Search Documents

By Robert T. Pring

In addition to the humanitarian aspects of a dust-control program, certain economic benefits are becoming more fully recognized and now furnish a greater incentive to the mill operator to eliminate the dust from his plant. It is well known that a favorable working environment promotes increased efficiency and improves the morale of employees, with a consequent decrease in labor turnover rates. One progressive mining company' estimates its direct and indirect labor turnover costs at from $40 to $50 per man; for skilled labor, such as machine operators, the cost may approach $100 per man. While these figures may not be typical of every plant, they do indicate a potential savings in operating costs. Dust signifies waste; its elimination decreases maintenance costs of mechanical and electrical equipment; its recovery is often justified from the standpoint of values contained in the collected material, although in large-scale milling operations handling low-grade ore this may not be an important factor, Perhaps the most significant trend in industrial sanitation is that dust-control equipment and procedures are becoming recognized as integral parts of the mill flowsheet. Exhaust ventilating systems are now afforded equal care in design and maintenance to that given other machinery more closely related to the Process. Methods of Dust Control "Dust control" is a general term descriptive of a number of methods used singly or together to prevent particulate atmospheric contamination. These methods are: in order .of their desirability, as follows: 1. Change of process. 2. Isolation of process. 3. Wet methods. 4. Ventilation. a. Local exhaust ventilation. b. General ventilation. 5. Personal respiratory protection. In ore-concentration plants it is seldom prac tical to change a process to eliminate the dust it creates It is possible in some cases to isolate a particular dust-producing operation from the rest of the plant, provided that constant attention on the part of the operators is not required. Frequently water is used in fine crushing and fine grinding, primarily to facilitate the processing of the ore and incidentally to allay the dust produced. Wet methods, however, are not feasible in coarse and secondary crushing operations, because of the relatively great volume of water required in instantly wetting each freshly fractured rock surface. Such quantities of water would entail conveying and storage difficulties. Local exhaust ventilation is the most effective and most commonly applied control measure, wherein directional air currents through the zone of dust production draw the contaminants into a duct system. Unfortunately, only a limited amount of reliable information on exhaust ventilation is available.2-5 of necessity these publications are rather general, being primarily concerned with engineering fundamentals. While many organizations have successfully

Jan 1, 1943

By R. G. Ward, D. R. Gaskell

Using the maximum bubble pressure technique, the densities of iron silicates at 1410°C have been measured blowing helium, nitrogen, and argon. By ensuring equilibrium between the melt and the blowing gas with respect to oxygen potential and by minimizing tempcrature cycling of the furnace, iron precipitation in the melt has been prevented. Thus the previously reported effect of blowing-gas composition on the densities of the melts has been eliminated. Consideration of the oxygen densities of the melts gives an indication of the structural changes accompanying composition change. The density-composition relationship of iron oxide-silica melts in contact with solid iron has been the subject of several investigations1-7 and considerable disparities exist among the various results obtained. Of these investigations, all except one5 have employed the maximum bubble pressure method. In the most recently reported of these investigations1 the density-composition relationship obtained blowing nitrogen differed from that obtained blowing argon. The measured densities obtained under nitrogen were greater than those obtained under argon, the difference being a maximum at the pure liquid iron oxide composition and decreasing with increasing silica content. This observation rationalized the disparities existing among the results of the earlier investigations, showing that two lines, one for nitrogen and the other for argon, could be drawn to fit all the earlier results. No explanation for this phenomenon could be offered. Chemical analysis of rapidly quenched samples of melt for dissolved nitrogen, and direct weighing measurements, excluded solution of nitrogen in the melt from being the cause of the increase in density. The range of blowing gases was extended by Ward and Hendersons who measured the density of liquid iron oxide bubbling helium, nitrogen, neon, argon, and krypton. The measured density was found to decrease smoothly with increasing atomic number of the bubbling gas. The work reported here is a continuation of the program initiated by Ward and Sachdev7 to study the densities in multicomponent melts in which the iron oxide-silica system is the solvent. As such it is necessary to explain or eliminate the anomalous densities of iron silicates under different atmospheres, and the present rede termination was carried out towards this end. EXPERIMENTAL The maximum bubble pressure method of density determination was again employed and the experimen- tal apparatus used was essentially the same as that used by Ward and Sachdev.7 A molybdenum-wound resistance furnace heated an ingot iron crucible of internal diameter 1 in. containing a 2-in. depth of melt. The bubbling gas was blown through a 1/4 -in.-diam mild steel tube onto the end of which was welded a 2-in. extension of 1/4 -in.-diam ingot iron rod, drilled out to 5/32 in., and chamfered to an angle of 45 deg. The blowing tube was introduced to the furnace through a sliding seal and its position was controlled by a vertically mounted micrometer screw which allowed the depth of immersion to be determined with an accuracy of ± 0.01 cm. A Pt/Pt-10 pct Rh thermocouple was located below the crucible and temperature control was effected initially by means of an on-off controller and later by a saturable core reactor. The bubble pressure was determined by measurement of a dibutyl phthalate manometer using a cathetometer. PREPARATION OF MATERIALS Iron oxide was produced by melting ferric oxide in an inductively heated iron crucible in air. The liquid was quenched by pouring onto an iron plate. Silica was prepared by dehydrating silicic acid at 650°C for 12 hr. RESULTS Before any measurements of the density of a melt were made, the density of distilled water at room temperature was measured bubbling helium and argon. Both gases gave the density as 1.00 ± 0.01 g per cu cm which showed that the density of the manometric fluid (dibutyl phthalate) was not affected by contact with the blowing gas. With the furnace controlled by an on-off temperature controller an attempt was made to measure the density of pure liquid iron oxide by bubbling argon. The furnace atmosphere gas and bubbling gas were dried over magnesium perchlorate and deoxidized over copper turnings at 600°C. It was found that the pressure required to blow a bubble at a given depth increased slowly with time, and thus it was impossible to obtain a unique value for the density of the melt. Inspection of the blowing tube after removal from the furnace showed that rings of dendritic iron had precipitated from the melt onto the immersed part of the tube. This is shown in Fig. l(a) where the various "steps" correspond to different depths of immersion. The precipitation of iron was considered to be due to one or both of two possible causes: i) The composition of the liquid iron oxide is that of the liquidus at the temperature under consideration and can be expressed by the equilibrium

Jan 1, 1968

By W. H. Gunniss

When the copper smelters were built, in Montana, all of the refractory products which were used in their construction were shipped in from Eastern factories. It was apparent that if a material suitable for the manufacture of these products could be found within the State, a saving of freight could be effected and a new industry built up. Prospecting for such materials was started and experiments were made with samples from numerous deposits. Considerable effort and funds were expended in an attempt to manufacture silica brick from Dillon quartz, a light colored sandstone that is found near Dillon, Mont., on the Oregon Shortline railroad. As a brick material the Dillon quartz was a failure, but it has proved very satisfactory as a silica cement. It is now being used in laying silica brick and for reverberatory furnace bottoms. A bed of quartzite that is particularly suited to the manufacture of high-grade silica brick was finally discovered at Browns Spur, 6 miles west of Anaconda. The material from this deposit is now being used in the manufacture of the Anaconda silica brick, which have gained a wide reputation for heat-resisting and lasting qualities. The analysis of an average sample of this quartz is: SiO2, 96.7; Fe2O3, 0.3; Al208, 2.5; CaO, 0.3 per cent. The very low percentage of fluxing ingredients makes this a most refractory material. All of the Anaconda silica brick are hand molded. Various types of brick machines have been tried, but the hand-mold process proved the best suited to the manufacture of silica brick from this material. The silica is first put through a jaw crusher and is then dumped into wet pans, where 4 per cent. of lime paste is added, to produce a bond. Without the lime the material would not be plastic enough to mold. After a few minutes' grinding in the wet pans the material is conveyed to the molder, who forces it into the molds by hand. The brick are then dumped on pallets and placed in the drier, where they are left. until dry enough to be set in the kilns. Some of the brick are re-pressed before being dried. They are burned about five days, in round down-draft kilns. It is not necessary to have as much heat in the kilns as the brick will be subjected to when in use in the furnaces. Montana refractories are used in smelters throughout the West. Prospecting for fire clays was also carried on. After experimenting with clays from several different deposits, it was found that a mixture of

Jan 1, 1914

* A complete magnetic picture of the Mesabi range is now available for the first time with the release of 21 additional maps based on the aeromagnetic survey of 30,000 square miles made jointly by the USGS and the Minnesota Geological Survey. A potential iron formation with a magnetic pattern remarkably like that of the Mesabi iron range lies buried beneath the glacial drift in a wide belt extending some 30 miles across the northwest corner of Itasca County in north central Minnesota. Identification of three new uranium minerals by the USGS was recently announced. The new minerals, andersonite, swartzite, and bayleyite are, respectively, hydrous uranyl carbonates of magnesium, of sodium and calcium, and of calcium and magnesium. They were first observed by Charles A. Anderson, Survey geologist, in a coating on gypsum on the 300-ft level of the Hillside mine, Yavapai County, Arizona. International Minerals and Chemical Corp. has announced a $4,000,000 expansion and development program in the Polk County phosphate fields in Florida, Plants for the manufacture of multiple superphosphate and phosphate chemicals including dicalcium p osphate for animal feed and chemical purposes, and a sulphuric acid are planned. Office building, machine shop, warehouse, and laboratory are also scheduled for erection. A new type of tramp iron detector is being used on the conveyor belt feeder at the South Agnew stripping operation. It is made of an alloy of aluminum, nickel and cobalt and does not have any electric coils. The alloy, known as Alnico, is remarkable for the permanence and intensity of the magnetism which it retains. The detector was made by Dings Magnetic Separator Co. and it is probable that Alnico will be used in magnetic separators. The American Society of Engineering Education estimates that unless the percentage of high school graduates entering engineering colleges is increased, there is a strong possibility that instead of a large surplus of engineering graduates, which has been propehsied rather widely, there will soon be fewer engineering graduates available than are needed annually by our national economy. This conclusion is reached from a survey that shows that engineering enrollments are only slightly higher than prewar whereas industrial expansion is increasing the need for engineers. * The natural resources of Korea are being investigated under ECA sponsorship. E.J. Longyear has just embarked 8 crews for that country; two drills are scheduled for base metal search while six are slated for development of coal reserves. An innovation in the development work in the iron country has been the adoption of the Drill-Mobile, jumbo-mounted drills on rubber-tired carriage, by Cleveland-Cliffs at the Mather A and B mines. The machine is being used to cut stations in the new Mathed B shaft and for driving development headings at the Mather A.

Jan 5, 1950

By Richard M. Foose

FIVE miles southwest of Mt. Holly Springs, Cumberland County, Pennsylvania, the Philadelphia Clay Co. has been mining and milling a white clay since 1896; for use in white cement, as a filler in rubber, pigment and paper, in the manufacture of insecticides, and for other minor purposes. Until late in 1942, when the mine was abandoned because of labor shortage, combined underground mining and surface stripping produced approximately 30,000 tons of clay a year, of which 10,000 tons were refined in the mill. GEOLOGY Lower Cambrian Antietam quartzite and the Montalto quartzite, stratigraphically older, trend N.55°E. and make a sharp 1600-ft. anticlinal ridge of the South Mountains. The Tomstown dolomite overlies the Antietam quartzite on the southeast flank of the anticline, which is displaced near its crest by a steeply dipping oblique fault. Erosion of the less resistant Tomstown formation has resulted in a narrow valley at an elevation of 700 ft. Directly on top of the Antietam quartzite and at the base of the Tomstown formation is a zone of sandy white clay, lenticular along the strike of the formations, ranging in thickness from 10 to 300 ft. within a mile. All of the formations dip steeply to the southeast, from 65° to vertical. The clay crops out at an elevation of about 1000 feet. ORIGIN AND CHARACTER The uppermost part of the Antietam formation is sericitic and schistose at many places, and at all these places weathering is in an advanced stage. The clay is the residual product of weathering of this sericitic quartzite or sandy schist. The crude clay has the physical aspect of a sandy kaolin, but there are large areas in the clay formation that are low in free silica, and commonly the 30 ft. adjacent to the Antietam quartzite is notably free of silica. The overlying Tomstown dolomite has been weathered almost completely to a mixed yellow and brown clay, which carries a zone of nodular and chunky limonite and manganese ore, 4 to 12 ft. thick, about l00 ft. from the contact with the white clay. In mining, this zone is an important "marker," indicating where the white clay may be expected in an advancing tunnel. The contact of the white clay formation with the yellow and brown clay of the Tomstown formation is relatively sharp, although the line of contact is irregular and therefore responsible for the lenticular character of the white clay formation. Following is an analysis of the refined clay containing less than one per cent of moisture when dried at 212°F.: Si02, 67.05 per cent; A1203, 22.27; Fe203, 0.40; CaO, 0.00; Ti02, 1.05; MgO, 0.52; Na20, 0.04; K20, 0.88; ignition loss, 7.65; total, 99.86.

Jan 1, 1944

By R. W. Brown, J. L. Loper

The process of sand erosion has been harnessed to perform a useful function—the directed perforating of oilfield tubular goods and formation rock. In this process the sand is carried by a liquid medium and the resrrlting slurry is forced through a nozzle at an extren~ely high velocity. Upon impingement against casing or formation, the fast-moving sand particles erode a perforalion in the casing followed by cavity formation in the medium behind the casing. The many factors influencing the rate and extent of this penetration are inherently difficult to establish experimentally. Equations describing submerged jet behavior have been applied to these conditions. The following paper develops the equations for predicting the performance of the perforating jet. INTRODUCTION The cutting power of high-velocity sand has been well known in the petroleum industry for many years. Sand carried by a high-velocity gas stream causes severe erosion. This effect has been the cause of numerous production problems in certain areas. It was logical to expect that this phenomenon would one day be adapted to useful industrial applications. Development of mobile, high-pressure, high-horsepower pumping equipment in the last decade has made this possible. The sand erosion process involves pumping a fluid containing abrasive solids through a set of nozzles at a high differential pressure (2,000 to 4,000 psi). The pressure conversion into kinetic energy imparts high velocity to the sand particles which, upon impact with the formation face or casing wall, erode the material in an organized pattern. This paper attempts to establish the effects of the various parameters and to predict the performance and limitations of the process. The following factors will be considered: nozzle differential pressure, perforation size, sand concentration, nozzle stand-off distance and backpressure. In general, this discussion is presented along theoretical lines and is substantiated by field results and data where reliable field data exists. POWER EQUATIONS The penetration rate of a high-velocity hydraulic jet is proportional to the power or energy of the jet at the point in quesion. The power of the submerged jet will decrease with distance (or penetration) into the cavity created by the jet. As the cavity is extended into the formation, the power of the jet diminishes to a value equal to the threshold cutting power of the particular formation which is defiined as the lowest power level at which the formation will be abraded by the jet. Obviously, when this point is reached, no further cutting of the formation can occur. It is assumed that this power level for a particular formation is a function of the hardness of that formation. The following equation can be written for the hydraulic process of formation abrasion.

By E. P. Shoub, J. P. Riott, R. C. Buehl

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By R. C. Buehl, J. P. Riott, E. P. Shoub

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By E. Miller, K. Komarek, W. Johnson

The activity of aluminum in solid Cr-A1 alloys has been measured by an isopiestic technique between Cr-A1890' and 1126" and 13 and 80 at. pct Al. The integral free energy of mixing has a minimum value of —5600 cal per g-atom at 59 at. pct Al. The maximum solid solubility of aluminum in chromium was determined to be 43 at. pct Al, and the composition limits of the compounds CrA14, Cr4A19, and Cr5Al, at 1000"~ were found to be 79 to 80, 66 to 70, and 59 to 63 at. pct Al, respectively. The thermodynamic properties of the Cr-A1 system have been investigated as part of a thermodynamic study of aluminum-transition metal systems.172 Little information is available on the equilibrium properties of the Cr-A1 system. The heats of formation of solid Cr-A1 alloys have been determined by Kubaschewski and Haymer at 600" and low-temperature specific heat data have also been obtained.~ More extensive work has been performed on the phase diagram, and a compilation has been provided by Hansen and Anderko,~ their phase diagram at elevated temperatures being essentially based on the work of Bradley and LU.~ The high-temperature portion of the phase diagram shows an intermediate phase CrA14 decomposing peritectically at 1018°C and existing at 82 at. pct A1 at 1000°C. They also identified the compounds with solubility limits of 72 to 75 at. pct A1 at 1000°C, and Cr5A1,, existing at 61 at. pct A1 at 1000°C. The maximum solid solubility of aluminum in chromium at 1000°C was found to be 46 at. pct Al. These elevated-temperature data were obtained by examination of quenched samples and were considered as less precise than the lower-temperature data. Koester, Wachtel, and Grube7 have revised the phase diagram as a result of their magnetic susceptibility and X-ray study. The results of this work differ appreciably from those of Bradley and Lu at temperatures above 800°C. The CrA1, compound is given as existing between 79 and 81 at. pct A1 at 1000°C, and they do not indicate the presence of a CrA13 phase reported by Bradley and Lu. They also report the compound Cr4Alg as having solubility limits of 66 to 70 at. pct A1 at 1000°C, while Bradley and Lu show this compound stable only up to 870°C. Koester et al. state that the high-temperature modification of the compound Cr5A18 is stable down to 1125"C, and not 980°C as stated by Bradley and Lu, and that the low-temperature modification of Cr5Al, has a range of homogeneity of 58 to 63 at. pct A1 at 1000°C. They also report that the maximum solid solubility of aluminum in chromium is 43 at. pct A1 at 1000°C. APPARATUS AND EXPERIMENTAL PROCEDURE An isopiestic method was employed which has been successfully applied to the determination of aluminum activities in solid ~e-All and Ni-Al alloys. Alloy specimens were held at different positions in a temperature gradient and were equilibrated with aluminum vapor from an aluminum reservoir kept at the temperature minimum of an impressed thermal gradient in a closed alumina system. Diffusion of aluminum into the specimens occurred until equilibrium was reached, at which the partial pressure of aluminum in each of the specimens was given by the vapor pressure of the pure aluminum reservoir. The activity of aluminum referred to liquid aluminum as the standard state in a given equilibrated sample at temperature T could therefore be expressed by: vapor pressure of pure aluminum at _ the temperature of the reservoir Vapor pressure of pure liquid aluminum, at specimen temperature T Since both the temperature of the aluminum reservoir and the specimen temperatures were determined experimentally, and the vapor pressure of pure aluminum is known as a function of temperature,' the activity of aluminum in a given aluminum alloy of known composition could be calculated. Initial runs were made with samples consisting of pure chromium chips placed in alumina crucibles. These runs exhibited large inconsistencies, indicating that equilibrium was not attained. High aluminum content Cr-A1 alloy powders were therefore substituted for the pure chromium specimens. The starting composition of the alloys was adjusted through experimentation until the concentration change necessary to attain equilibrium was small. In this manner, consistent results were obtained in reasonable times. SPECIMEN PREPARATION Alloy specimens were prepared from chromium of 99.997 pct minimum metallic purity: with 0.028 to 0.038 pct H, 0.0002 pct N, and 0.27 to 0.46 pct 0 (Aviquipo, Inc.). The aluminum had a purity of 99.99+ pct and the following impurities: 0.003 pct Cu; 0.002 pct S; 0.002 pct Fe; 0.001 pct Pb; 0.001 pct Ga (Aluminum Corp. of America). Alloy powders were prepared from weighed mixtures of chromium and aluminum by double-arc melt-

Jan 1, 1969

By Robert T. Pring

IN addition to the humanitarian aspects of a dust-control program, certain economic benefits are becoming more fully recognized and now furnish a greater incentive to the mill operator to eliminate the dust from his plant. It is well known that a favorable working environment promotes increased efficiency and improves the morale of employees, with a consequent decrease in labor turnover rates. One progressive mining company1 estimates its direct and indirect labor turnover costs at from $40 to $50 per man; for skilled labor, such as machine operators, the cost may approach $100 per man. While these figures may not be typical of every plant, they do indicate a potential savings in operating costs. Dust signifies waste; its elimination decreases maintenance costs of mechanical and electrical equipment; its recovery is often justified from the standpoint of values contained in the collected material, although in large-scale milling operations handling low-grade ore this may not be an important factor. Perhaps the most significant trend in industrial sanitation is that dust-control equipment and procedures are becoming recognized as integral parts of the mill flowsheet. Exhaust ventilating systems are now afforded equal care in design and maintenance to that given other machinery more closely related to the process. METHODS OF DUST CONTROL "Dust control" is a general term descriptive of a number of methods used singly or together to prevent particulate atmospheric contamination. These methods are, in order of their desirability, as follows: I. Change of process. 2. Isolation of process. 3. Wet methods. 4. Ventilation. a. Local exhaust ventilation. b. General ventilation. 5. Personal respiratory protection. In ore-concentration plants it is seldom practical to change a process to eliminate the dust it creates It is possible in some cases to isolate a particular dust-producing operation from the rest of the plant, provided that constant attention on the part of the operators is not required. Frequently water is used in fine crushing and fine grinding, primarily to facilitate the processing of the ore and incidentally to allay the dust produced. Wet methods, however, are not feasible in coarse and secondary crushing operations, because of the relatively great volume of water required in instantly wetting each freshly fractured rock surface. Such quantities of water would entail conveying and storage difficulties. Local exhaust ventilation is the most effective and most commonly applied control measure, wherein directional air currents through the zone of dust production draw the contaminants into a duct system. Unfortunately, only a limited amount of reliable information on exhaust ventilation is available.2-5 Of necessity these publications are rather general, being primarily concerned with engineering fundamentals. While many organizations have successfully

Jan 1, 1940

By A. L. Mincher, W. F. Sheely

Mechanical properties of columbium have been studied over the temperature range of -196 to 1093oC. The decreased strengthening influence of cold-work at temperatures below ambient has been interpreted in terms of the Peierls-Nabarro effect. Maxima in the rate of strain hardening observed during tensile testing in the range 250-600°C. have been correlated with interstitial impurities to indicate the temperature ranges at which carbon, oxygen, and nitrogen, respectively, are responsible for strain aging. THE growing need for structural materials for use above the useful service temperatures of the iron-, nickel-, or cobalt-base alloys has caused the refractory metals to be considered as potential engineering materials. These metals, which include columbium, tantalum, molybdenum, and tungsten, are called refractory because the lowest melting point among them,that of columbium, is about 1000°C higher than the average melting temperatures of conventional high-temperature alloys. They are all body-centered cubic transition metals and, as such, their mechanical properties have basic characteristics which distinguish them from the face-centered cubic metals. For example, all show a much steeper rise in strength with decreasing temperature below room temperature than do the face-centered cubic metals, and their mechanical properties are strongly influenced by interstitially dissolved impurities. In order that these new metals may be used efficiently, it is necessary that their characteristics of behavior be fully known. In this paper, the mechanical properties of columbium will be examined over a wide range of temperatures. In particular, the influences of cold-work and individual species of interstitial impurity atoms on mechanical properties will be described, and basic mechanisms which may control the observed characteristics will be explored. EXPERIMENTAL The material used in this investigation was Union Carbide Metals Co. columbium roundels consolidated to four 4-in. diam ingots, three by consumable-electrode arc melting and one ingot by electron beam melting. Impurity contents of the ingots and methods of ingot conversion and treatment are summarized in Table I. The only metallic impurity occurring in any significant quantity was tantalum at about 0.1 pct. Iron, silicon, titanium, and zirconium were each less than 0.015 pct; boron was 1 ppm or less. This should have no appreciable influence on properties. The electron beam melted material, being the purest, will be used as the basis for comparison in the discussions to follow. Tensile tests were conducted from-196 to 1093oC, on both cold-worked and fully recrystallized arc-melted and electron-beam melted columbium using standard 1/4-in. diam, 1-in. long gage length test specimens. A strain-rate of 0.005 in. per in. per min was employed until the 0.2 pct yield strength was achieved and then the strain-rate was increased to 0.05 in. per in. per min for the balance of the test. Samples were protected in an inert atmosphere at tests above 300°C. The tensile properties obtained on the electron-beam melted columbium, E, in both the cold-swaged and recrystallized conditions are given in Fig. 1. The yield strength data of Dyson, et al.,' obtained on recrystallized electron beam melted columbium and the tensile strength data reported by Tottle2 on powder metallurgy columbium are included in Fig. 1. The material used by Tottle had been purified by vacuum sintering. There is excellent agreement between Dyson's data and those obtained in the present investigation. The tensile strengths obtained by Tottle were slightly greater than those obtained in this investigation on electron-beam melted columbium but varied with temperature in a similar manner. Tottle's data showed a maximum in tensile strength near 500°C, as did our data on electron-beam melted material, and also showed a small maximum at 300°C. The significance of these maxima will become evident later in the discussion. The tensile properties of cold-swaged and recrys-tallized arc melted columbium are plotted in Fig. 2. It was found that the properties of the recrystallized arc-melted columbium from all three heats showed very close agreement except at temperatures between about 500" and 800°C. A reason for this range of disagreement will be suggested in the discussion. The generally good agreement, however, attests to the ability of cold-working and subsequent recrystal-lization to erase the effects of the three different primary breakdown procedures and to produce nearly equivalent structures in the samples derived from the three different heats. wesse13 reported tensile data on columbium having interstitial impurity contents between those of the

Jan 1, 1962

ARTICLE I Name and Object Sec. 1. This Division shall be known as the Institute of Metals Division of the American Institute of Mining and Metallurgical Engineers. Sec. 2. The object of the Division shall be to furnish a medium of cooperation between those interested in the field of physical metallurgy; that is, the nature, structure, alloying, fabrication, heat treatment, properties and uses of metals; to represent the AIME insofar as physical metallurgy is concerned, within the rights given in AIME Bylaw, Article XI, Sec. 2, and not inconsistent with the Constitution and Bylaws of the AIME; to hold meetings for the discussion of physical metallurgy; to stimulate the writing, publication, presentation and discussion of papers of high quality on physical metallurgy; to accept or reject papers for presentation before meetings of the Division. ARTICLE II Members Sec. 1. Any member of the AIME of any class and in good standing may become a member of this Division upon registering in writing a desire to do so, but without additional dues. Sec. 2. Any member not in good standing in the AIME shall forfeit his privileges in the Division. ARTICLE III Funds Sec. 1. The expenditure of the funds received by the Division shall be authorized by the Executive Committee of the Division. ARTICLE IV Meetings Sec. 1. The Division shall meet at the same time and place as the annual meeting of the AIME, and at such other times and places as may be determined by the Executive Committee subject to the approval of the Board of Directors of the AIME. Sec. 2. The annual business meeting shall be held within a few days before or after the annual business meeting of the AIME. Sec. 3. At a meeting of the Division, for which notice has been sent to the members of the Division through the regular mail or by publication in the Journal of Metals at least one month in advance, a business meeting may be convened by order of the Executive Committee and any routine business transacted not inconsistent with these Bylaws or with the Constitution or Bylaws of the AIME. Sec. 4. For the transaction of business, the presence of a quorum of not less than 25 members of the Division shall be necessary. ARTICLE V Officers and Government Sec. 1. The officers of the Division shall consist of a Chairman, a Senior Vice-Chairman, a Vice-Chair -man, a Secretary and a Treasurer. The office of Secretary and Treasurer may be combined in one person, if desired by the Executive Committee. Sec. 2. The government of the affairs of the Division shall rest in an Executive Committee, insofar as is consistent with the Bylaws of the Division and the Constitution and Bylaws of the AIME. Sec. 3. The Executive Committee shall consist of the Chairman, Senior Vice-Chairman, Vice-Chairman, past Chairman, Secretary, and nine members, all of whom shall be nominated and elected as provided hereafter in Article VII. Sec. 4. The Chairman, Senior Vice-Chairman and Vice-Chairman shall serve for one year each, or until their successors are elected. Each member of the Executive Committee shall serve three years. The Chairman shall remain a voting member of the Executive Committee for one year after his term as Chairman. Sec. 5. The Treasurer of the Division shall be invited to meet with the Executive Committee, but without ex-officio right to vote. He shall be appointed annually by the Executive Committee, from the membership of the Executive Committee or otherwise. Sec. 6. The annual term of office for officers of the Division shall start at the close of the Annual Meeting of the Institute and shall terminate at the close of the next Annual Meeting. ARTICLE VI Committees Sec. 1. There shall be standing committees as follows: Programs Committee. Finance Committee, Membership Committee, Annual Lecture Committee, Technical Publications Committee, Mathewson Gold Medal Committee, Nominating Committee, Education Committee and such other Committees as the Executive Committee may authorize. Sec. 2. It shall be the duty of the Programs Committee to secure the presentation of papers of appropriate character at meetings of the Division. Sec. 3. It shall be the duty of the Finance Committee to inquire into and examine the financial condition of the Division and to consider proper means of increasing its revenue and limiting its expenses. The Finance Committee shall audit the accounts of the Division and report to the Executive Committee prior to the Annual Meeting of the Division. It shall render a budget to the Executive Committee estimating receipts and expenses for the ensuing year so that action can be taken on same at the first meeting following the Annual Meeting.

Jan 1, 1953

By W. B. Dancy, A. Adams

Laboratory test work on heavy liquid separation of sylvite from halite is reported. Numerous tests were run on sylvite ore sized in the ranges of 4x20 mesh, 10x65 mesh, 8x100 mesh, -8 mesh and -10 mesh with heavy liquids in the range of 2.05 to 2.15 sp gr. From the test results, it was concluded that, with the type of ore under study and a size in the range of -8 mesh, a recovery as high as 90% could be achieved with a product grade of 70% KCl. However, a final product at an acceptable recovery cannot be made with one pass, and the float must either be further processed with heavy liquids or dried and sent to a conventional froth flotation circuit. Potash ores occurring in this country consist essentially of sylvite and halite plus minor amounts of magnesium sulfate salts and montmoril-lonite-type clays. Recovery of potash minerals from evaporite ores in the North American potash fields is accomplished almost exclusively by use of amine flotation. European practice involves froth flotation as well as solution-crystallization processes. Laboratory and pilot plant test work has been reported in Europe and the U. S. on the application of heavy media separation to potash ore beneficiation. Work was probably discontinued because of lack of ore with the required very coarse liberation characteristics (1/8 to 1/2 in. liberation size). Sylvite, with a gravity of 1.99, and halite, with a gravity of 2.17, appear to be ideal for separation by heavy liquids, which are now available in gravities from 1.59 to 2.95. This paper reviews preliminary results obtained from laboratory test work on heavy liquid separation of sylvite from halite. TEST WORK The heavy liquids used in the tests under discussion were chlorobromethane, with a specific gravity of 1.923, and dibromethane, with a gravity of 2.490. These liquids, completely miscible, were combined in the proportions needed to give a mixture having the desired specific gravity. Feed for the laboratory tests was mine-run ore screened to the desired mesh sizes. In conducting the tests, the sample was fed at a constant rate into a stream of heavy liquid and the mixture directed into a small separatory vessel. The float overflowed into a collecting pan while the sink collected in the bottom of the separatory vessel and was removed at the end of the test. Approximately 500 g of feed constituted a charge. Pulp density of the feed was kept low to prevent particle to particle interference in separation. With feed in the range of 8x100 mesh, a pulp density of under 10% solids by weight was found advisable. With coarser feed the pulp density could be carried as high as 15% solids. Time of separation was very rapid. In the case of 4x20-mesh material, separation was effected in 15 to 30 sec; with -10-mesh feed, separation required about 1 to 2 min. SPECIAL EQUIPMENT Since heavy liquids are toxic to varying degrees, all separatory work was carried out in a standard laboratory fume hood. It was noted that complete removal of fumes was not being effected; therefore the hood construction was modified, resulting in a completely satisfactory arrangement for heavy liquid test work. In the interest of safety, details of this fume hood are reported here. Unlike most fumes, heavy liquid fumes tend to settle and flow like water, rather than to rise like a gas. Working on this assumption, a standard water drain was installed in the hood. Across the front of the hood a 1-in. barrier was constructed. In the rear of the hood a false back was installed, with an adjustable sliding door on both the bottom and top of this panel. As shown in Fig. 1, the exhaust fan pulled a vacuum behind the barrier, sucking the heavy fumes from the bottom of the hood. Another addition was the drying box, shown to the right of the hood. This is simply a box covered on top with hardware cloth and connected by a 6-in. inlet to the hood. Sample trays made of fine mesh wire filter screens were found ideal for drying samples. With this arrangement, air flowed completely through the sample and all fumes were drawn into the hood. In use, it was found effective to cover with a

Jan 1, 1963

By R. W. Stormont, F. Wald

The phase diagram of the Pseudobinary section PbTe-Fe was determined. It was found to contain a monotectic and a eutectic reaction, the latter one taking place at 14 at. pct Fe and 875° * 5°C. The solid solubility of iron in PbTe was found to be 0.3 at. pct by electronmicroProbe analysis. No solubility of PbTe was detected in iron. Slight deviations from true pseudobinary behavior were found to occur in the range of - 5 to 10 at. pct Fe. In the course of a general investigation of reactions of various metals with lead and tin telluride,' the lead telluride-iron system was reinvestigated. It had been established much earlier than iron does not chemically react with lead telluride but forms a eutectic with a melting point of 879" The eutectic composition or other related information has never been reported, but for a number of years iron has been in general use for contacting of lead telluride and lead telluride alloys for thermoelectric applications. It seems therefore desirable to clarify the exact constitution of the system to furnish a base for the long-term evaluation of bonds made between lead telluride and iron either by pressure contacting or by brazing methods. I) EXPERIMENTAL METHODS Lead telluride-iron alloys were prepared in 10-g charges, using premelted lead telluride. This material was prepared from high-purity, semiconductor-grade lead and tellurium obtained from the American Smelting and Refining Co. and described as 99.999 pct pure. The iron used was "Armco" iron; the major impurities found here were 0.02 pct C, 0.018 pct Si, and 0.015 pct Cr. All remaining impurities were less than 0.01, the total of all impurities not exceeding 0.15 pct. Charges were prepared in closed quartz arnpoules which were evacuated and in some cases backfilled with high-purity argon to retard excessive lead telluride evaporation and deposition in slightly cooler parts of the ampoule. For high iron concentrations, this can lead to total separation of the constituents, since the vapor pressure and the sublimation rate of PbTe are quite high.4 Nevertheless, since the ampoules are closed, no change in overall composition was expected and the nominal composition of all alloys was assumed to be retained. X-ray diffraction analysis, thermal analysis, and microsections were used in the evaluation of the alloys. The nature of the system was such that X-ray diffraction was not particularly helpful. It merely served to establish that at all concentrations PbTe and a! iron were in equilibrium at room temperature. Thermal analysis was carried out by taking direct temperature vs time curves on a Sargent recorder where a width of 10 in. was kept as 1 or 0.5 mv by use of an automatic bucking voltage network. Quartz ampoules with minimized dead space, coated with boron nitride and fitted with a thermocouple reentrant, were used as containers for the charge. At high temperatures and over long periods of time, boron nitride reacts with iron. For the thermal analysis runs, however, this was not significant. More significant was the fact that the vapor pressure of PbTe at some of the meas -uring temperatures apparently exceeded I atrn quite considerably. This, in some cases, caused the slightly softened quartz tubes to blow out if great care was not taken to contain them and minimize time and temperatures used. As containers pure nickel tubes were used which also served to avoid temperature gradients in the quartz ampoule. Nevertheless, the experimental difficulties at high temperatures were severe and the monotectic temperature could therefore not be determined accurately. In general, the accuracy reached by the thermal analysis setup in this case is *4"C as determined with gold, silver, and tin, under the conditions of analysis here. Inherently, the apparatus is capable of reaching accuracies better than i 1°C. Also, difficulties were encountered in microsection-ing. They were related to polishing, since it is rather difficult to avoid pulling the iron out of the weak and brittle lead telluride matrix. It proved best to follow a procedure where, after grinding to 600 grit on carborundum paper, a polish with 6 p diamond was used on nylon cloth. Finally, #3 "Buehler" alumina and an automatic polisher were used for -5 min only, to avoid relief. The best etching results were achieved with

Jan 1, 1969

By R. M. Waterstrat

The phases occurring in the V-Mn system were studied by means of X-yay diffraction and metallo-paphic techniques, using are-melted alloy specimens annealed in the temperature range 800° to 1150°C and quenched. The bcc solid solution extends at 1250°C all the way from vanadium to 6-manganese. Below 1050°C the a-phase is formed, and the terminal a-manganese phase is stabilized up to about 900°C by vanadium in solid solution. IN the only previous general survey of the V-Mn system Cornelius, Bungardt and Schiedtl reported the existence of three intermediate phases corresponding to the approximate compositions VMn,, VMn, and V5Mn. The phase VMn8 has recently been identified as a o phase2 but the alloy VMn was found to have a bcc structure2 corresponding apparently to the vanadium solid solution rather than to the large cubic unit cell reported by Cornelius et al. 1 Subsequent work by Rostoker and Yamamoto3 has shown that the vanadium-base bcc solid solution extends to at least 15 pct Mn at 900°C. An alloy corresponding to the composition VMn, was examined by Elliott,4 who reported that the as-cast sample as well as samples annealed at 1200o and 1300°C had bcc structures, but that annealing at 1000°, 800") and 600°C produced two phases. One of these phases was apparently the bcc solid solution and the other resembled the o phase structure. Hellawell and Hume-Rothery5 established the phase relationships in manganese-rich alloys above 1000°C, and showed that the o phase in this system is replaced by the 6 Mn (bcc) solid solution at temperatures above 1050°C. These results suggest that a continuous bcc solid solution may exist above 1050°C between vanadium and 6 Mn. The present investigation was undertaken in order to develop more complete information in regard to this system. EXPERIMENTAL METHODS The alloys used in the present work were prepared by arc-melting electrolytic manganese having a minimum purity of 99.9 pct and vanadium lumps with a purity of 99.7 pct. The major impurities present in these metals were carbon, nitrogen, and oxygen and this would account for the small percentage of nonmetallic inclusions observed metal-lographically. The arc-melting was at first performed under a helium atmosphere and it was necessary to keep the melting times as short as possible in order to minimize the loss of manganese by vaporization. It was later found that the evaporation of manganese was considerably reduced when the melting was done under argon atmosphere. The final composition of each alloy was calculated by assuming that the total weight loss during melting was due to evaporation of manganese. Compositions which were calculated in this manner agreed reasonably well with the results of chemical analysis, as shown in Table I. Spectrographic analysis revealed the presence of contamination by tungsten, but in no case was the percentage of tungsten greater then 0.4 at. pct. The specimens were in each case broken in half and the fractured section was examined visually and microscopically for evidence of inhomogeneity. Each specimen was homogenized at temperatures near l100°C, as shown in Table I. After this treatment most specimens consisted of large columnar grains of the bcc vanadium solid solution. The etchant used in most of the metallographic work consisted of 20 pct nitric acid, 20 pct hydro-flouric acid, and 60 pct glycerine. It was found that this etchant would clearly delineate the phases present in these alloys although it does not produce any striking contrast between the phases. For certain manganese-rich alloys, a 1 pct aqueous solution of nitric acid was used. This etchant gave a brown color to the a-manganese phase, whereas the o phase was virtually unattacked and appeared very light as shown in Fig. 1. The etchants used by Cornelius et a1.l were found to produce spurious effects in some of these alloys. In particular, the vanadium-rich alloys etched in hot sulfuric acid often appeared to consist of two phases when both X-ray diffraction and etching with the glycerine-acid mixture indicated the presence of single phase bcc solid solution. A few percent of what appears to be an oxide or nitride phase was found at the grain boundaries and in the interior of the grains, especially in the vanadium-rich alloys. All alloys were annealed in sealed silica tubes containing 1 atm of pure argon and these tubes were then quenched in cold water. Although some manganese loss occurred during annealing, the loss seemed to be confined to the surface of the speci-

Jan 1, 1962

By R. S. Shoemaker, F. W. McQuiston

The principle of cyanide recovery or regeneration of cyanide from solution, be it a weak wash or a foul solution, is by acidulation. All or part of the cyanogen is converted thereby into hydrogen cyanide which is fixed by an alkali (generally lime) and returned to the cyaniding system (the Mills- Crowe process). The solution is made acid by bringing it into contact with sulfur dioxide. The acidified solution is stripped in a packed tower by a counter flow stream of air in which the air and the solution are brought into intimate contact. The air leaving the tank charged with hydrogen cyanide is then passed to another tank in which it is mixed with an alkaline solution, the latter absorbing the HCN and leaving the accompanying air clean for reuse in removing more HCN from the acidified solutions. Of the recoverable cyanide, 9770 is usually recovered in a three-unit disperser tower system. When there is a considerable content of copper, the solution is made acid following standard procedure; however, the HCN is not removed until the copper is reduced by regeneration. Copper removal at this point may be as high as 90%. After reduction of the copper, the acidified solution then is contacted with a lime solution which brings down calcium sulfate as a second step. The silver, gold, and copper are recovered in the form of a precipitate which can be shipped to a smelter. At the Hudson Bay Mining and Smelting Company, Ltd., most of the precious metals have been recovered in the copper and lead concentrates, but it has been found profitable to submit the flotation tailings to a scavenger cyanide treatment for the recovery of exposed gold contained in pyrite rejected during flotation. Due to the fine particle size, maximum dissolution is achieved in a four-hour contact period. In the cyanidation plant, the thickened flotation tailing at 70% solids is given a four-hour agitation in weak calcium cyanide solution and filtered. Filter cake is repulped and discharged either to the tailing storage area or to the mine as backfill. Clarification, de-aeration and precipitation of pregnant solutions are carried out in a standard Merrill-Crowe unit. Typical analysis of precipitate is as follows: Au, 0.47%, Ag, 4. 88%; Cu, 46.0%; Zn, 13. 8%. Cyanicides in the form of copper minerals limit the reuse of barren solution. To recover the substantial quantities of cyanogen chemically combined with base metals in this solution, a cyanide regeneration plant is in operation. The barren solution is acidified with sulfuric acid in a spray tower where an upward current of air from a blower sweeps the liberated hydrocyanic acid gas to an adsorption unit where it meets a fine rain of lime suspended in plant solution with which it combines to form calcium cyanide. The lime--water solution carrying calcium cyanide is reused in the cyanide circuit. The return gases are drawn into the blower and back to the disperser tower thus

Jan 1, 1975

1. What Is a Rock Burst? . . . . 569 Jack Spalding . . 569 A. F. Robertson . 569 W. R. Crane . 569 G. S. Rice......569 P. J. Shenon........... 569 L. C. Graton........... 569 2. What Are the Causes of Rock Bursts? 573 Jack Spalding..........573 A. B. Yates and P. J. Shenon 574, 578, 592, 596 A. F. Robertson . . . . . 575, 595 W. R. Crane. ... 579 G. S. Rice.........582 Oscar Weiss .........582 P. B. Bucky..........582 P. H. Bird.....588, 592, 593, 595 M. K. Hubbert.........593 G. Wheeler.......... .593 B. F. Tillson........593, 595 T. T. Read............594 3. How Can Rock Bursts Be Prevented?. 597 Jack Spalding..........597 A. B. Yates and P. J. Shenon ... 598 A. F. Robertson.......599 W. R. Crane....... .602 G. S. Rice............602 Page 4. How Can Rock Bursts Be Predicted?. 602 Jack Spalding.........602 A. F. Robertson.......602, 603 A. B. Yates and P. J. Shenon ... 602 W. R. Crane......602 G. S. Rice......602 B. F. Tillson...........603 5. How Can Rock Bursts Be Controlled? 603 Jack Spalding..........603 A. F. Robertson.......603 W. R. Crane...........604 A. B. Yates and P. J. Shenon 604 Oscar Weiss........605 G. S. Rice......- 604, 609 R.D.Hoffman..609, 610 R. D. Hall..........609 R. H. Cromwell . . 609 P. J. Shenon...........609 P. B. Bucky......609, 611 B. F. Tillson.........609, 611 A. B. Yates...610, 611 P. H. Bird.......610 6. General Discussion........ 611 Jack Spalding.......... 611 R. D. Parker....... . 616 C. Baring Horwood........ 618 P. B. Bucky........... 619 Foreword This symposium brings together points of view, experiences and ideas on rock bursts from a goodly portion of the globe. Some conceptions, particularly those regarding the necessity of depth, have been fairly well shattered, as we find that they are present in the water tunnels of New York, on the surface at marble quarries in Vermont, and in the coal mines of the United States, as well as in the deep mines of the South African and Kolar gold fields, and in the nickel, copper and gold mines of Canada. Their occurrence under such a variety of conditions is, therefore, a challenge, which, if successfully met and overcome,

Jan 1, 1946

By J. A. B. Forster

The operations of the Electrolytic Zinc Co. of Australasia Ltd. involve the preliminary roasting of zinc concentrate from Broken Hill, New South wales, at a number of acid-making centers on the Australian mainland. The partially roasted material is shipped to the Tasmanian plant where it was formerly re-roasted prior to leaching,' but this practice was gradually abandoned in favor of the flotation of leach residue for the recovery of a concentrate containing about 22 pct sulphide sulphur and known as secontlary or Risdon concentrate.2 This change, together with increased zinc production, called for new furnaces at Risdon for the roasting of 10-50 tonst of secondary concentrcite and 90 tons of new concentrate per day. The war situation made it imperative that construction be commenced without delay and to save time it was decided to base the design on that of Skinner type furnaces which had been built in South Australia twenty years earlier. This involved some sacrifice of desirable features but did not prevent the incorporation of provisions which had enabled small Herreshoff type furnaces to roast secondary concentrate autogenously. The roasting rate in existing furnaces was 10-12 Ib sulphide sulphur per sq ft of hearth area per day, and eleven hearths, 20 ft in diam, were provided in the two new furnaces in anticipation that they would prove capable of the autogenous roasting of 130-140 tons of concentrate per day from 28-29 pct sulphur down to 4.5-5 pct, with an oxidation rate of 11-11.5 lh S/S per sq ft of hearth area per day. It was soon realized, when the first of the furnaces went into operation, that a much better performance than this might be given, and the purpose of this paper is to show how a treatment rate of 100 tons per furnace day, and oxidation rates approaching 15 lb S/S per sq ft per day have been reached. Furnace Design General specifications for the design were based on the results of experience with several other types of furnace and on the special operating conditions

Jan 1, 1950

By B. H. Bergstrom, D. D. Crabtree, C. L. Sollenberger

Experimental data is given for several natural minerals on the load required for the fracture of single specimens over a very wide size range. The experimental results are compared with published data, such as for coal specimens from 30 to 64 in. in size. Correlations of the specific energy (per unit mass) required for fracture with specimen size are given as functions of specimen shape and the experimentally observed load-size relationships. These correlations are discussed with reference to comminution theory. Comminution, through compression or impact, can result only if sufficient force, or energy, is applied to a specimen of a given size. The relation of specimen size to force requirements has not been adequately considered previously. For single particle fracture, it was previously shown 1 that the specific energy (energy, e, per unit mass, M) at fracture of single glass spheres was inversely proportional to the product size modulus, k, as indicated in the equation: e/M = C/k [1] This was found true not only for numerous glass specimens of the same nominal size but was true for all nominal sizes tested from 1/8 to 1 in. Also, it was observed that the smaller spheres required a greater specific energy at fracture than the larger spheres and consequently had a smaller product size modulus. However, the relationship giving the required fracture energy as a function of specimen size was not further elaborated upon. This paper will extend the development of the feed size effect on specific fracture energy for brittle materials. SPECIFIC ENERGY-FEED SlZE RELATIONSHIPS FROM PREVIOUS DATA For the particular glass spheres tested in the previously described experiments,' it can be shown, as on Fig. 1, that the specific energy varied inversely as the 4/3 power approximately of the feed size, F, in the span of nominal sizes ranging from 1/8 to 1 in. The load at which a 3-in. diam glass sphere (a crystal gazer's ball) was broken was within the range predicted by an extrapolation of this energy-feed size relationship. The product size distribution from this

Jan 1, 1963

By C. A. Rowland, R. C. Nealey

Until recently there were very few dry grinding rod mills used. The early installations performed as anticipated. In the last seven years a number of dry grinding rod mills have been made, most of which have been used for grinding iron ore. The recent installations have not operated as anticipated in accordance with practices learned from wet rod milling. The paper does not establish any definite rules for the application of dry grinding rod mills but does attempt to establish some guidelines that can be followed. The paper reviews a number of flowsheets describing the problems encountered and some of the solutions found. Completely dried compared to partially dried feed is reviewed. Problems encountered in scaling up from pilot plant to production-size mills are described. Rod sizing, effect of rod length, rod bending, rod breakage, and rod tangling are discussed. While wet grinding rod mills have been used for 50 years to grind ores and make specifications sands, dry grinding rod mills have only recently come into significant use in the mineral industries. Among the materials now being processed in dry grinding rod mills are (1) lead sinter, lead slag, and limestone; (2) iron ore; (3) silica; (4) cement clinker; and (5) coke. Because the application of wet grinding rod mills was well understood at the time that dry grinding rod mills were introduced, the criteria and guidelines for selecting the former were carried over to the selection of the latter - in some cases with the anticipated results, but in many cases not. It is the purpose of this paper to describe, through a series of flowsheets, some of the recent problems encountered with dry grinding rod mills, to point out the solutions to these problems, and to set forth some guidelines for the selection and application of these mills. When dry grinding rod mills came into use, the following facts about their wet grinding counterparts were already well established: (1) The performance of large-sized mills can be accurately predicted from the performance of rod mills 5 or 6 ft in diameter. Wet grinding mills as large as 14 ft in diameter (inside shell) by 20 ft long with rods 19.5 ft long can operate satisfactorily.

Jan 1, 1970