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By D. H. Gieskieng

THE advent of curved jaw crusher wearing plates made an approach other than segmental layout analysis desirable for prediction of capacities. For some time it had been known that the drawing board capacities of crushers using these plates had to be considerably modified by complicated experience factors to achieve agreement with results. Because these apparent capacities could be readily increased severalfold by minor crushing chamber shape changes, it was necessary that the utmost precaution be taken in predicting capacities of jaw plates modified for nonchoking, special wear characteristics, or any other reason. To this end the laboratory and field tests outlined by the author in a previous paper1 were made on Blake-type jaw crushers. The results of these tests were summarized in a simple first degree equation applicable to crushers using either straight or curved jaw plates. This equation first outlines the maximum capacity potential of a given crusher, then reduces this figure in accordance with installation circumstances by means of a realization factor. It was found subsequently that this equation, with the addition of an eccentric throw factor, is applicable to standard types of single-toggle or overhead eccentric jaw crushers as far as maximum capacity potential is concerned. However, these crushers have realization factor curves somewhat different from those outlined for the Blake type. While this paper is concerned principally with standard type single-toggle crusher capacities, the evaluation of data obtained with these machines is simplified by comparative reduction to the 10 x 7 in. Blake-type equivalents upon which the summary of the preceding paper was made. Convertibility of data from one type of crusher to the other also tends towards confirmation of both. The agreement of these data is sufficient to be considered complimentary. Consequently the feed factors, f, previously reported for Blake crushers are slightly adjusted to an average with the single-toggle crusher results. Blake-type equation: C = f.d.w.y.t.n.a.r [I] Single-toggle type equation: C = f.d.w.y.t.n.a.e.r [21 where C is the capacity in short tons per hour through the crusher, f is a feed factor, dependent upon the presence of fines in the feed, and the surface character of the jaw plates used. Values of f : Smooth Plates Corrugated Plates With normal fines 0.0000414 0.0000319 Fines scalped out 0.0000368 0.0000252 Large pieces only . 0.0000312 0.0000215 d is the apparent density of the broken product in pounds per cubic foot. (If the true specific gravity of the feed is known, 40 pct voids may be assumed and d becomes 37.4 times sp gr). w is the width of crushing chamber in inches. y is the openside setting of the crusher, in inches. In the case of corrugated jaw plates it is measured from the tip of one corrugation to the bottom of the valley opposite. t is the length of jaw stroke in inches at the bottom of the crushing chamber. It is the difference between open and close-side settings. n is rpm, or crushing cycles per minute, a is the nip-angle factor. It is unity for 26" and 3 pct greater for each less nip-angle degree. A nip-angle of 20" has an a value of 1.18, and an angle of 30" has an a value of 0.88, see Fig. 1. r is the realization factor. It is unity for perfectly uniform choke feeding and usually less for actual operating conditions according to the method of feeding used and the probabilities of hang-ups involving the size of feed and crusher opening. Approximate values are given by the curves in Fig. 2. These values are further reduced by intermittent feeding, e is the throw or diameter of gyration of the single-toggle crusher eccentric in inches. As evident in Fig. 1A, variation of feed size will generally have little effect on nip-angle if both jaw plates have flat areas. Jaw plates having continuous curvature, as in Fig. 1B will have different nip-angles, depending upon the size of feed. For test work as described in this paper this effect was accounted. For general compilation of capacities for average feeds it is suggested that the nip-angle be taken at the various settings computed, at an arbitrary level, such as is indicated in Fig. 1C. Data Evaluation To bring the Blake and single-toggle type crusher capacity test results to common terms for evaluation, all data are converted to terms of 10 x 7 in. Blake-type performance at conditions of 100 lb per cu ft, 10 in. chamber width, 250 rpm, 0.65 in. stroke, 3-in. openside setting, and 18° nip-angle. (The nip-angle of the 10x7 in. Blake is 18° at 3-in. setting.) The single-toggle crusher performances are also divided by the eccentric throw to bring this effect to unity. As outlined,' laboratory and field tests made on Blake-type crushers ranging from 10 x 7 in. to 60 x 48 in. were summarized along the foregoing conditions of speed, stroke, etc. This resulted in groups of data which correspond to feeds with fines, feeds

Jan 1, 1952

By M. M. Leighton

THE engineers of Illinois have been submitting to the State Geological Survey an increasing number of requests for advice on their geological problems, including landslides, unequal settling of fills, laying pavements across peat bogs, shattering and heaving of pavement by frost action, etc. The authors discussed the behavior of landslides, their effects, their causes, a tentative classification, and remedial measures. In the latter connection the efficacy of drainage and how it should be handled in order to be most effective was considered. The principles and efficacy of other remedial measures, such as subsurface blasting, rock anchoring, installation of wood, concrete or steel piling, construction of retaining walls, construction of cribbing and reduction of slope were also discussed. Attention was also given to the matter of applying geological knowledge to avoid undesirable highway locations, and a specific example was cited.

Jan 1, 1929

By C. D. King

The advances in basic open-hearth practice which have occurred during the past decade have been principally in the direction of the physical development of larger units, decreased fuel consumption and increased length of furnace campaigns. However gratifying the economies which have resulted from these improvements are, it must be granted that like advances in chemical, physical-chemical and metallurgical knowledge pertaining to open-hearth practice would have added greatly to these already attained reductions in costs and it is the purpose of this paper to stimulate the discussion and study of these phases of cooperation and their relation to ingot yields. Definition and Calculation oF Losses In discussing basic open-hearth yields, the writer refers to the ingot yields obtained in the process of converting pig iron, scrap and ore to steel. In this conversion some scrap is produced. The difference between the total weight of ingots and scrap produced and the original metallic charge constitutes the loss. Basic open-hearth yields vary considerably, and while some of the divergencies can be explained metal-lurgically because of differences in slag analysis, proportion of pig charges, and character of steel made, there often remain differences that apparently do not readily disclose themselves. In reconciling yields by the usual accepted methods of calculation, one is often forced to evaluate large losses as ((unaccounted for." Such a system is prone to accept these large losses as inherent to the operat.ion and not, susceptible of appraisement and correction. It is obvious that if the ingot yield plus the scrap recovered subtracted from 100 per cent. determines the loss, a drop in yield may increase the loss or the scrap, and both contribute to a monetary loss. To be sure, one can subdivide these unaccounted for losses into various components, such as metallics in slag, free Fe in slag, spillage, etc., and thereby reduce the amount of losses considered as "unaccounted for." If empirical methods could indicate with some degree of accuracy all losses, however slight, the undetermined losses become appreciably smaller and resultant calculations will check actual

Jan 1, 1929

Richard Moldenke, New Pork, N. P. (communication to the Secretary*):—Iu following the discussion of the specifications for cast-iron and finished castings, I mas strongly impressed with two points which might be ventilated to advantage more fully from the view of the practical foundry-man. First, the reasons why only silicon and sulphur were specified and not phosphorus, manganese and carbon; and second, why but little was said concerning the presence of iron oxide in the material.

Jan 1, 1905

By George W. Small

The ore, as it is raised from the mine, has all average assay-value of $11 per ton, chiefly in the form of free gold. All the ore goes directly to the stamp-mills, of which there are two. The older and larger mill contains sixteen batteries of five stamps each, with one Frue vanner to each battery. The new mill has eight batteries of five stamps and two Frues to each battery. The large mill is driven by Leffel turbine wheels, with a pressure of eighty feet, and a consumption of 600 miners' inches of water. The smaller mill is driven by "hurdy-gurdy" wheels with a pressure of about 550 feet and a consumption of 150 inches of water. At both mills the tailings from the stamps pass over about 20 feet of plates on their way to the Frues. In each set of plates the first or upper one is copper, the rest are so-called silver plates. The bullion from the stamps is about 800 fine in gold and 200 in silver. The concentrates from the Frues average from 11/4 to 1 ; per cent. of the ore stamped. They very rarely exceed 2 per cent. I was unable to get the exact assay-value of the concentrates, but it is said to vary between $100 and $200 per ton. The concentrates are treated at the chlorination works at the rate of one hundred tons per month. The capacity of the works is somewhat greater than this, but as the supply of concentrates is limited it is not deemed advisable to work them up any faster. Care is taken to keep the concentrates always damp until they are put into the roasting furnace. If this is not done, a decomposition of the pyrites begins, forming lumps which do not roast, and which consequently cause a loss of gold in the residues from leaching. A Fortschaufelungsofen is used for roasting. Its dimensions, incloding fire-box, are 12' X 80'. The hearth is one continuous plane, but the charges, of which there are three in the furnace at one time, are kept entirely separate. The furnace-men call the three compartments, the "drying," the " burning," and the " cooking" compartments. In the middle, or "burning" compartment, the ore is spread

Jan 1, 1887

Bailey, Lewis Newton, Master Engineer, Senior Grade, 4th Regiment, U. S. Engineers, Headquarters Company, died of pneumonia at Camp Merritt, N. J., on April 30, 1918. Baird, Louis, Lieut., Royal Field Artillery, British Army, died on the battlefield in 1915. Ballamy, John H., Capt., 103c1 Engineers, killed in action near Fismes, Aug. 9, 1918. Bowles, Martin F., 2d Lieut., Co. B, 355th Infantry, killed in action, Sept. 3, 1918. Burt, Andrew, died in active service, 1916. Cobeldick, William Morley, Royal Engineers, died from gas poisoning on Oct. 7, 1915. Dougall, Ralph,. 4th University Co., Princess Patricia Regiment, killed in action early in the war. Evans, Alfred Winter, Lieut.-Col., New Zealand Rifle Brigade, D. S. 0., D. C. M., killed in action on Oct. 12, 1917. Gordy, Sheppard B., died in service, Oct. 9, 1918. Gorman, Thomas C., Lieut., Canadian Engineers, killed in France, Mar. 18, 1918. Hague, William, 1st Lieut., Engineer Officers' Reserve Corps, died in active service, Jan. 1, 1918.

Jan 1, 1919

By G. S. Rice

Some form of longwall mining is generally used in Continental Europe; also in Great Britain where the coal is weak and friable, or the coal bed provides material for pack walls and filling, or where the bottom is soft and squeezes up easily, or the roof is pliable, or the bed is thin and brushing provides building material, or the thick multiple seams are mined in layers, such as the 24-ft.seam at Weymss, Scotland, or the 10-yd. pitching bed near Coventry, England. Pillar methods have been retained in British fields where the coal beds are from 5 to 9 ft. (1.5 to 2.7 m.) thick, and free from thick partings or binders or without draw slate, which would provide waste rock for pack walls, or where the roof is hard and requires shooting to bring down, and also the bottom or floor is comparatively hard. The room-and-pillar system, by which probably 95 per cent. of the coal of the United States is produced, is now found in only a few mines in Wales, where it is known as the pillar and (single) stall method. The American room-and-pillar system is not equivalent to the bord-and-pillar or stoop-and-room system. The essential difference between the bord-and-pillar and the average American room-and-pillar system is probably due to natural conditions. The majority of the coal mines in the United States have a depth of less than 700 ft. (213 m.). The bulk of the coal mined in Europe comes from a depth of more than 1200 ft., while the deepest mines in Great Britain, Belgium, and France reach 3500 to 4000 ft. In deep workings wide spans are impracticable, even with the use of abundant timber; accordingly, as the mines in the United States become deeper the rooms will become narrower and the pillars thicker, and the bulk of the coal will be extracted on the retreat. It is generally conceded, without reference to the cost of production, that the larger the pillars left on the first mining, the more thoroughly can the coal be extracted. In Europe complete extraction is generally compelled either by the lessors or by the governments; whereas in this country it has been more largely a question of competitive cost of production, or the support of the surface in the flat farming districts of the

Jan 1, 1922

By Frank L. Antisell

Certain physical and chemical properties of copper are so intimately related that a change in variation of the physical properties indicates a certain chemical change. The standard specifications of copper of the American Society for Testing Materials are based upon physical characteristics with one exception, namely, the copper contents. The physical standard includes electrical properties, defects in the "set or casting," variation in dimensions, weight, etc. Paragraph 4 of the specifications states, " Wire bars, cakes, slabs and billets shall be substantially free from shrink holes, cold sets, pits, sloppy edges, concave tops, and similar defects in set or casting." Shrink holes and concave tops indicate both physical and metallurgical defects; cold sets, sloppy edges, etc., indicate physical defects only; small pit holes may be indicative of either or both physical and metallurgical defects, depending on conditions under which they originated. To the refiner, a copper casting, no matter how made, is known as a "shape," irrespective of whether it is wire bar, ingot, cake, slab, billet, etc. As practically all shapes are cast in an open mold, that is a mold without cope, the surface of the shape cools in contact with the atmosphere and has a coating of oxide. Certain phenomena of the surface indicate the adaptability of the metal for forging or rolling. The writer believes that it is much simpler to produce shapes that are metallurgically correct than those that are mechanically satisfactory. When copper is to be rolled into thin sheets or drawn into very fine wire, the slightest physical defect becomes evident. It is, therefore, of the utmost importance to consider every characteristic of the shape and ascertain the physical and chemical relationship. Pitch and Set.—The appearance of the surface of the copper when cooled is considered as to its "pitch," or the general contour of the surface of the shape (which may vary from a concave to a convex); and the '" set,"' or the detail appearance of the surface of the shape (the wave-like structure). When the shape presents a concave surface, it is known as "low pitch;" a convex surface is known as "high pitch" or "tough pitch;" and a level surface is called "flat pitch." The set of the copper is, in a very large measure, directly related to the pitch; and so intimately are

Jan 1, 1921

By T. G. Decker, A. F. Witt, A. M. Gaudin

The historical development of the concept of contact angle hysteresis is reviewed. The measurements of contact angles reported in literature have all been made under static conditions. For the measurement of dynamic advancing and receding contact angles, four methods are described together with an evaluation of their precision. Three of these methods were adapted from devices already used for the measurement of static contact angles. The fourth method uses a new principle and apparatus and is the most accurate. While the relationship between surface tensions and the contact angle was extablished as early as 1805 by T. Young,' it was only in 1890 that the existence of contact angle hysteresis was reported by Lord Rayleigh. His observations, however, remained almost unnoticed. With the invention of froth flotation, the contact angle and its behavior became the subject of investigation. In 1907, L. Sulman and collaborators investigated the cause of discrepancies between recorded values of contact angles. They found that certain variations existed in most cases, but that definite maximum and minimum values of contact angles are obtainable. The difference between the maximum and the minimum contact angle in a given system was termed "hysteresis" by Sulman. First attempts to explain hysteresis were made by E. Edser. In the following years extensive studies of contact angles in various three-phase systems

Jan 1, 1963

By A. M. Gaudin, C. S. Chang

IN general, fatty acids or their alkali salts do not cause clean quartz to float. However, the presence of multivalent ions, such as the alkaline-earth metal ions and heavy-metal ions, can activate quartz and thus make it respond to soap flotation. The mechanism of activation and soap flotation of quartz has been studied.1-4 The problem was attacked by making flotation tests rather than by careful adsorption measurements on suspensions of ground quartz in solutions containing the activator and the collector. Attempts have been made to determine the adsorption of barium ion and calcium ion on quartz. The results have only qualitative value as the analytical methods were not adequately precise. Artificial radioactivity and progress in radioactivity detection have introduced a new analytical tool, which can have a sensitivity several orders of magnitude greater than that of other methods.5-6 In the experimental work described in this paper the radioactive tracers, barium 140 and carbon 14, were employed. Lauric acid was synthesized, having a radioactive carbon atom and was used with Ba140 to study the adsorption of barium and laurate ions from aqueous solution on ground quartz. Flotation tests, using the vacuum flotation technique,4 were run in conjunction with the adsorption experiments. The object of this investigation was to understand better the reaction mechanism of barium activation and soap flotation of quartz. Methods and Materials Quartz Sample Preparation: Lumps of hand picked quartz, washed free of stains, were reduced in size to -20 mesh in laboratory crushers. The product was boiled in concentrated hydrochloric acid to remove the abraded iron. The cleaned product then was ground dry in a laboratory porcelain mill, using siliceous pebbles, to pass through a 200 mesh screen. The ground product was deslimed by sedimentation at 37 microns. This -200 mesh +400 mesh material was subjected again to boiling with concentrated hydrochloric acid, followed by washing with distilled water and further desliming. The cycle was repeated until no trace of iron could be detected in the wash liquor. During the final washing, the pH of the wash water was checked to insure the complete removal of the acid from the mineral. The quartz then' was rinsed with conductivity water. The resulting product was dried in a drying oven and stored in mason jars. The impurities (spectrographic) were as follows: Al, 0.01 to 0.001 pct; Li, Mg, Mn,, Na, each less than 0.001 pct; Fe and Cu not detected. The specific surface of the quartz sample was determined by the krypton gas adsorption method.7 The average of two measurements was 881 sq cm per g. The calculated specific surface for 37/74 micron quartz particles, using a surface factor of 1.96,8 is 900 sq cm per g. This is in good agreement with the measured value of 881 sq cm per g. Beta-Counting Method: The radioactive isotope of barium used in the experimental work was barium 140. It has a sufficiently long half-life, 12.8 days, emits high-energy beta particles (up to 1.0

Jan 1, 1952

By H. O. Hofman

Lime-roasting is a term proposed by Ingalls' for the operation of forcing air under pressure through a mixture of galena and lime at the kindling-temperature with the object of oxidizing lead and sulphur and of fritting or fusing the charge. If finely-divided galena were treated in this manner without the addition of lime, the heat set free by the oxidation of part of the lead and the sulphur would be sufficiently great to fuse un-decomposed sulphide, and thus stop desulphurization. Besides the chemical action that the addition of lime, limestone or gypsum to the charge may have, the admixture has the physical effect that it keeps the particles of galena separated from one another and accessible to the oxidizing effect of the air. At present, three methods of lime-roasting are carried out on a working-scale,—the Huntington-Heberlein, the Carmichael-Bradford and the Savelsberg2 processes. In the last, which interests us here, an 8-ton charge is made up of galena, limestone and perhaps some siliceous or ferruginous flux; the whole is crushed to pass a screen with 3-mm. holes and moistened with 5 per cent. of water. It is fed gradually into a bowl-shaped converter, 6.56 ft. in diameter, supported by trunnions attached to a truck. On the bottom the converter has a grate with blast-inlet beneath. In starting, the truck with the converter

Jan 1, 1908

By W. I. Duvall

A brief review of the resonant frequency and ultrasonic pulse methods for obtaining elastic constants of rock samples shows that the equations for an iso-tropic elastic solid commonly used to calculate Young's modulus, modulus of rigidity and Poisson's ratio can result in appreciable error if the rock is even slightly anisotropic. Consideration of the equations that relate bar velocities and free medium velocities to elastic constants for orthotropic, transversely isotropic and isotropic solids shows that a combination of the two methods can be used to determine average elastic constants for a single sample of rock. The determination of elastic constants in different directions requires a large number of tests on oriented samples. The use of both the resonant frequency and ultrasonic pulse methods is recommended so that the determination of both Young's modulus and modulus of rigidity is independent of the determination of Poisson's ratio. The dynamic elastic constants of rock can be determined in the laboratory from measurements on small samples which may be in the form of right circular cylinders (diamond drill core). Either the resonant frequency6 or ultrasonic pulse method1 can be used to determine the longitudinal and shear velocities in the same specimen. The resonant frequency method provides a measure of the longitudinal bar velocity, V,, and the torsional bar velocity Vt; whereas the ultrasonic pulse method provides a measure of the free medium longitudinal velocity, Vp, and the free medium shear velocity, V, . The elastic constants (Young's modulus E, modulus of rigidity G, and Poisson's ratio v) are calculated from the density of the specimen, p, and one of the pairs of velocities by means of equations from isotropic elastic theory. For the resonant frequency method the equations are: It should be noted that Eq. 5 is derived from Eq. 4 and 6 and the relation between the elastic constants for an isotropic solid as given by the first part of Eq. 3. The in situ determination of dynamic elastic constants is accomplished by generating a complex seismic pulse in the rock and measuring the arrival times for the longitudinal and shear phases of the pulse at two distant points.5 From these travel times and the distance of the travel path both Vp and Vs are determined. The density of the rock is determined from samples in the laboratory, and Eq. 4 to 6 are used to calculate the elastic constants. These three methods involve the assumption of isotropy and perfect elasticity, and require the use of the relation v = (E/2G) - 1, which holds only if the elastic constants are the same in at least two directions. The calculation of v by means of Eq. 3 introduces considerable error, as a 1% error in V, and V, results in a 20% error in v. If frequencies for samples with a length-to-diameter ratio greater than four and a length greater than 8 in. are measured carefully, the experimental error can be kept to 1%. However, it is not unreasonable to expect that the variation in Vb and Vt with direction may exceed 1%. Thus, for even a small percentage of anisotropy the use of Eq. 3 can result in considerable error in v. The calculation of v by Eq. 5 also involves some uncertainty as a 1% error in Vs and Vp can result in a 6% error in v. By means of the pulse method both

Jan 1, 1965

The recent dedication of Magma Copper Co.'s San Manuel, Ariz., electrolytic copper refinery and rod casting plant marks the culmination of a three year, $250 million expansion-modernization program. The new facilities will eventually permit an annual production of 200,000 tons of marketable copper for direct sale to fabricators. The Newmont Mining Corp. subsidiary will produce and market continuous cast 5/16 in. copper redraw rod, full plate cathode copper and sheared cathode plate under its own brand name. Among the four largest in the nation, the new refinery and plant, was built at a cost of $34 million and will be fully automated. In addition, all materials used in the plant will be recycled and no waste, discard or emissions will occur.

Jan 3, 1972

By Harrison R. Cooper

Computer process analysis and control techniques are overcoming major hurdles to acceptance in the mineral industry. With the growing list of new advances characteristic of a rapidly maturing discipline, management of the companies on the leading edge are keenly aware of the advantages of using computer methods in process applications. Significant advancement is already evident in such technical areas as operations research, mathematical modeling, statistical methods, and computer-based procedural techniques. The early advances during the past decade are already documented in prior symposium proceedings of this series. Thus, the currently active areas of the advancing technology are emphasized here. This section deals with applications in pyrometallurgical and kiln processes. The two sections which follow deal with concentrating operations from two points of view: 1) mathematical modeling, and 2) on-line control. In the metallurgical field, emphasis on developing effective process computer applications for the steel industry began in the late 1950's. Simulation techniques also received substantial attention. The steel industry was soon joined by the cement industry in the development of computer methods for process applications. Computers were in use for feed-material blending and kiln control in the early 1960's. Computer development in other pyro processes, such as aluminum reduction and glass-making, also marked significant advances by the mid 1960's. Aluminum smelter potline control is a nearly universal application at this point, and the development of glass furnace computer process control represented a recent important technical advance in digital control techniques. Steel Hot strip rolling was the earliest significant process computer application in the metallurgical industry. Computer control of rolling mill operations is now a steel mill standard component, particularly in plants newly placed into production. 1, 2 Initial computer installations principally carried out procedural routines in setting mill clearances for production lot control. Recent developments are leading toward feedback control based on mathematical model techniques. 3

Jan 1, 1969

By Linn Bradley, W. W. Strong, H. D. Egbert

Marked differences of opinion have been expressed by engineers interested in cleaning iron blast-furnace gases for use in hot-blast stoves and under boilers, in reference to the advantages of a hot-dry method over a cold-wet method. One point at issue involves the sensible heat energy in the moisture contained in the gas. Some advocates of the cold-wet methods claim that the condensation and resultant removal of the greater portion of this contained moisture by wet scrubbing, spraying or similar methods, results in a saving of some of this sensible heat energy, by reason of the fact that water vapor has a high capacity for sensible heat energy and may carry from the exit of a hot-blast stove, for example, more heat units than are sacrificed or lost when the gas is cleaned by this cold-wet method. Other advantages claimed for the latter method are: That gas burns more readily when it is free from moisture in any form; that because gas is made denser by cooling and removing the moisture it has a higher calorific value than hot gas carrying moisture; and that higher flame temperatures are obtained when the gas is cleaned by the cold-wet method. A search through the available literature fails to disclose any extensive calculations or records of conditions obtained in practice, bearing upon these phases of gas cleaning. This paper deals with these problems from the standpoint of the economy of using a gas of higher flame temperatures, of improved stove design, and of economy in the gas-cleaning department. Calculations have been made which show that by cleaning the gases by a cold-wet method the sensible heat energy of the blast-furnace gases is greatly reduced; and, in general, this loss of heat energy is far greater than that lost from the stoves or boilers due to the sensible heat capacity of the water vapor if carried away by the exit gases. By making use of the

Jan 1, 1917

By William Victor Vietti

A METHOD of determining the average density of the fluid column in a flowing oil well is presented. Actual field data from several wells are used to illustrate the application of the method in the Yates Field, Pecos County, Texas. The depth of the wells in the Yates Field varies from 900 to 1700 ft. depending on surface elevation and the location of the wells on structure. Production rates over 8000 bbl. hourly have been determined and many

Jan 1, 1930

By O. Katz, A. Spilners, M. Simnad

THE oxidation of titanium has been studied in recent years by a number of investigators.'-W ne of the most important aspects of the reaction of titanium with oxygen, which has been observed by several of the investigators, is the extremely high solubility of oxygen in the metal. A detailed study of the Ti-0 systeml% as shown that the solubility limit of oxygen in a-titanium is approximately 14.5 wt pct 0 from 800" to 1700°C. The unusual oxidation behavior of titanium may be ascribed to this phenomenon, since part of the oxygen uptake enters into solid solution in the metal and only a part is available for scale formation. A more complete study of the kinetics of oxidation of titanium would require information on the influence of oxygen in solution in the metal upon the oxidation rate. Such information would also be of interest in regard to the theory of oxidation of metals. The authors have measured the oxidation rates of oxygen-saturated and oxygen-free titanium between 800" and 1200°C and have found appreciable differences in their rates of oxygen uptake and in the activation energies. The experimental set-up was similar to that described in a previous paper.u The course of oxida-dation was followed continously by attaching the specimen to a sensitive spring balance and observing the extension of the spring with time of oxidation. High purity electrolytic oxygen was used and precautions were taken to purify the gases employed. The specimens were prepared from iodide titanium and were ground on metallographic polishing papers to a fineness of 4/0. The furnace was first evacuated and filled with argon; the specimen was lowered into the furnace, which was then evacuated and filled with oxygen, after the specimen had reached the oxidation temperature in argon. The samples were presaturated with oxygen by means of the following procedure: Each specimen was first heated in oxygen at 1200°C until its increase in weight indicated 13.8 pct by weight uptake of oxygen. The furnace was then immediately evacuated and filled with purified argon five times. The specimen was heated in the argon atmosphere for a period of 72 hr at 1200°C in order to saturate the metal with oxygen by complete absorption of the scale. The temperature was lowered to the oxidation temperature, the furnace was evacuated and filled with oxygen, and the oxidation rate was followed by noting the extension of the spring balance. No oxide film was visible on these specimens at the start of the oxidation experiments. Titanium samples in the oxygen-free as-received condition were also oxidized and their oxidation rates measured. The results are shown in Figs. 1, 2, and 3, which indicate that the initial rates of uptake of oxygen by both the oxygen-saturated and oxygen-free specimens are parabolic at these temperatures. The rate of uptake of oxygen is, however, appreciably lower

Jan 1, 1956

By Arthur F. Taggart

THIS paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. CAPACITY Operators know that the tonnage of rock that can be put through a particular crusher in a given time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: TR = TR80KK'K" [I] in which TR, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R50 = "80-per cent reduction ratio"; and K, K', and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific-gravity range-2.5 to 2.7-T should be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppard' has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about I.5: I: 0.6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:0.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for -150 + 200-mesh pyrite to be I.4: I : 0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value I.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciable tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 8o per cent of the feed, divided by the aperture that would pass S0 per cent of the product. It is best determined from screen tests.

Jan 1, 1941

By Richard Akerman

At the International Sessions of the Iron and Steel Institute of Great Britain, the American Institute of Mining Engineers and the Verein Deutscher EisenhUttenleute, held in Allegheny City, Pa., in October, 1890, Sir James Kitson, then President of the Iron and Steel Institute, presented an account by Sir Henry Bessemer of the origin of the process which bears his name.? The paper graplhically told how, after the first successes immediately following the first reports given at Cheltenham of the new process, there rapidly succceded such obstacles and difficulties in its practical application, that the whole matter was nearly abandoned in 1858. Suddenly, however, the situation changed, so that at length it became possible

Jan 1, 1894

By Alvin Kaufman

At the turn of the century the U.S. had a gross ALVIN KAUFMAN Mineral Economist U.S. Bureau of Mines area slightly in excess of three million square miles, a population of 76 million, a gross national product of $37 billion (constant 1929 dollars) and a mineral production valued at $2 billion (1929 dollars). Some 60 years later its area had increased by 600 thousand square miles, population had risen 140%, gross national product was seven times greater and mineral production had grown to a value five times that of 1900. Of perhaps greater significance, however, were the changes wrought by technology. It was during this period that the "American Way of Life" was revolutionized by the introduction of the airplane and the automobile; man also took his first probing steps in the utilization of nuclear energy and pre- pared to unlock the secrets of space. It was during these six decades, also, that he learned to create additional new materials, some of which were from old sources such as plastics from petroleum and coal. These new materials have created competition and attendant problems for the traditional materials such as iron, copper, lead and zinc. Additional problems undoubtedly will be created by technologic advances in the future.

Jan 2, 1963