1967

By Subhas G. Malghan

Feldspathic deposits occur widely throughout the United States, but North Carolina, California, Connecticut, and South Carolina accounted for over 80% of the total domestic feldspar output for the year 1973.1 Pegmatites and granites constitute the major feldspar reserves of the United States, 2 and in addition, feldspar is produced as a byproduct by firms whose major products are spodumene and mica. 3 Feldspar flotation is practiced in the United States, Mexico, Finland, Norway, West Germany, Japan, and the USSR. According to the US Bureau of Mines estimates, the production of feldspar in the United States and rest of the world were 704,000 and 2,514,000 t, respectively. Feldspar is principally used as a flux in making glass, pottery, porcelain, enamel, tile, and other ceramic products. In recent years, the domestic feldspar industry is faced with a number of problems. As a result of increased cost of energy and the introduction in 1972 of new legislative programs relating to air, water, and noise pollution, land-use restrictions, and mined-land rehabilitation, the production costs have increased. Apart from the increased cost of operation, an operational problem exists with the feldspar producers in North Carolina, especially those in the Spruce Pine area. This problem is concerned with the use of hydrofluoric acid in the feldspar flotation. Feldspar producers in Spruce Pine, NC, have been discharging process waste water into the North Toe River. The mill waste water contains active fluoride ions. Fluoride in excessive concentrations is undesirable in waters used for drinking. It is stated that water containing 0.9-1.0 ppm fluoride will seldom cause mottled enamel on children's teeth; and for adults, concentrations less than 3-4 are not likely to cause endemic cumulative fluorosis and skeletal effect. Although the literature on this subject is rather confusing and inconclusive, the inference that concentrations over 4 ppm may affect bone structure is clear. According to the recent Federal Register,4 maximum fluoride level of 1.4 ppm at 79°-98°F is considered to be adequate and safe for the protection of health of the consumers. The EPA and state environmental agencies have accelerated their drive to reduce water pollution on a faster schedule. By 1977, the EPA specifications for maximum contaminant level of fluoride ions in the discharge waters of feldspar milling operatins is expected to be reduced to 2.0 ppm. The other problems of using hydrofluoric acid are the following: toxicity, hazards of handling and storage, and high cost as compared to other inorganic acids (almost 16 times the cost of sulfuric acid). In order to reduce fluoride levels in mill discharge waters, mining companies have taken the following steps: I) Recirculating a part of their mill water. Even though recirculation of mill water seems to be a novel method of reducing fluoride contamination of the discharge water, there are certain operational and handling problems. 2) Conducting research on the treatment of fluorides in mill waste water. The technology of treatment of fluoride ions has received considerable interest in the recent years. A recent research report5 estimates that the cost of fluoride ion removal to meet the present specifications in the feldspar industry is approximately $0.25/t of ore processed, and about $0.50/t of feldspar produced. At the request of the feldspar producing industries in North Carolina, the North Carolina State University (NCSU) Minerals Research Laboratory has taken up a research program directed towards overcoming the fluoride ion pollution problem. After a close study of the operational, technical, and economic aspects of feldspar flotation, it was decided to attempt to replace the conventional hydrofluoric acid process. Conventional Feldspar Flotatin Using Hydrofluoric Acid Since the inception of the feldspar flotation process using hydrofluoric acid by O'Meara2,6 the process has achieved a great commercial success. In the conventional flotation separation of feldspar-quartz, hydrofluoric acid is used to suppress quartz and activate feldspar, and a long chain amine salt (acetate or chloride) is used as a collector. The bulk of amine collectors used in current feldspar operations are applied as the water-soluble acetates of the free-base amines. These products are pastes or waxy solids which are available in a range of acetic acid neutralization levels generally 50 to 100% neutralized. In spite of the problems mentioned in connection with the use of hydrofluoric acid, this process is extremely stable with respect to changes in the process variables. With the exception of a few operations, alaskite and pegmatite are the major sources of feldspar.7 In the flotation treatment of a pegmatite that contains iron-bearing minerals (heavy minerals), mica, feldspar, and silica, a logical order of removal presents itself considering the following:8 1) Mica is readily floated by an amine collector in a pulp pH 3.5. Some iron-bearing minerals will usually respond to amine collector in acid circuit with the mica concentrate. 2) Since iron minerals tend to occur invariably in the feldspar concentrate, it is desirable to remove most of these prior to feldspar flotation. A fatty acid or a petroleum sulfonate collector used in this flotation step in acid circuit will float iron-bearing minerals. 3) Using hydrofluoric acid to maintain pH of 2.5, an amine collector employed in a final step will float feldspar away from silica, usually leaving the latter in the tailing as a high-grade silica concentrate. The problems connected with the flotation separation of feldspar-quartz arise due to the similarities in their chemical structure. Therefore, any reagent system that is likely to succeed in feldspar-quartz separation should have adsorption affinity towards one of the minerals and depression effect towards the other mineral.

Jan 1, 1979

By George B. Clark

COMMERCIAL strip mining of coal was first begun in the state of Illinois in 1911.1 The annual tonnage of coal produced from coal strip mines in the state was very small until 1924, when the strip mines shipped 1.68 pet of the total state shipping mine tonnage. Since then, strip mines have gradually come into their own by producing a sizable percentage of the coal produced in the state. Table 1 shows the coal tonnage shipped from both underground and strip coal mines in the state of Illinois for selected years. The table illustrates the major role that strip mines play in the state's coal industry. Today, strip mining is generally conducted in areas where the coal seams lie so close to the surface of earth that a stripping method of mining the coal seam is cheaper than an underground method. Strip mining not only recovers 100 pct of the coal deposit, but may also be used as a method of mining large acreages of coal abandoned by the closing of underground mines.

Jan 1, 1949

By B. B. Gottsberger

A NOTABLE feature of the practice of the American mining engineer is the fact that 'his field has been world wide, and the results of his work may be found in all countries. For this reason, the mining engineer is particularly interested in any legislation tending to restrict his freedom in this respect. In their zeal to improve the status of the professional engineer, those responsible for the enactment of the registration laws now in effect in various of the United States may have introduced a practice that will result very much to the detriment of the American mining engineer. He is preeminently interested in the maintenance of the professional "open door," and fame may be the reward of him who calls the first "International Conference for the Limitation of Restrictive Legislation." There is no reason why these United States should be the only country to regulate the practice of professional engineering and, in fact, the movement is spreading beyond its borders. Already, it is necessary for the mining engineer to be thoroughly conversant with the various state laws in order to insure his right to pursue his calling. It may soon be necessary for him to make sure of his status in foreign countries before undertaking work in them.

Jan 1, 1921

By AIME AIME

NO ADEQUATE salary scale, at the present time, can ignore the increase in the cost of commodities 'during the last few years or- afford to assume that this increase is merely temporary. A study of the change in prices and wages during and after the Civil War is of interest in connection with changes that have taken place from 1913 to 1919. The curves on the accompanying figure, based on reports of the War Industries Board and Department of Labor, show these changes. During the Civil War retail prices rose to 164,2 per cent. of the prices of 1860, during the next 8' or 10 years. It is interesting to note that the rise in wages during the Civil War was almost simultaneous with the rise in the prices of commodities. During the World, War, the rise in wages has lagged about a year behind the rise in prices. Furthermore, in the period immediately, following the Civil War, wages continued to rise after prices started to fall, The present price level is not considered merely temporary by such of our Government agencies as the Department of Labor and the Federal Reserve Board or by such economists as Irving Fisher' and J. S. Holden.2 Sub-

Jan 1, 1920

By James Douglas

PROBABLY no well-informed engineer would ques¬tion the accuracy of the statement that the piece of equipment that comes nearest to being in universal use in modern hydrometallurgical and ore-dressing plants is the Dorr continuous thickener. The function of the device is to separate a thin mixture of liquid and solid particles into a thick pulp and a clear liquor. Another machine in wide use is the Dorr classifier, the purpose of which is to separate the coarse and fine particles in a pulp containing solids of various sizes. It is in recognition of the invention of these two devices and the vast amount of technical work in the successful application of them in the beneficiation of ores that the James Douglas Medal for 1930 has been awarded to John Van Nostrand Dorr. To gage the importance of the development of the thickener, particularly, it is only necessary to imagine the loss of time, capacity, and metallurgical efficiency that would result if all the Dorr tanks in the world were put out of operation for a day and replaced by the old methods of intermittent settling and decantation. Truly, the record is one of "distinguished achievement in non-ferrous metallurgy" that deservedly ranks the medalist with those who have preceded

Jan 1, 1930

By George S. Rice

THE wide scope of the causes and effects of ground movement and their interrelation to various kinds of mining and geological conditions are not always understood. Minimizing of roof movement by selection of suitable mining methods is closely allied to the problem of obtaining maximum safety in mining, but there are other questions involved: for example, mechanical developments which increase the production of ore and other minerals per shift or unit of labor but which, though highly ingenious, may increase hazards. More- over, the newer mobile loaders need much space for operation with longer protecting roof spans. Another problem emphasized by mechanization is the incompleteness of extraction of mineral, as there is no governmental requirement (except for limited leases on the public domain) in this country, as in European countries, and loss of coal in pillars is accepted by mine operators rather than risking the loss of an ex- pensive loading machine from a heavy roof fall. The extent to which losses of coal in mining are now being incurred is conjectural, for it has been fifteen years since there has been a comprehensive survey. In that study a loss of slightly over one-third of the coal in both anthracite and bituminous mining was indicated.

Jan 1, 1939

By R. Dawson Hall

WHAT are the opportunities for the services of engineers in the coal mines? The best answer perhaps can be made by detailing the present lines of development in the bituminous coal mining regions. They are just at the beginning of a period that can be conservatively termed revolutionary. A few glimpses of new possibilities in economic mining that will reduce the cost of bituminous coal at the mines by at least one half have convinced almost everyone that a change must shortly come about. The cost of housing has become such a burden, especially in mountainous regions, that development which demands the hiring of more men and constructing more houses would not be profitable. The introduction of mechanical aids to mining, transportation and dumping suggests a way of obtaining a large tonnage with the same or a smaller number of men with no further house construction. At the same time the restriction of immigration has made it increasingly difficult to obtain men inured to hard labor. Modern civilization is not productive of muscular energy and at the same time it provides men who are not indisposed to the use of the mental effort required to direct the physical forces now available for the replacement of such energy.

Jan 1, 1924

By AIME AIME

The following changes of address of members have been received at the Secretary's office during the period of March 15 to May 1, 1907. This list, together with the list of changes of address published in Bi-Monthly Bulletin, No. 14, March, 1907, supplements the annual list of members corrected to Jan. 1, 1907, and brings it up to the date of May 1, 1907. The names of members who have accepted election during March and April, 1907 (new members), are printed in italics. The large number of changes of address since Jan. 1, 1907; shows the importance of publishing these changes as frequently as possible, and the Bi-Monthly Bulletin has. been selected as the means to present this information to the members of the Institute. By the simple method of cutting out these names and addresses and pasting them directly over the corresponding names in the annual list of members, the record can be kept practically up to date, and the value of the list correspondingly increased. For this purpose the changes of address have been printed only on one side of the page. The names of new members, being in italics, are readily distinguished from the others, and can be pasted in approximate alphabetical order on the margins of the pages.

May 1, 1907

By Frederick Forscher

Frederick Forscher (Nuclear Materials and Equipment Gorp.)—It is gratifying to see another series of experimental results that essentially reconfirms our reported observations of "Strain-Induced Hydrogen Embrittlement in Zirconium" (F. Forscher: AIME Trans., 1956, vol. 206, p. 536, noted as Ref. 3 of paper). However, several important considerations may have been overlooked, or—at least—have not been specifically pointed out. It is the purpose of this discussion to present some of these considerations. The statement is made that "the equilibrium concentration of hydrogen in solution at 20°C in zirconium is approximately 10 ppm; i.e., no zirconium hydride precipitates are formed at or below this hydrogen content." This is perhaps the case in the unstressed condition. In the presence of triaxial (plastic) strain however, even as little as 1.4 ppm H2 (Forscher's Figs. 6 and 11) will cause porosity near fracture in 20°C tests. The necked-down region that develops during a tensile test provides the triaxiality necessary for these pores to open. Both Beevers' and Forscher's micrographs show that porosity is concentrated in the necked region. So-called "pure zirconium" containing about 10 ppm H2 also shows pronounced porosity near fracture (Beevers' Fig. 3). Since it is generally agreed that such porosity is

Jan 1, 1965