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Papers in Classes A-Metal Mining, B-Milling and Concentration, H-Indus¬trial Minerals, and I-Mining Geology are distributed in MINING TECHNOLOGY, which is issued every other month. Papers in Classes C-Iron and Steel Division, D¬-Nonferrous Metallurgy and E-Institute of Metals Division are distributed in METALS TECHNOLOGY, which is issued twice quarterly. Papers in Class G-Pe¬troleum Division are distributed in PE¬TROLEUM TECHNOLOGY, which is issued every other month. TECHNICAL PUBLICA¬TIONS in Classes F-Coal Division, J¬-Mineral Industry Education Division, K¬-Mining Administration, and 1-Geophysics, are issued as separate pamphlets. Contributions are papers that when printed or mimeographed are available but only on request.

Jan 1, 1943

By Ned Kuenhold

AN equipment manufacturer swallows a bitter pill when he sees his machinery improperly in- stalled and not used to the full potential. This frequently happens with vibrating screens. Screens can be classified according to the path of vibration imparted to the screen surface. On this basis there are: 1) vertical circle or modified circle throw, 2) horizontal circle or modified circle throw, 3) vertical linear reciprocation, 4) tilted linear reciprocation, and 5) horizontal linear reciprocation. The most versatile vibration for medium to coarse sizing is generally conceded to be the vertical circle produced by an eccentric or unbalanced shaft, but other types of vibration may be more suitable for certain screening operations, particularly in the finer sizes.

Jan 6, 1957

By James R. Evans

Forecasting demand is an art as well as a science, and much personal judgment is required. National forecasts made for sand and gravel, crushed stone, and/or aggregate may be misleading or unhelpful locally. This is because of significant variation in demand within different geographic and/or political areas in the US, and/or the use of an inadequate data base and forecasting methodology. The best forecasts may be made for local or regional market areas, irrespective of geographic or political boundaries, provided a proper data base is available.1 The Bureau of Mines compiles detailed statistics on commodities sold or used in the construction industries and, therefore, is in an advantageous position to make national aggregate demand forecasts.

Jan 1, 1979

By K. K. Clark

Excessive injection pressures in water injection wells may create deeply penetrating fractures, or may open up existing reservoir fractures. If these fractures are oriented toward offset producing wells, oil recovery at injected water breakthrough and ultimate recovery will be impaired. A method of calculating fracture lengths from pressure fall-off test data is presented. The method is based on a linear flow model that simulates conditions present during the early-time period after shutting in an injection well. Fracture lengths can be calculated directly if formation permeability is known. A graphical technique is presented that provides fracture lengths based on permeabilities calculated from normal fall-off tests, where those permeabilities are adjusted for flow geometry. Test data from four injection wells and results obtained from applying the method to these wells are discussed. Introduction Rates of water injection into nonstimulated wells in low permeability reservoirs frequently fall below economically desirable levels. Therefore, some form of stimulation such as hydraulic fracturing often is performed on these wells. Usually these fractures penetrate a short distance from the well — short as compared with interwell spacing — and therefore should not create problems of water channeling to offset producing wells. In the past, some operators have been tempted to increase injection pressures continually to maintain a specified injection rate. But pressures cannot be increased indefinitely. Field data indicate that continued injection above fracture opening pressure can cause excessively long fractures to approach between well distances. If these long fractures intersect or closely approach offset producing wells, premature water breakthrough can result. It is difficult to seal effectively these fractures by various workover procedures after premature water breakthrough has occurred. This paper describes testing procedures that are potentially capable of detecting premature water breakthrough before it occurs. Knowledge of fracture lengths and fracture opening pressures can assist field engineers in selecting optimum operating conditions for water injection wells. Fracture lengths should be calculated from pressure fall-off tests on injection wells, and maximum permissible iniection pressures should be determined from step-rate injectivity tests. Equipped with this information, the field engineer is in a better position to recommend for a particular injection well the optimum operating conditions that will lessen the chances of channeling to offset producers. Historically, pressure transient tests have been conducted primarily on producing wells and analysis procedures have been based on radial flow concepts. Several authors have suggested that these same procedures be applied to tests conducted on water injection wells. This application is justified for wells that are not connected to extensive fracture systems. However, if the formation is severely fractured, the interpretation methods must be modified. Several authors have presented the pressure response of various systems that are intended to simulate reservoir conditions. A paper by Russell and Truitt* on vertically fractured svstems is probablv the most useful from the standpoint bf the fieid engineer. Russell and Truitt presented the pressure-vs-time behavior of a well connected to vertical fractures with various fixed lengths. Their permeability adjustment technique is used in this paper to refine methods of calculating fracture length and interwell permeability. Theory and Definitions Three flow-system geometries are considered in analyzing injection well test data. These various geometries are applicable during specific time intervals of transient pressure tests. Schematic drawings of the flow models, and the time intervals over which they are applicable, are shown in Fig. 1 in the chronological order that they are applied to the pressure response of water injection wells in fractured formations. The very early response is simulated by an infinite linear system as shown in Fig. 1A. This representation permits the calculation of matrix surface area exposed to the fracture. The model shown in Fig. 1B is identical with the one studied by Russell and Truitt. They evaluated the transient pressure response of this vertically fractured system by finite difference techniques. The results were presented as tabulations of dimensionless pressure drop functions for a range of fracture lengths. A semilog plot of these data shows that the apparent matrix permeability can be related to the true matrix permeability by a simple exponential function. Their adjustment factor is used in this paper to improve the accuracy of matrix-permeability and fracture-length calculations.

Jan 1, 1969

By N. K. Chen, R. B. Pond

IN the study of slip band* formation, there have been many examples to show that they do not always appear as lines traversing the entire crystal, but as segments whose ends seem to vanish in their path through the crystal. This characteristic appearance of slip bands has been witnessed under the optical microscope at various magnifications'-' and also under the electron microscope.' A typical example of this behavior seen with the optical microscope is shown in Fig. 1. However, the slip band studies were generally conducted on polished surfaces of single or poly-crystalline metals which had been previously deformed; i.e., the load was generally released so that the observation could be made. Any picture of slip bands so obtained can represent the surface phenomenon only in a static state of the strained material. The conditions prior to their formation cannot be definitely and clearly assigned. Thus, while a segmented slip band may suggest that slip is a growth process as supposed by the theory of the nucleation of slip,V he usual appearance of suddenly and fully developed slip bands around the crystal has generally been considered as a consequence of uniform shear of the entire slip plane akin to a cataclysmic process. Little clear-cut information is available with regard to the speed at which a slip band forms, its direction of motion, the geometry of its position with regard to its neighbors, and its dependence on orientation. A description is given in this paper of experimental apparatus by which the progressive formation of slip bands can be recorded while the specimen is undergoing deformation. Qualitative and quantitative data on the dynamic formation of slip bands will be presented with special interest concerning the propagation of slip bands, the spacing of slip bands, and their relations to strain hardening. Views on the formation of slip bands are discussed and a mechanism of the unit process involved in the formation of a slip band is proposed. Preparation of Specimens Single-crystal specimens of high purity aluminum (99.997 pct), 1/8 in. square in cross-section and 1% in. long in gage length, were made by the method of gradual solidification from the liquid state. Since no machining work could be introduced in preparation of crystals of such small size, a special mold was designed for casting them to final shape. The mold consisted of two, separate, high purity graphite blocks. Generally, 20 molds were packed together in one container so that 20 specimens could be obtained by one casting operation. This was desirable since it was possible by this method to obtain groups of crystals with similar orientations. The as-cast crystals were carefully clipped from the gate, etched, and homogenized for 24 hr at 600°C. They were then very gently polished using a 4/0 paper, re-etched, and finally electrolytically polished after the method previously described by Chen and Mathewson.6 The crystallographic orientations were determined using a back-reflection, Laue method. Tensile Testing and Photographic Method The tensile testing equipment for these tests was composed of a specially designed microtensile machine, microload cell and microclip gage with necessary appurtenances. The members of the microtensile machine consisted of three parts, as shown in Fig. 2. The chassis is equipped with an oil cylinder, A, and piston, the piston being part of the movable cross-head, B. Pressure in the oil cylinder is controlled and regulated by an external pneumatic-hydraulic cell, C, which is connected to the cylinder by a Vs in. high pressure copper tube. This cell is half filled with hydraulic oil and has a needle valve, D, on the oil exit side as well as a needle valve, E, on the gas inlet side. A quick-acting valve, F, as well as a pressure gage is provided for the gas side. The oil exits into the load piston on the microtensile machine. By connecting a tank of inert gas to the gas inlet, regulated pressures were provided over the oil so that the oil would leave the cell at a rate determined by the setting of the exit needle valve, the gas pressure, and the pressure of the oil in the load piston of the microtensile machine. With this pneumatic-hydraulic appurtenance it was possible to

Jan 1, 1953

By C. R. Fulton

The idea of making permanent magnets by sintering is not a new one, but until the introduction of the iron-nickel-aluminum permanent-magnet alloys there was little need for such a process. Permanent-magnet alloys in use up to the time of this development could be forged and machined to the exact size or shape required without too much difficulty. These new Fe-Ni-A1 alloys, which have become known as Alnico in this country, soon became very popular because of their superior magnetic properties, but when they were made by casting they did have some disadvantages when compared with the older type of permanent-magnet steels. Among these disadvantages were the following: 1. It was almost essential that they be cast to size, since they could not be readily forged or machined. The material could be ground but this was quite expensive, particularly for small magnets in large quantities. 2. The material had a large grain size and was mechanically weak. 3. Close tolerances could not be held and on small magnets this meant grinding almost the entire magnet. Because of the brittle nature of the material, there was considerable loss due to chipping and cracking during these grinding operations. 4. Because of its high aluminum content, the molten alloy had a high surface tension. It was difficult to prevent loss of aluminum by oxidation during the melting operation. In view of these difficulties it is not hard to understand why a process for sintering alnico was developed, or why sintered alnico has come into such widespread use in recent years. It might be well to qualify this by saying that it has come into widespread use for small magnets weighing 2 oz. and less, of the type shown on Fig. I. In this size range, sintered alnico has been able to compete in price with cast alnico. Magnets larger in size, of the type shown on Fig. 2, can be made but they do not compare favorably in price with the cast material. An idea of the range of sizes that have been produced is given by Fig. 3, which shows a small magnet weighing 0.006 oz. and a large magnet with a steel insert weighing 4 lb., both made from sintered alnico. Magnets weighing more than 2 oz., however, should not be made from sintered

Jan 1, 1945

By C. O. Brawner

Introduction The procedures used to design slopes in rock have only been developed over the past 10 years. Therefore, continued review and assessment of field experience by the publication of case studies is necessary to determine the accuracy of the theory and the application of the design techniques being used. Where failures have occurred, mine management should insist that the slides be 'back analyzed' to determine the shear strength parameters at failure. These values should be used for any pit design as they reflect site conditions more accurately than any other procedure. The previous papers have dealt with the basic concepts and factors that influence stability. To illustrate the application of these concepts in the study of practical stability problems, six ease studies have been selected for review. The examples include failure of tailings dams due to an earthquake, failure and redesign of a slope in overburden, review of failures of rock slopes, calculation of the amount of seepage into a pit and influence of the groundwater on stability, and the design of slopes for a proposed 2,500 ft. deep open pit mine. All but the first example are drawn from the authors personal experience. EXAMPLE 1 Failure of Tailings Dams Due to an Earthquake On March 28. 1965 a serious earthquake occurred in Chile with the epicenter located near the town of La Liqua, about 80 miles north of Santiago. The earthquake had a rating of 7 to 7.25 on the Richter scale. A record of the acceleration was taken at Santiago where a maximum horizontal acceleration of 0.18 g. was recorded. Many small mines are located in this region. Each of the mines provided impoundment storage for tailings with one to three dykes. As a result of the earthquake, most of the dams failed. The most serious failure involved two dams at El Cobre where more than 2,700,000 cu. yd. of tailings flowed into the valley, travelled over 7 miles in a few minutes, destroyed part of the town of El Cobre and killed more than 200 people. The Chilean government commissioned an investigation into the failures. A summary of the findings was published by Dobry and Alvare7 in 1967 (1). The location of the earthquake and the tailings dams which were investigated is shown

Jan 1, 1971

By Herbert Haas

ChEap power is essential to large-scale mining and metallurgical operations, particularly where fine grinding of large tonnages has to be resorted to, as is the case with an increasing number of mines treating disseminated copper ores, or low-grade gold ores, in the grinding of which considerable power is consumed. A comparison of the cost of generating power with Diesel engines and steam turbines and a brief discussion of the different factors governing the profitable use of either type of prime mover may be timely in view of the importance of a cheap supply of power to the mining industry, and the prominence that has lately been given to Diesel engines in America. As many of our most important mines and largest producers of metals are located in sections where neither cheap coal nor water power is available, and the price of oil fuel is increased by long hauls, the Diesel engine will become a prime mover of increasing importance. This applies particularly to the Southwestern States, where an ample supply of cold water for condensing purposes is unusual and local climatological conditions interfere in securing the high vacua essential to high turbine efficiency. About 12 times as much water will be required for condensing purposes with steam turbines under general conditions as is needed for jacket-cooling of two-stroke cycle Diesel engines; four-stroke cycle Diesel engines require only one-twentieth of the cooling water. Another condition favoring the use of Diesel engines is the high load factor at which most mine power plants operate. The influence of load factor on the cost of generating power is shown in Tables 1. and 2 and in Fig. 1. The higher the cost of fuel and the load factor, the greater will be the proportion of the fuel to the other costs making up the total cost of generating power. Capital charges (interest and redemption), maintenance, operating labor, and lubrication are practically constant amounts regardless of the load-output of the station; the fuel consumption and its cost are practically the only variables with the load. The costs here given are based on conditions as generally found in the Southwestern States, and on Diesel engine operating experience in European countries, where this type of prime mover has reached a high state of perfection, units upward of 4,000 hp. having proved entirely successful in operation.

Jan 1, 1917

By S. Le Fevre

The application of electricity to the mining and beneficiation of the magnetic iron ores of the Mineville, N. Y., district, on Lake Champlain, has resulted,in economies and enhanced quality of product, without which only a comparatively small quantity of the present output could be marketed. The output of the Witherbee, Sherman & Co. mines is about 700,000 tons per annum, and of this about four-fifths is treated by magnetic separation to increase the iron content and reduce the phosphorus. The use of electricity for the magnetic separation of the ore from the gangue is therefore of primary importance, and makes possible the application of electric power for mining and preparing the ore for concentration. The fact which speaks loudest in favor of the application of electricity is this: It is possible for about 25c. per ton of shipping ore to furnish all the power required to put the ore on cars. This includes mining a hard ore underground, from a depth of 900 ft., hoisting to surface, crushing to $ or 8 in. in size, and separating from the gangue. This also includes lights about the works, and all miscellaneous power, such as about the shops. The first plant for the generation of electric power for use at Mineville was completed in 1903, and was one of the first for the application of electricity to mining iron ore. This was a Nordberg compound condensing steam engine, direct connected to a 750-kw. alternating-current generator, 3,300 volts, 3 phase, 25 cycles. All the plants added later use the same current but with higher voltage. The map, Fig. 1, shows the power plants and the transmission system; Figs. 2 and 3 are views of the shaft houses and the mills. The power from the first plant was distributed to Mills 1 and 2, one 2,500 cu. ft. compressor, and to the Harmony hoist, which was driven by a 300-h.p. motor operating four 10-ft. drums. Our experience with the hoist covers largely the development of electric hoisting from 1903 to date, and in some cases improvements were worked out to satisfy our needs. The motor is still operating as installed, but the control apparatus has been radically changed several times. When installed, the iron re-

Jan 1, 1915

By P. J. Carroll, H. Colijn

Feeders must be considered an integral part of the over-all bin and feeder system. Improper design of either one of these parts will affect the performance of the whole system. Feeder load, material properties, feed rate, physical layout, and environment as well as first cost determine the selection of feeder type. This paper provides a new method of calculating feeder loads with some experimental data as verification. General design criteria for belt and apron feeder, rotary table and plow feeders are discussed and a special section is devoted to vibratory feeders. Feeders are used to provide a means of control for the withdrawal of bulk materials from storage units such as bins, bunkers, silos and hoppers. This control function can only be performed properly as long as the bulk materials flow by gravity to the feeder in a uniform and uninterrupted fashion. During the past few years various papers have been written and presented on feeders and their application.1-9 This paper will be confined to some specific bin-feeder design criteria such as feeder load and power requirements. Special attention is given to the basic design criteria for vibratory feeders. HOPPER AND BIN DESIGN Feeders must be considered an integral part of the over-all bin and feeder system. Improper design of either one of these parts will affect the performance of the whole system. A lot of effort has been devoted in the past decade to the development of functional design criteria for hoppers and bins.1,10 The integrated concept of bin and feeder design requires quantitative analysis of the bulk material characteristics before the design and the selection of components can be attempted. The design of a feeder system must begin with the proper dimensioning of the hopper outlet to prevent arching or ratholing. The hopper opening should be large enough to allow passage of the bulk solid at the required maximum discharge rate." A feeder can only throttle the flow. FEEDER LOAD DETERMINATION Over the past ten years, there have been various attempts to determine the exact load or pressure on feeders mounted directly under hopper openings. Many designers, however, still consider this pressure equal to the hydrostatic head of material above the opening, and because such pressure would be very

Jan 1, 1969

By V. W. Aubel

Among the iron-bearing ores of the world, the titaniferous iron sands of New Zealand are probably the least known to American engineers. This is not surprising in view of the fact that American ironmasters look with disfavor on any, even domestic, ores containing over 1 per cent. titanium. The possibilities of smelting titanium-bearing ores in the blast furnace are still undetermined, but the problem of their efficient reduction is becoming more important each year. There seems to be little inclination on the part of American furnace operators to tackle this problem, and they will no doubt continue to utilize the non-titanium ores at their command, leaving research and experiment to those countries not so fortunately endowed. New Zealand has, however, been intensely interested in her deposits of titanium-bearing iron sand, and during the last 50 years considerable English and New Zealand capital has been expended on this problem. The extent of the New Zealand iron-sand deposits is almost beyond measure. From Patea north to Manuka Heads, along almost the entire west coast of the North Island, these deposits exist. Wherever there is beach sand along this 300 mi. of coast line, the sand is iron bearing. It varies considerably in grade; but wherever the contour of the coast affords small bays and inlets, large dunes are found that are remarkably uniform in quality. Several of these large deposits, especially those at New Plymouth and Patea, are noted for their high iron content. The dune at New Plymouth runs between 55 and 60 per cent. of metallic iron; the panning action of the prevailing westerly winds and northwesterly currents has no doubt resulted in this concentration. Whatever the origin of this material may be, there is no doubt that deposition is going on continuously. At New Plymouth, a small artificial harbor, protected from the drifting sand by a large breakwall, has been dredged. Here sailing vessels have for years been loaded with iron sand for ballast by means of a crane and grab bucket operated from the wall of the breakwater. Thousands of tons of sand have been removed from this spot yet each succeeding storm renews the supply. The sand dune shown in Fig. 1 has been tested to a depth of 50 ft. (15 m.) without losing its

Jan 1, 1920

By T. R. Ingraham

The thermal decomposition of Nickel (II) sulfate was examined by determining the total pressure of SO3, SO2, and O2 developed over a sample when it was heated in an evacuated system fitted with a Pyrex-bellows mercury manometer. The equilibrium pressures were combined with previously published data on the oxide and sulfide systems to establish a predominance volume diagranz for the Ni-S-O system over the normal range of roasting temperature and gas composition used in the treatment of nickel minerals. The protective coating of NiSO4 which forms on the oxide during roasting, and which causes the sulfation reaction to almost cease, may be disengaged readily by temperature cycling. When nickel sulfide is roasted in air, either nickel sulfate or nickel oxide may be formed by altering the experimental conditions. Because of the differing stabilities of various metal sulfates and oxides, possibilities exist for separating some of the constituents of mineral mixtures by the technique of selective sulfation and oxidation roasting. The experimental information required to make reasonable predictions of the possibilities of this technique is incomplete and, in some cases, in conflict. As part of a general program of providing some of the required information, a study EXPERIMENTAL Materials, Apparatus, and Procedure. Fisher certified reagent-grade nickel sulfate hexahydrate was used in the source material for both nickel sulfate and nickel oxide. The manufacturer's analysis indicated: 0.001 pct Fe, 0.005 pct Cu, 0.003 pct Co, 0.001 pct C1, other 0.15 pct. The material was dehydrated, before use, by heating at 420°C for 5 hr. The oxide was prepared by roasting the sulfate in an open tray in a muffle furnace for 10 hr at950°C. Preliminary experiments which were made by thermogravimetric analysis and X-ray diffraction analysis established that only one crystal form of nickel oxide and one of nickel sulfate were involved in the equilibrium shown in Eq. [I]. The apparatus used in the experiments was described previously.1,3 The main feature is a corrosion-resistant, flexible Pyrex bellows which is used to protect the mercury in a u-tube manometer. Before a run was begun, the apparatus was washed thoroughly with hot hydrochloric acid and water and dried by heating under vacuum. The mixed sulfate -oxide sample was contained in a platinum bat rest-

Jan 1, 1967

By A. W. Schillinger

Drilling was not entirely satisfactory in the search for native copper in the Osceola amygdaloid footwall, for oreshoots mined were more continuous than drilling indicated. The Geophysics Dept of Michigan Technological University developed a portable induced polarization drill hole probe with a 3-electrode configuration and a sampling radius of 5 ft ±. A 20 ma current of alternately reversed polarity is pulsed at 3.5 sec intervals and readings are taken every 2 ft. Potential is read during the "on" cycle for resistivity determinations. During the "off" cycle, after 10 ms delay, the 1P potential is sampled for 10 ms and recorded. Resistivity (ohm-ft) and "S" values (mv/v) are plotted on semi-log paper. A lithologic log is plotted based on resistivity and anomalous S values noted. The low resistivity amygdaloid contains sills of high resistivity basalt. Verified by core logs, contacts between amygdaloid and basalt could be picked to ± 0.5 ft from resistivity alone. Lithologic logs of long-steel holes, based on resistivity, then replaced sludge logging. An S value of 30 mv/v was established as the lower limit of diagnostic 1P effects in mineralized amygdaloid and 50 mv/v as a threshold value for ore. Of 914 footwall zones drilled, 315 were probed yielding 203 anomalies. Success increased from 34% for drilling to 72% for drilling and probing. As checked by mining, 46 probe results were 78% correct. Failures were due to the instrument not detecting copper or to the misinterpretation of an anomaly. No false anomalies were found. During the past ten years, the Calumet Div. of Calumet &Hecla, Inc has conducted extensive exploration and mining in the footwall of the Osceola amygdaloid lode in the native copper district in Michigan's Upper Peninsula. In an attempt to delineate more clearly the extent and nature of high-grade footwall copper mineralization, 876 diamond drill and long steel holes aggregating slightly more than 47,000 ft have thus far been drilled from the accessible hanging-wall workings in the Osceola No. 6, 13, 14 and 15 shafts. Early in this program it became obvious, through inspection of footwall workings in progress, that due to the erratic and bunchy occurrence of the native copper in the footwall zone, the exploration drilling was failing to outline completely the extent of the oreshoots. Early drilling re-

Jan 1, 1964

By Robert Bedford

THE objects of this paper are: 1. To give cost data on drilling obtained at the North Star mine in California; 2. To describe the methods of testing drills employed there, giving the results of shop tests, and the correlated data obtained underground; 3. To demonstrate that these tests form a satisfactory basis upon which to judge rock drills. The rock at this mine is either close-grained diabase in the upper levels, or tough grano-diorite in the lower levels. The vein, which has an average dip of 23°, consists of about 5 ft. of "formation" lying between walls of unaltered country rock. Of the 5 ft., solid quartz and stringers make up 18 in. of pay ore. The formation, which consists of but slightly altered country rock, is almost as hard as the unaltered walls. From 4 to 42 ft. is sent to the mill. One-third of the holes are drilled in the quartz; two-thirds in the formation. The holes are about 42 ft. deep and break on the average of 1.15 tons per hole. The average number of holes per stope drill shift is at present 5.65. The number of drill shifts throughout the mine in 1913, was 18,679, and the cost of labor for drilling, power, supplies, upkeep of machines and air lines, tool sharpening, and distribution, amounted to 33 per cent. of total mining expense. To do this work there are in commission 43 No. 12A, three No. 17V and eight No. 16V Waugh stopers, three Jackhamers, two Waugh pluggers, and 17 No. 8 ,and two No. 7 water Leyners. For the past three years a record has been kept of the repairs on each machine and number of shifts operated by it. The figures given in Tables I and II for the cost of repairs per drill shift have been taken from this record. The underground air pressure is about 90 lb. The air consumption per drill shift was measured by, aerometers. The leakage of pipe lines between compressor and meter is not included. From the number of pieces of steel sharpened for each type of drill during the year the average used per drill shift has been ascertained. For other supplies the figures given represent a year's average.

Jan 8, 1914

By F. E. Bash

Previous to the war, this country depended on Europe for its supply of a number of alloys of great importance in the manufacture of electrical apparatus and equipment. When this source was cut off shortly after the beginning of hostilities and the small stocks in the United States were exhausted, either inferior substitutes had to be used or the manufacture of the apparatus suspended. Since electrical equipment was of primary importance in the prosecution of the war, the substitutes had to be used until the composition and processes of manufacture of the alloys could be determined. Included among these alloys was one generally known as manganin, which is composed chiefly of copper and manganese and which, beside having a low-temperature coefficient of resistance, has a low thermal electromotive force against copper. These qualities make it valuable for the manufacture of electrical instruments and machines in which it is used in the shape of wire, rod, and strip. As a result of these circumstances, an investigation was started to determine how manganin could be made on a commercial basis so that a product of uniform quality and the necessary electrical characteristics could be produced. The object was to procure an alloy with temperature coefficient of resistance not greater than 1 X 1X per " C. and a thermal e.m.f. against copper no greater than that of the imported material. It is characteristic of manganin that the temperature-resistance curve is not a straight line but has a peak, which varies in position with different samples; it is most desirable, however, to have it at a temperature of 25' C. if possible. The coefficients of manganin from all records available ranged from a large negative to a large positive value. The composition as reported by different authorities is given in Table I. Prof. Matthew A. Hunter, of Rensselaer Polytechnic Institute, made up a number of small melts of manganin in a graphite crucible and the one that he reported as having the best cocfficient gave the following analysis: Copper, 83.6 per cent.; nickel 2.5 per cent.; manganese, 13.6 per cent.; iron, 0.3 per cent. An analysis of a sample of German manganin gave the following: Copper, 82.6 per cent.; nickel, 3.8 pcr cent.; manganese, 12.8 per cent.; iron 0.7 per cent.

Jan 1, 1921

JULY was a month of great turmoil, both economic and political, in many parts of the world. Instead of conditions improving, they seem to have been getting worse. In Ireland, civil war is still in progress. Germany was in the throes of a political upheaval. The meaning of the assassination of Rathenau became clear. It evinced the development of a class struggle between the bourgeois and socialistic parties which may rock the country to its foundation and will increase the difficulties of arranging things with the French and solving the reparations problem. In Poland, a similar contest is developing. Austria is already so completely contaminated by the socialist element that it is unable to get outside assistance and is thus falling into a state of despair but little short of what prevails in Russia. The conference at the Hague came to an abortive end for the reason that it was impossible to discuss affairs with the Russians in any reasonable way. The Germans declared their inability to meet the reparation payment that fell clue on July 15 and the Commissioners for the Allies granted them an abate-ment of a portion thereof and are listening attentively to the supplication of the Germans for a moratorium of one or two years. In the United States, July was a month of widespread labor troubles, accompanied by a great deal of disorder and violence, which led to detachments of the National Guard being called out in several states. The railway shopmen, about 400,000 in number, struck on July 1. The road maintenance men decided not to strike, but many thousands of them quit work without the sanc-tion of their union. There were more strikes in the textile industry in New England and in similar industries all over the country. Early in July, it was estimated that about 1,250,000 men were out on strike. The coal strike continued. Early in the month the President summoned operators and miners to Washing-ton and submitted a plan for the compromising of disputes. After a week's consideration the miners re-fused flatly to agree to anything. The majority of the operators consented to accept the proposed arbitration, but a minority refused. Following this abortive at-tempt the President dismissed the conference with the injunction "to return to your mine properties and resume operation." Production has not increased.

Jan 8, 1922

By J. R. Campbell

In order to discuss the subject intelligently, it will be necessary to touch briefly on the forms in which sulphur is supposed to exist in coking coal to be carbonized in beehive or byproduct ovens. Sulphur is known to exist in coal as sulphides and sulphates, as can be determined experimentally. Then there is the so-called organic sulphur, i.e., sulphur in combination with the carbon, 'hydrogen, and oxygen of the coal, about which much has been written, but nothing definitely proven in an experimental way. In fact, so far. as we know, there never has been developed a satisfactory and conclusive method in the laboratory for the direct determination of organic sulphur. For most coking coals, it can be safely assumed that the preponderance of sulphur is in the form of pyrite (FeSz). When exposed to comparatively low temperatures during the coking process, it loses its sulphur according to the following chemical reaction: 7FeS2 = Fe7Ss + 6S. It will be seen, therefore, that six out of the 14 atoms of sulphur (or 42.8 per cent.) are expelled, or volatilized. FelSa remains in the coke as pyrrhotite, or magnetic sulphide of iron; this accounts for the highly magnetic properties of the powdered coke. A more commonly accepted reaction is FeS2 = FeS f S, in which the straight sulphide of iron is produced with 50 per cent. volatilization of sulphur. In the beehive methods of coke making, air is introduced in sufficient amounts to carry on the distilling and the coking processes, and the sulphur is oxidized along with the other volatile products: First, into sulphur dioxide (SO2), known by the pungent andsuffocating odoremitted from the trunnel head; second, into sulphuric anhydride (SOa); and, finally, into sulphuric acid, as it comes into contact with the air and moisture. In a properly regulated draft on a beehive oven, there is never complete combustion of the gases. In other words, the coking process should be carried on in a reducing atmosphere, and a low-grade producer gas kept issuing from the trunnel head. For coals of about 30 per cent. volatile matter, the ratio of air to gas is 3½ to 1. In complete combustion, the ratio is 6 to 1, producing an extremely high temperature

Jan 1, 1916

By F. A. Forward, A. H. Ross, J. Halpern

IN treating carnotite ores by carbonate leaching it is often necessary to subject the ore to a prior roast. Among the advantages that may result from roasting are: 1) improvement in settling and filtering characteristics of the ore; 2) destruction of carbonaceous impurities that would otherwise contaminate the leach solution and consume carbonate reagent; and 3) substantial increase in extraction of vanadium, which is usually low when the un-roasted ore is leached. Recovery of vanadium may be an important factor in processing carnotite ores, and the fact that vanadium is usually recovered in good yield by acid leaching has often been an economic advantage over results obtained with carbonate leaching. It has been found that response of an ore to carbonate leaching, following roasting, depends primarily on its composition and on the conditions of the prior roast treatment, particularly with regard to the roasting temperature, the atmosphere maintained above the charge during roasting and cooling, and the presence of certain reagents. These factors, especially the last, virtually determine the extractions of uranium and vanadium attainable by carbonate leaching, as well as the reagent consumption on leaching, whereas adjustment of the leaching conditions has only a limited influence. These variables were examined systematically to elucidate roasting and carbonate leaching of carnotite ores and to establish a treatment yielding high recoveries of both uranium and vanadium. Tests were conducted on samples of four different Colorado Plateau ores, described in Table I. Roasting Procedure: The roasting tests were conducted in an electrically heated furnace in which the temperature was controlled to +15oC. The ore (usually a 500-g charge) was ground to —100 mesh, mixed with the desired reagent, and spread in a silica tray to a bed depth of about 1/2 in. During the roast the charge was rabbled at intervals of 15 min. To provide free access of air the furnace was usually ventilated; however, in a few experiments a controlled atmosphere such as carbon dioxide or hydrogen was maintained above the charge. Leaching Procedure: The pulp, generally comprising 500 g of ore and a 2000-cc solution containing 80 g per liter of Na2CO3 and 20 g per liter of NaHCO3, was agitated in an autoclave at 120°C, under 30 psi O2. Samples were withdrawn periodically, and the solution and ore residue were analyzed. Under the conditions employed, maximum attainable extractions were generally achieved in 4

Jan 1, 1958

By F. S. Buffington, Morris Cohen

THERE is little quantitative information available concerning the effect of applied stress, in both the elastic and plastic ranges, on diffusion kinetics. Accordingly, a program has been undertaken to determine the rate of self-diffusion in high purity iron, with the specimens subjected to stress during the course of diffusion. Results are reported here for body-centered cubic iron at a diffusion temperature of 890°C. The iron used was 99.97 pct Fe, the principal impurities being 0.002 pct C, 0.008 pct 0, and 0.02 pct Si. The surface-decrease method of self-diffusion analysis was employed, and Fe55 Mn x-radiation, 2.9 years half-life) was the tracer element. Details of specimen preparation and equipment are given elsewhere.' Disks of the high purity iron, % in. diam x 1/4 in. thick, were annealed at 890°C for 18 hr in dry hydrogen to produce large grains of a ferrite. The faces of the disks were then surface ground, and possible grinding stresses were relieved by annealing at 890 °C for 45 min in vacuum. A layer of radioactive iron 0.00186 cm thick was electroplated on one face of each disk. The self-diffusion coefficient was calculated from the ratio of the radiation intensities emanating from each of these faces, as measured before and after diffusion.' The heating was carried out in a vacuum furnace (10-5 mm Hg) controlled to ±1.5°C. The specimens were loaded in compression between two quartz disks during the diffusion runs by means of a lever system operating through a vacuum seal. Thus one quartz disk was pressed directly against the radioactive face of the specimen, minimizing the vapor loss of radioactive iron. Steady-state creep occurred in all instances where permanent deformation could be detected, but the specimens did not exhibit appreciable transverse barreling until natural strains of 0.2 to 0.3 were attained. Essentially, the diffusion took place under uniaxial stress. The plastic strain in each case was obtained at the diffusion zone by measuring the diameter d there before and after the run: e = 2 In — do Let Du equal the self-diffusion coefficient which would have prevailed at the given temperature if no stress had been applied. This was calculated from the temperature-dependence equation which had been previously determined: —59,800 D, = 6.2 exp ( I sq cm per sec* [I]

Jan 1, 1953

By J. Parke Channing

THE production and use of copper in the world, like that of other metals, has been of slow growth. In 1880, production in the United States, was only about 60,000,000 lb. and the world's production 325,000,000 lb. Up to 1913 there was a gradual and regular increase in production; in 1912 the production was 1,250,000,000 lb. for the United States and 2,200,-000,000 lb. for the world. At the outbreak of the war there was a slight falling off, both in the United States and elsewhere, but when it was realized that the war was not to be a short one and that copper was one of the most important things for carrying it on, there was a tremendous increase in production every-where. Roughly speaking, the average production for the years 1916, 1917 and 1918 was very close to two billion pounds in the United States and three billion pounds in the world. WAR DEMANDS STIMULATED PRODUCTION Not only did the production of copper in the United States increase, but the same may be said of .all the base metals, particularly of zinc. Much of the zinc used by the Allies came from Australia, but had been smelted in Belgium and Germany, now closed to them. We were called upon to make good this lost supply and in feverish haste built new zinc smelteries and opened new mines. The armistice found the United States with a large surplus of copper on hand, variously estimated at from 800,000,000 to 1,000,000,000 lb. In addition to this surplus in the United States, there were considerable stocks in the hands of the Allies and also immense quantities of unused brass. To. complicate the situa-tion still further, the reclamation of old brass shell cases and other used material was instituted, still further curtailing the demand for new material. This situation was- not appreciated at the time and it was assumed, that the Central Powers would immediately come into the market ? replenish their exhausted stock of copper. During the war Germany had torn down copper trolley wires and roofs, and melted up bronze bells, doors, statues and other copper-bearing material to reclaim this metal for military purposes. But Germany-was unable or at least unwilling immediately to replace her copper substitutes.

Jan 1, 1923