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By P. C. Kapur

An approximate but simple and practical solution to the discretized batch grinding equation has been obtained. The solution is based on a new grind function, a combination of selection and breakage parameters and particle-size distribution of the feed. The grind function, which is readily computed from limited experimental data, has significant implications for analysis of the particle-size reduction processes. Using the practical solution, three ball-mill comminution systems—dry dolomite, wet quartz, and limestone—have been simulated with acceptable precision. The application of the discretized batch-grinding equation1 to routine comminution analysis is frustrated by two principal difficulties. First, the selection and breakage parameters of the grinding equation must be computed separately. This is usually a difficult and uncertain task although, recently, a number of experimentally and computationally efficient strategies have been proposed for determination of these parameters?" Second, for actual computation, the solution to the batch-grinding equation becomes progressively tedious and unwieldly for finer particle-size range.1,2 The purpose of this part of the report is to present an approximate solution to the batch-grinding equation, which is practical in that it seems to overcome the aforementioned obstacles. The solution involves the concepts of a new grind function, a lumped parameter of the selection-breakage functions, and particle-size distribution of the feed, which is readily computed from a limited amount of experimental data.

Jan 1, 1971

By G. R. Love, M. L. Picklesimer

Aboue 950°C the Nb-Zr system consists of a completely miscible bcc solid solution, commonly called the phase. Between 950 and 600°C, and between 20 and 85 pct Nb, the phase deconlposes, after sunciently long times, into two bcc solid solutions. The pct Zr alloys are conveniently descibecl with T-T-T (time-temperature-transformation) curves having a nose at about 2 hr at 700°C. The reaction rate varies only slowly with zirconium content and negligibly with oxygen contanzination; it is speeded up by a factor of 10 to 15 by 90 pct cold ulork and slowed dou by n factor oj 10 to 30 by a two-hundrecljold increase in grain size. Nb-r alloys with compositions between 40 and 85 pct Nb have been the basis for the majority of commercially important superconducting materials. In part because of their commercial promise, more is known about these alloys than about most other high-field superconducting materials. At the same time, there is considerable disputed or incomplete metallurgical information. For example, although Rogers and tkins&apos; indicate a monotectoid reaction at approximately 600°C and a two-phase 01 + 0, field extending between 20 and 85 pct Nb and to a maximum of 95OGC, erhout&apos; has reported that this entire region would be a single homogeneous B were it not for oxygen contamination. Again, although it has been shown that relatively short-time heat treatments in the vicinity of 700CZ significantly improve the ability of short wire samples to carry high currents in high magnetic fields at 4.2K, these observations have never been fully correlated with the structural change or changes occurring during the anneal. We intend to investigate in detail the effect of metallurgical variables, including heat treatment, on the superconducting properties of hard superconductors. To verify that our experimental techniques are valid and to establish a relative standard against which other materials may be measured, we feel it advisable to know the behavior of the Nb-Zr alloys under a variety of processing conditions. As an initial step toward this goal, we have determined in detail the kinetics of the transformations in Nb-Zr alloys. EXPERIMENT A number of problems had to be solved before beginning any fruitful work on the reaction kinetics in this system. While solving some of these problems, either by chance or by design, small amounts of information were obtained about alloys containing 40, 50, 60, 65, 67, 70, and 75 pct Nb, bal. Zr. In addition, a large range of grain sizes and a range of temperatures considerably greater than the range indicated by Rogers and Atkins phase diagram were examined. We will, however, report in detail only the results obtained for the Nb + 33 pct Zr and Nb + 25 pct Zr alloys at three grain sizes, two levels of oxygen contamination, and the temperature range 550 to 950°C. These data are most complete, but the other data are sufficiently complete to indicate the kind and magnitude of the variation of the transformation kinetics outside this range. The first and most difficult problem encountered in this inquiry was one of sample homogeneity. When Nb-Zr alloys are arc- or electron-beam-melted on a cooled copper hearth, solidification is sufficiently slow that there is appreciable coring in the cast structure and a large variation of grain size across the button thickness. Both these factors significantly affect the apparent reaction rate in the system. A two-step solution to the problem was attempted; an arc-melting and drop-casting technique has been developed by conald that greatly reduces the as-cast grain size and virtually eliminates coring segregation. Ingots made in this way exhibited no detectable (3 pct maximum) zirconium segregation. Before it was evident just how good this technique was, we attempted to supplement it with rather long-time, high-temperature annealing of the cast ingots. This annealing was carried out in evacuated and sealed (seal-off pressures < 1.0 x 106 torr) quartz capsules lined with tantalum foil at 1400 to 1450 C for 8 to 72 hr. There were two principal effects of this treatment: the grain size increased to a fairly uniform 150 p, and the surface and all grain boundaries near the surface acquired a film of a second phase, tentatively identified as an oxide (possibly additionally contaminated with silicon). There was no evidence that this 1400 C treatment had affected the zirconium segregation. High-temperature annealing was subsequently used only for grain-size control, but anneals of longer than 4 hr at temperatures greater than 1000°C were performed in dynamic vacuums (pressure no greater than 1.0 x lo torr). Any contamination resulting from these treatments was well below the limits of detection of our techniques. All samples, as cast, were cold-swaged to at least 85 pct reduction in area. The samples called cold-worked were tested as swaged. The minimum re-crystallization anneal for these alloys was about 12 hr at 1050 C; this produced an equiaxed grain diameter of about 4 to 8 P. Annealing for 4 hr at 1450°C produced a grain size of about 80 to 150 p; and annealing for 4 hr at 1650aC, close to the melting point of many of these alloys, produced a grain size of 0.5 to 1.0 mm. At all temperatures, the larger grain size was

Jan 1, 1967

By E. P. Palmatier

RECENTLY a cooling system has been in process of installation on the 3400 and 3600-ft. levels of the Magma copper mine at Superior, Ariz. The general system of ventilation employed at this inclined-vein mine is as follows: Ventilating air enters the shafts on the western side of the mine. Booster fans at each level each deliver approximately 38,000 cu. ft. per &apos;min. to the drifts. This air passes diagonally upward to the upper levels of the mine on the eastern side, where it passes up the shafts to large exhaust fans on the surface. Owing to the dip of the vein, all operations on the eastern side of the mine are above the 2500-ft. level, and those on the western side below the 2500-ft. level. At present, 2500 lineal feet of new drifts are being driven-on each of the 3400 and 3600 levels. The rock temperature is about 135°F. Until the new installation is completed, ventilating air for the two levels is supplied through Flexoid tubing from small booster fans -on the 3200-ft. level. The air picked up by these booster fans has a dry-bulb temperature of approximately 85°F. and a wet-bulb temperature of 82°F. By the time this air reaches the end of the Flexoid tubing at the 3600-ft. level, the dry-bulb rises to 101°F. and the wet-bulb to 86°F. The conditions at the head of the new drift were 102&apos; dry-bulb and 93&apos; wet-bulb.

Jan 1, 1937

By R. F. Dickerson, D. A. Vaughan, A. F. Gerds

ALTHOUGH the metallurgy of uranium has been under intensive study since the early 1940&apos;s, no systematic effort has been made to identify the non-metallic inclusions in uranium. Uranium carbide (UC), which is probably the most common inclusion found in graphite-melted metal, has been tentatively identified by previous investigators, but the other nonmetallic inclusions have received little attention. Since metallography is a valuable tool in metallurgical studies, the metallographic identification of the nonmetallic inclusions in uranium is important. Such an investigation has been completed and the identification of slag-type inclusions and of uranium monocarbide, uranium hydride, uranium dioxide, uranium monoxide, and uranium mononitride is described. Metallographic Preporation It is often possible to prepare specimens for metal-lographic examination equally well by several methods. The specimens which were examined in this work were prepared by one of two acceptable methods. For the convenience of the reader, both methods will be discussed in detail and will be referred to simply as Method I or Method II in the subsequent sections. For both Methods I and 11, specimens for microscopic examination usually were mounted either in bakelite or in Paraplex room temperature mounting plastic. Method I—Specimens were ground in a spray of water on a revolving disk covered successively with 120-, 240-, and 600-grit silicon carbide papers. It was necessary to perform the final grinding operation carefully on worn 600-grit paper to keep the scratches as fine as possible. After washing and drying, the specimens were polished for 3 to 4 min on a slow speed wheel (250 rpm) covered with a medium nap cloth. Diamet Hyprez Blue diamond polishing paste, Grade 00, 0 to 2 µ, was used as abrasive with kerosene as lubricant on the wheel. Specimens were washed thoroughly in alcohol and final polished electrolytically in an electrolyte composed of 1 part stock solution (118 g CrO, dissolved in 100 cm3 H2O) with 4 parts of glacial acetic acid. A stainless steel cathode was used. At an open circuit potential of 40 v dc, a polishing time of 2 sec retained inclusions well with the bath at room temperature. If additional etching was required to sharpen the interface between the metal and the inclusions, an electrolyte composed of 1 part stock solution (100 g CrO3 and 100 cm8 H20) and 18 parts glacial acetic acid was used at room temperature. Best results were obtained by etching for from 10 to 15 sec at 20 v dc in the open circuit. Surfaces obtained by this method are suitable for microscopic examination. However, if desired, they may be etched further with other chemicals. Method 11—Rough grinding was done on a wet 180- or 240-grit continuous grinding belt. The specimen was then ground by hand successively on 240-, 400-, and 600-grit silicon carbide papers in a stream of water. Final polishing was accomplished on a 4 in. high speed wheel (3400 rpm) covered with Forstmann&apos;s cloth. Linde B levigated alumina, suspended in a 1 volume pet chromic acid solution, was the abrasive. Specimens usually were polished in 5 min or less by this technique. Often the inclusions present in the metal were identified in the mechanically polished condition. When etching was required to outline inclusions more sharply, one of the two following methods was used. In the first method, the specimen is etched lightly while electropolishing in the chromic-acetic acid solution described above (1 part of stock solution to 4 parts of acetic acid). The electrolyte was refrigerated in a dry ice-ethyl alcohol bath and specimens were etched at 60 v dc on the open circuit for 2 or 3 cycles of 3 to 4 sec each. The second technique utilizes electrolytical etching at about 10 v dc (open circuit) in a 10 pet citric acid solution at room temperature. X-Ray Diffraction Technique The major problem in the identification of inclusions in metals by X-ray diffraction techniques is the extraction of a sufficient amount of each type of inclusion to obtain an X-ray diffraction pattern. In the present study, X-ray diffraction patterns were obtained from individual inclusions of the order of 10 µ diam. The polished and etched samples shown in the micrographs were examined at a magnification of X54 or XI00 with a binocular microscope. This allowed sufficient working distance to extract the inclusions with a needle probe for powder X-ray diffraction analysis. Friable inclusions such as MgF2, CaF2, UO2, and UH3 could be freed from the metal by probing the as-polished and etched surface. The fine particles then were picked up on the end of a Vistanex-coated glass rod (0.002 in. diam) which was held in a brass adapter made to fit the powder X-ray diffraction camera. The end of the glass rod was centered in the path of the X-ray beam. In the case of the UC, UO, and UN inclusions which are smaller in size, more metallic in appearance, and less friable than the other inclusions, it was necessary to etch the inclusion in relief before extraction. UN inclusions etched sufficiently in relief in the electrolytic polishing solution described in Methods I and II by increasing the polishing time. UN inclusions were relief etched by extending the

Jan 1, 1957

By F. S. Knoll

The expanding role of electrostatics as applied to minerals and related fields is covered in this paper. New and existing forms of electrostatic separation are classified into two broad groups: separation of charged particles and separation of neutral particles. The basis for each form of separation is discussed along with particle examples and results obtained.

Jan 1, 1986

By A. L. Bailey, B. A. Landry

All who are even generally aware of the tremendous rate of increase in coal washing operations must realize the growing importance of the float-and-sink test. I believe it is conservative to estimate that, in the past decade, the dollar volume of float-and-sink testing has increased tenfold. It is a simple matter of economy, then, to examine the factors that determine the cost of adequate float-and-sink testing. When the Coal Preparation Section of the Bureau of Mines entered upon a greatly expanded program of such work in connection with the synthetic liquid fuels investigations, it seemed advisable to examine these factors experimentally. The principal consideration that differentiates float-and-sink test sampling from general purpose sampling, is that the original particle size must be preserved. Therefore, the total cost of the test will be directly affected by any standard that might be proposed to limit sample bulk reduction at any given particle size. For this reason, the relationship of sample size to variability of results was the first factor to be studied experimentally. Of course, the matter is rendered complex by the circumstance that the float-and-sink test, not a simple analytical measurement but a process test, comprehends a number of more-or-less independent items of fundamental data; and as shown in the report, the&apos; variability of the samples differs with respect to these different items. This condition and the wide variety of situations in which float-and-sink test data are used, in combination with other factors, to study complex process operations, indicate the difficulty of setting up fixed standards for float-and-sink sampling and testing. At this stage at least, it is the intent rather to obtain experimental data on the principal Factors involved so that the reader may arrive more intelligently at a procedure adapted to his problem. The authors of this paper have presented experimental data showing the relationship between size of sample and particle size for different variability tolerances with respect to percentage of sink. In further studies, data are being collected to appraise also the variability of the samples with respect to float-ash content and size consist. The scope of this work will be broadened to cover particle sizes up to 4 in., and a third series of tests has yielded similar data for a much cleaner type of raw coal. Thus, the further studies will make available a fairly comprehensive meas- ure of variability with respect to size consist, percentage of sink, percentage of middlings, and percentage of ash in the float, for three coals ranging from 3.87 to 17.68 pct in refuse content (heavy sink material) and from 4.18 to 17.52 pct in middlings. Introduction At present there are no published standards for float-and-sink test sampling. During the rapid expansion of float-and-sink test work, varying procedures have been based on adaptations of the ASTM standards for sampling coal for analyses. For this reason, a special study of gross sample reduction has been undertaken to determine the limits for this step in the operation where no reduction in particle size is to be made before testing. Float-and-sink tests are made whenever a thorough study of coal characteristics is desired. The tests may be made on samples from coal-cleaning units such as jigs or tables, or coal samples may be tested which are taken from a loading boom, railroad car, or the coal seam itself. The resultant gross sample may be large and pose a problem of sample reduction. The question is, then, how much can the sample be reduced and still fall within preassigned limits of accuracy of the original gross sample of coal? Coal from channel samples may be crushed to liberate impurities and then separated into various gravity fractions from which washability curves are drawn; from these curves, it is possible to determine the cleaning characteristics of the coal. However, coal samples from coal-cleaning units cannot be

Jan 1, 1950

By Fordyce Coburn

RICHARD ore mine of the Colorado Fuel & Iron Corp. is an underground magnetite operation located in northern New Jersey approximately four miles northeast of the town of Dover, and 40 miles west of New York City. The mine is situated on the southeastern flank of one of the many gneissic ridges having a northeast-southwest trend

Jan 11, 1953

By Samuel M. Brock

Problems related to strip and auger coal mining have grown with the development of the industry and the economy. New technology has produced equipment that has made mining by surface methods not only highly productive, but potentially highly destructive of water resources and natural environments. On the other hand, growth of the economy has produced an increased demand for clean water and created new demands for amenity values that natural environments provide. Thus, opposing forces have created conflicts in resource use. And these conflicts have made surface mining highly controversial. Resolving the controversy will require the solving of complex economic and social problems. The essence of the economic problem created by surface mining is that, in the absence of proper regulation, substantial costs resulting from the process are borne by the public rather than the mine operator. The major goal for the regulation of the surface mining industry is to provide controls which will promote the attainment of maximum social welfare. Deciding upon the type of controls required, and how to apply them to attain a social optimum, however, poses some very difficult problems. In a comprehensive article on this subject, David B. Brooks defined the problem and suggested that benefit- cost analysis might be helpful in providing guidelines for making decisions about surface mining regulations.1

Jan 5, 1969

New York LOUIS D. HUNTOON, Chairman. ARTHUR S. DWIGHT, Vice-Chairman. THOMAS T. READ, Secretary, Woolworth Bldg., New York, N. Y. E. MALTBY SHIPP, Treasurer. GEORGE F. KUNZ, W. De L. BENEDICT. Boston ROBERT H. RICHARDS, Chairman. ALBERT SAUVEUR, Vice-Chairman. AUGUSTUS H. EUSTIS, Secretary-Treasurer, 131 State St., Boston. Mass. TIMOTHY W. SPRAGUE, HENRY A. WENTWORTH.

Jan 1, 1917

By G. L. Oldright

THE development of the reverberatory furnace for smelting copper ores up to 1912 was described by E. P. Mathewson1 with details concerning the great changes in dimensions of the furnace. Hayward2 tabulated similar data, giving developments up to about 1924, in his revision of Hofman&apos;s Metallurgy of Copper. DEVELOPMENT OF PRESENT SMELTING PRACTICE Dimensions of Furnace Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may be considered that the improvements made were due largely to increases in the size of the furnace used. Doubt-less some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By D. C. Shelton

Eq. 4 and 5 on page 5 of the March 1965 issue of Transactions, and Eq. 12 on page 8 of the same issue should read respectively:

Jan 1, 1965

By G. A. Parks, C. Kovacs

The xanthate flotation of chrysocolla can be improved by preheating the ore to 500° to 600°C. Addition of water vapor reduces the preheating temperature required. The temperature range of flotation improvement correlates well with the range over which dehydroxylation occurs. This observation is consistent with changes in the structure and properties of the mineral as detected by measurement of solubility, D.T.A., T.G.A., X-ray diffraction, and infrared absorption and with recently proposed explanations of chrysocolla flotation. A brief experiment by one of the authors several years ago showed that exposure of chrysocolla to high temperature water vapor resulted in increased recovery by subsequent xanthate flotation. Recently renewed interest in flotation and thermal properties of chrysocolla1-5 prompted further study of this effect and more detailed reporting.6 This paper explores the response of pure chrysocolla recovery in Hallimond tube flotation tests to the time and temperature of exposure and to changes in the water vapor pressure. Brief attention is paid to the effects of such treatment on accessory minerals. A single experiment with flotation of chrysocolla from an ore is reported. MATERIALS Chysocolla from Inspiration, Gila County, Ariz. was crushed in a porcelain mortar, cleaned of visible and magnetic impurities, and screened to obtain -48 +65 mesh (Tyler) material for flotation test work. The clean material contained 28.5% Cu and was identified as chrysocolla by comparison of X-ray diffraction, D.T.A., and infrared absorption data with that reported for similar materials by others.4,5,7 Samples of malachite and chalcopyrite, prepared in a similar way, were also identified by X-ray diffraction. The sources of these materials are unknown. A sample of chysocolla-bearing oxidized copper ore from Pima, Ariz. was crushed and dry ground for thermal activation and subsequent flotation. PROCEDURES Activation: The apparatus used for exposure of materials to high temperature water vapor is illustrated in Fig. 1. The furnace temperature was controlled to ±5°C Water vapor pressure was controlled by pre-saturating a nitrogen carrier stream with water in an adjustable thermostatted saturator (multiple bubbler). A water manometer and capillary orifice flow meter were used to measure nitrogen flow rates for activation and for subsequent Hallimond tube tests. Gas flow rate and temperature were adjusted prior to introduction of samples. Exposure time was measured from the moment of introduction of the sample to the furnace, hence includes an initial drop below the control temperature. After heating, the samples were removed from the furnace and stored at room temperature in air. Representative samples of activated materials were examined by X-ray diffraction and infrared absorption.

Jan 1, 1967

Kaolin deposits in Georgia lie in a belt that stretches across the central part of the state from Alabama to South Carolina. In 1977 the state&apos;s kaolin production was an estimated 4.58 million tons (5.05 million st) or 77% of US production. While mining methods vary, stripping contractor W. F. (Billy) Jackson of Wrightsville, Ga., depends on a fleet of 18 twin-engine elevating scrapers to move the hulk of overburden and clay. Owner Billy Jackson explains that the soil in this part of the state is perfect for scraper operation; overburden ranges from Georgia red clay to a sandy gumbo material. Apart from some shale, very little rock is encountered, thus eliminating any blasting re¬quirement. Most of the material can be removed with cutting edge teeth or with drop center blades.

Jan 6, 1978

By W. S. Stahmann, F. W. Bowdish

Previously published experimental data showing that both collectable and non-collectable sulfide films may be formed on chrysocolla were analysed, and chemical mechanisms were proposed for the formation of the films. A special technique was developed by which the quantity of sulfur adsorbed was determined for samples sulfidized under different conditions of pH and soluble sulfide concentration. The data on sulfur adsorption were used together with the expression for the equilibrium constant for the proposed reaction to indicate the validity of the chemical mechanism theory. A recent paper 1 about contact angle studies on chrysocolla reported the effects of pH value and sodium sulfide concentration during sulfidization upon the character of the surfaces produced. In each experiment the surface of a specimen of chrysocolla was prepared by grinding it with fine abrasive and washing it in tap water. The specimen was then placed in the sulfidizing solution for 10 min, rinsed in tap water, treated for 15 min in a solution containing 30 mg of potassium amyl xanthate per liter, rinsed in distilled water, and tested for air adhesion in a contact angle apparatus. The results, when plotted as in Fig. 1, showed an island representing conditions giving air adherent or collectable surfaces surrounded by conditions giving non-collectable surfaces. Particular interest may be attached to the region of non-collectability to the right of and above the island in Fig. 1, and to the boundry curve separating these two regions. It is apparent that at any pH from 4.0 to 6.0, where a collectable surface may be produced, a sufficient increase in the sulfide concentration results in a non-collectable surface. Thus it appears that too much soluble sulfide has a depressing effect upon the mineral, just as it has in sulfide flotation. This paper presents some speculations regarding the chemical mechanisms by which the collectable and non-collectable sulfide films may be formed and reports on a physical-chemical study intended to support these speculations. THEORETICAL CONSIDERATIONS The experimental results presented in the previous paper were analysed in several ways in an attempt to understand better the chemical mechanisms involved. For several points on the boundary to the right of the region of contact in Fig. 1, molar concentrations of H+, OH-, HS-, and S= ions, as well as various ion concentration products, were calculated. Following the reasoning of Wark and Cox, 2,3 no single kind of ions may be credited with causing depression of the surface because none of them have a constant con-

Jan 1, 1968

By H. H. Lowry

Science attracts the attention and interest of an individual or an industry in general only in proportion to the apparent direct application to its immediate welfare or benefit. Engineering accomplishments, as aphlied science, therefore receive more recognition than the scientific discoveries on which they are based. Even self-evident truth needs endless repetition before recognition: engineering as applied science cannot advance beyond the science on which it is founded. Scientific research, and discovery, is sometimes forced by the limitations recognized by those active in its development for the use of society and industry and sometimes results from an apparently superficially illogical curiosity about natural phenomena. Research has received inadequate support except in industries born in the laboratory, the chemical and electrical industries, from which many lessons might well be drawn by their elder associates. The rapid growth of these industries has been largely due to close cooperation of scientists and engineers. Most rapid advance is made when the research worker and the engineer keep in close touch with each other&apos;s ideas and problems. Many of the developments of modern industry are based on discoveries buried for years in scientific journals because the "time was not ripe" at their discovery. The situation with regard to utilization of coal is quite the oppositre. The fundamental facts in regard to the scientific aspects of the problem of developing the energy from coal are scarcely known. The engineer, without much scientific cooperation, has accomplished much in increasing the number of kilowatt-hours per pound of coal from 0.31 in 1919 to 0.69 in 1934, in increasing the 1000 gross ton-miles of freight hauled from 12.2 to 16.4 per ton of coal in the same period, in fuel economy in production of coke, and in developing the production of motor fuel in Germany and England by hydrogenation of coal. This lack of scientific research on the utilization of our most important natural resource, coal, has been due undoubtedly in large part to the reluctance of investigators to study the interrelationships involved in the complicated processes in which coal is used. These complexities are made

Jan 1, 1936

By H. H. Lowry

Science attracts the attention and interest of an individual or an industry in general only in proportion to the apparent direct application to its immediate welfare or benefit. Engineering accomplishments, as aphlied science, therefore receive more recognition than the scientific discoveries on which they are based. Even self-evident truth needs endless repetition before recognition: engineering as applied science cannot advance beyond the science on which it is founded. Scientific research, and discovery, is sometimes forced by the limitations recognized by those active in its development for the use of society and industry and sometimes results from an apparently superficially illogical curiosity about natural phenomena. Research has received inadequate support except in industries born in the laboratory, the chemical and electrical industries, from which many lessons might well be drawn by their elder associates. The rapid growth of these industries has been largely due to close cooperation of scientists and engineers. Most rapid advance is made when the research worker and the engineer keep in close touch with each other&apos;s ideas and problems. Many of the developments of modern industry are based on discoveries buried for years in scientific journals because the "time was not ripe" at their discovery. The situation with regard to utilization of coal is quite the oppositre. The fundamental facts in regard to the scientific aspects of the problem of developing the energy from coal are scarcely known. The engineer, without much scientific cooperation, has accomplished much in increasing the number of kilowatt-hours per pound of coal from 0.31 in 1919 to 0.69 in 1934, in increasing the 1000 gross ton-miles of freight hauled from 12.2 to 16.4 per ton of coal in the same period, in fuel economy in production of coke, and in developing the production of motor fuel in Germany and England by hydrogenation of coal. This lack of scientific research on the utilization of our most important natural resource, coal, has been due undoubtedly in large part to the reluctance of investigators to study the interrelationships involved in the complicated processes in which coal is used. These complexities are made

Jan 1, 1936

By E. A. Has, E. Mcl. Tittmann

Double roasting of sinter carrying a high percentage of lead concentrates, gave rise to the problem of removing the sheets of metallic lead formed in the wind boxes. The solution of the problem has been found in the water sealed wind boxes. The metallic lead is granulated and can be removed without seriously interfering with the operation of the machine and eliminates necessity of men going in wind boxes with air guns. THE practice of double sintering the charge for lead blast furnaces, which is now standard throughout the lead smelting industry, has introduced the rather difficult problem of cleaning the wind boxes on sintering machines used for the final pass. Two factors have more recently accentuated this problem: (1) the growing demand of workmen for better working conditions and less arduous labor, and (2) the growing dearth of crude lead ores and the improved grade of lead concentrates demand a charge carrying the maximum percentage of lead. One to 2 pct of such high lead charges (30 to 40) when processed on machines in reasonably good condition will pass through the grates and report in the wind boxes. One-half to three-quarters of this material will be metallic lead reduced by the coke breeze or other fuel mixed with the charge. This lead forms large cakes or chunks which are hard to remove and which substantially increase the time the machine is out of operation for cleaning the wind boxes. A high lead fall can reduce the operating time 30 pct. As a specific example, a 30 ft machine having 63 in. pallets, and equipped with three conventional wind boxes, requires 3 to 4 hr per day for removing the accumulation in the wind boxes when operated as a final pass machine. This same machine when operated on raw charge requires less than an hour per day for cleaning. To eliminate formation of large chunks of lead in the wind boxes of the final pass machines and time lost in cleaning the boxes, R. C. Rutherford of the Chihuahua smelter of the American Smelting and Refining Co., devised and patented* a wind box having a water bath to granulate the lead as it falls from the grate bars. Later, C. H. Harris of the San Luis plant made and patentedt an improvement on Rutherford&apos;s device in which the water bath seals the wind box in a manner permitting the latter to be readily cleaned without destroying the vacuum. The first machine to our knowledge to use the water bath box was at the Chihuahua plant along the lines of the Rutherford patent. In 1942, a machine was placed in operation at the El Paso plant which incorporated the features of both Rutherford and Harris. This machine is 40 ft long equipped with pallets 63 x 24 in., and having 4 wind boxes 10 ft 4 in. x 7 ft 2 in. x 51 ft 3 in. in size (fig. 1). The air from the fourth box is returned to the first three, and the gas from the first three discharged to a baghouse. The machine has conventional drive, discharge hopper,

Jan 1, 1951

By Edward Adarns Richardson

As developed up to the end of the Great War, an airplane was essentially a mechanism of wood and fabric, joined and held together by metal fittings and fastening. The engine and accessories, wire for wheel spokes and other tension members, and the occasional use of metal for inter-plane struts and the shafts of surface controls required some metal, but essentially it was a wood and fabric structure. Attempts had been made to use metal for other purposes, such as struts in the drag-trussing of wing panels and for short interplane struts, and Junkers had produced an airplane of iron in 1915. By 1917 he was using steel tubes and duralumin sheets, securing noteworthy advantages aerodynamically and from the standpoint of weight and upkeep. He was a pioneer along lines in which progress is still being made. Principles of Airplane Construction The loaded weight of airplanes ranges from 500 lb. for the light plane of which we heard so much a short time ago, to about 40,000 Ib. The great majority of planes range from 2000 to 10,000 Ib. Of this weight from 15 to 30 per cent. is for the dividend-earning load. This percentage is so low that every per cent. saved on the rest of the weight increases the pay load from 3 to 6 per cent,. The economic value of every pound that can be saved in designing or constructing a plane is sufficiently great to warrant spending money on refined methods of structural analysis, materials of superior strength properties for a given weight, methods of manufacture which may make possible the use of refinements in structural design, and the like, to an extent unknown heretofore. The weight of the airplane, without fuel, oil and crew, and pay load, ranges from 50 to 65 per cent. of the total weight loaded for planes suitable for commercial use. It is doubtful if the structural weight will ever be brought down to as low as 40 per cent. for commercial planes. The airplane may be roughly divided into: (1) Power plant, which includes engines, radiators, cooling water, oil and fuel tanks, piping, propellers, engine-starting systems, pumps, super-

By Enid C. Plante, K. L. Sutherland

Practical metallurgists are unanimous in stating that oxidation of mined sulphide ore adversely affects separation of the constituent minerals under standard conditions in a mill. Frequently, the need for different operating conditions in different plants is attributed either to the extent of oxidation of the mineral surfaces or to soluble oxidation products. The investigations reported in this paper study the effect of oxidation on the primary condition necessary for flotation, namely that of adhesion of an air bubble to the mineral. Oxidation Products from Sulphide Minerals Weinig and Carpenter1 state that the relative rates of oxidation of some sulphides are: FeAsS > FeS2 > CuFeS2 > ZnS > pbS > Cu2S tetrahedrite. They also show that oxidation is rapid at first, that sulphides of older geological formations are more stable than those of recent date and that mixtures oxidize more rapidly than any one mineral alone. The more rapid oxidation of mixtures is borne out by results of Gottschalk and Buehler2 who found that PbS, zns and cus oxidized 8 to 20 times faster when pyrite was present. Only the end products of oxidation, sulphate and the metal ion, were identified in these tests. Gaudin3 states that pyrite is difficult to oxidize in the laboratory. Winchell,4 using pyrite containing 99.98 pct FeS2, allowed water to percolate through a bed of this ground material. At the end of three months only a trace of Fe+++and So4- were found. Even after 10 months the quantity of Fe+++ was only 7.8 mg and SO4- 25.5 mg per liter of percolating solution. Car-michael5 confirmed Winchell&apos;s results on the slow oxidation of pyrite. Usually the oxidation of sulphide surfaces has been conducted in neutral or acid solutions, the acidity being caused by the oxidation products. Eliseev and Nagir-nyaka found that treatment of pure pyrite with lime gave only thiosulphate, but considered it possible that thiosulphate was obtained not by oxidation of the pyrite itself, but by oxidation of calcium disul-hide. Mitrofanov7 suggested that thiO~ulphate was formed by the following process: oxidation of sulphide to give free sulphur (this frequently happens with marcasite), followed by oxidation of the sulphur to sulphur dioxide and reaction of the sulphur dioxide with sulphur to give thiosulphate. Ferric and ferrous hydroxides, or similar compounds, were shown to be present. The Research Staff of the United Verde Copper Co.8 treated ore from the United verde mine with lime. Thiosulphate, sulphide and polysulphides were stated to be

Jan 1, 1949

By W. R. Opie

The object of this paper is to show the manner in which typical electrodeposits of hexagonal-close-packed metals—zinc, titanium, and zirconium—tend to form. The conditions of electrodeposi-tion markedly affect the nature of the deposits in regard to crystal size, orientation, and appearance. No attempt is made to correlate cell conditions with deposit quality but, rather, characteristic types of deposits were selected and a study made to show the relationship between outward appearance and crystal structure. The zinc deposits were made in aqueous electrolytes and the titanium and zirconium deposited from fused salt baths. ZINC deposits studied were made in the American Smelting and Refining Co.&apos;s Central Research Laboratories. They were representative of bright, shiny, smooth deposits, which can be obtained by using proper current densities and electrolyte addition agents, and of deposits which are dull and tend to be nodular. Four deposits were selected which ranged from smooth and bright to very nodular and dull, and their classifications are shown in Table I. These were examined microscopically and were found to differ quite markedly in outward surface structure and in microstructure. The outward surface variation is illustrated in Figs. 1 to 4. At X50 the texture could be seen to progress from granular to platelike to smooth. The microstructures of the deposits&apos; cross sections from starting sheet to outer surface are shown in Figs. 5 through 8. The dull nodular deposits (sample Nos. 1 and 2) show some tendency for acicular growth. The platelike sample (No. 3) shows tendency for columnar growth development and the smooth deposit (No. 4) has a well developed columnar structure. Back-reflection X-ray patterns of the deposit surfaces revealed some interesting information. The crystal orientation was random in all deposits except the smooth deposit. Fig. 9 is a typical back-reflection pattern of a nodular deposit. All planes are reflecting in the proper relative intensity. The pattern of the smooth deposit is shown in Fig. 10. The only planes which reflect are the (10.4). This series of planes is in the right position, with the FeKa radiation used, to satisfy the Bragg equation if the basal planes {00.1) are at right angles to the X-ray beam. The results were interpreted to mean that in the bright smooth deposit preferred orientation has occurred with the basal hexagonal planes oriented in the plane of the sheet but with all other planes randomly oriented around the c-axis.

Jan 1, 1957