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By E. M. Norris

In the summer of 1917 it was decided to fireproof the main tramway hoisting shaft of the Anaconda Copper Mining Co. at Butte, Mont. The shaft has three hoisting compartments and one pump compartment; it is timbered with 12 by 12-in. (30.5 by 30.5-cm.) fir timber, and is 2475 ft. (754 m.) deep. Subsidence and displacement of the surrounding country rock had produced exceedingly heavy ground, and had carried the shaft out of line, in several places, by as much as 2 ft. (0.6 m.) displacement. Constant repairing and realigning of the timbers had been necessary in order to maintain clearance for the cages. The most feasible method of fireproofing, therefore, seemed to be to cover the timbers with a coat of concrete applied with the cement gun. The cement gun, which is operated by compressed air at ordinary mine pressures, feeds a mixture of sand and cement through a hose to a nozzle having a water connection. The mortar in fluid form is thus sprayed upon the prepared surface in thin layers, which can be built up to any desired thickness. The cement gun can be set up at any convenient point; satisfactory results have been obtained with the nozzle a distance of 500 ft. (152 m.) from the gun. The shaft was thoroughly overhauled and the timbers put in the best possible state of repair. Between the 1000 and the 1400 levels, much loose ground lay against the shaft timbers on the north side; this was breast-boarded back and a 30-in. (76-cm.) reinforced-concrete retaining wall was erected, leaving a space of 12 in. (30.5 cm.) outside the shaft timbers to allow for future movements of the ground. For convenience in handling men and materials; it was arranged to concrete the auxiliary hoisting-cage compartment and the pump compartment first, using the cages of the main hoist to serve the cementing crews. The auxiliary cage was then available while the remainder of the shaft was being concreted.

Jan 3, 1918

By David R. Mitchell

IF you are the owner of a mine, or a mine executive, or just an ordinary miner, and have a son about to go to college, do you urge him to take up mining engineering or do you try to dissuade him from such a course? If you live in the East, or the Middle West and your interest has been in coal rather than metal mining, the chances are that your young hopeful will be urged to do almost anything else rather than follow in father's footsteps.

Jan 1, 1939

By C. M. Brinckerhoff

Safety work in mining at the Andes Copper Mining Company, Potrerillos, Chile, is divided into three parts: (1) accident prevention, (2) fire prevention and protection, and (3) silicosis prevention and control. In the company's operations a steady campaign has been waged to reduce accidents and silicosis, and to prevent fires. In the following pages the safety record for the past twelve years will be examined and the methods explained by which the accident rate was reduced. Table 1 contains yearly data pertaining to the scale of the mining operations and the accident record. Block caving is used for all ore production and the tonnage mined has averaged 30,000 short tons per day for the past seven years. Table 2 classifies 44 fatal accidents which have occurred since 1935. The leading causes of accidental death are transportation of ore and use of explosives, followed in order by persons falling, falls of ground, and by asphyxiation. In the case of serious accidents, however, the greatest cause is falls of ground, followed, in turn, by ore transportation, persons falling, explosives, and asphyxiation. Accident Prevention The improvement of the safety record was obtained by a combination of many factors. In surveying the results at Andes for the past twelve years, it is possible to select some features of the present operation which have contributed to the accident record. The most important of these are: (1) good staff planning, (2) good equipment, (3) an alert safety organization, and (4) education and selection of the workmen. GOOD STAFF PLANNING As a result of good staff planning, the mine has good working conditions, namely, good ventilation, moderate temperatures and humidity, and good lighting. These factors not only reduce accidents but also improve efficiency. In haulage work, accidents have been reduced by having traffic move in a forward direction only. The latter makes possible a one-man train crew. Trains are centrally dispatched from loading drifts which are interconnected by signal lights, so that when a train in a lateral drift is given permission to move into a main line drift, a red light signals in all other lateral drifts in that area. Good track is maintained, and safety niches are required at all switch throws. Since safety niches were provided there have been no fatal or serious accidents at such points. All main line haulage drifts are illuminated by electric lights. Haulage drifts are kept clean. Adequate clearance has been provided between cars and between cars and the sides of the haulage drifts. The 4 1/2 ton Cranby type cars are equipped with roller bearings, and the 10 ton loconlotives with trolley shoes instead of wheels. These two features are responsible for improved ore haulage conditions and contribute to greater efficiency and safety.

Jan 1, 1950

By Charles H. Hopkins, J. M. Lohr

A COMMERCIAL development requiring a suitable alloy or alloys for sealing into various grades of glass made it desirable to have a more exact knowledge' of the expansion characteristics of the nickel-iron alloys than appeared in the published literature. The expansion of the more common [ ] types of glass under consideration varies from 2.7 to 10.4 X 10-s per degree Centigrade, over the range 20° to 400° C. For reference, the thermal expansion curves of several types of glass used for sealing by the lamp and tube industries are given (Figs. 1 and 2). This paper covers a detailed investigation of the thermal expansion of a series of nickel-iron alloys in which the nickel content varies from 30 to 70 per cent. Particular stress is laid on the alloys that contain between 40 and 55 per cent Ni.

Jan 1, 1938

By J. A. Danko, C. D. Pears, L. D. Schmidt, J. P. McGee

This paper describes a pilot plant built on the campus of West Virginia University by the Federal Bureau of Mines station at Morgantown, West Virginia, for gasifying pulverized coal in a mixture of oxygen and steam at pressures in the gasifier up to 450 psig. The work upon which this report is based was done under a cooperative agreement between the Bureau of Mines and the West Virginia University. The goal of Bureau of Mines work at Morgantown is development of a low-cost process for producing mixtures of carbon monoxide and hydrogen for the modified Fischer-Tropsch synthetic-fuel plants, which may be shifted to produce hydrogen for coal-hydrogenation plants. Gasification with oxygen offers the inducement that the gas may be generated at the pressure at which the gas will be utilized in a gas-synthesis plant. Since the optimum pressure for the synthesis appears to be about 400 psig, the gas should be produced at approximately 450 psig to take care of pressure drops in the purification train and control valves. If the gas is generated at the pressure of utilization, only the oxygen need be compressed, not the raw synthesis gas. This means a saving in power of 5 hp per hour per 1000 cu ft (CO + H2), which is approximately a saving of 4 cents per gallon of gasoline. For example, consider a commercial plant generating 18,000,000 standard cubic feet per hour of raw gas that would produce approximately 10,000bbl of liquid product per stream-day. Assume that this gas would be 85 pd CO + H2 and that the oxygen required would be 320 cu ft per 1000 cu ft of CO + H2. The following estimate is based on a commercial centrifugal compressor. If the 18,000,000 std cu ft per hour of synthesis gas is compressed to 450 psig with 74 pd efficiency, 107,000 hp is required for compression. If the synthesis gas is generated at 450 psig pressure, only

Jan 1, 1953

By R. H. Richards, A. E. Woodward

In Rittinger's Aufbereitungskunde of 1867 occurs (p. 195) the following table, which shows the rate of falling in water of fragments of minerals, of irregular shapes and of five different specific gravities: Diameter of particles in millimeters. 10 8 6 4 3 2 1 .5 Sp. Or. Velocity in meters per second. Auriferous silver, 16..9 .8 .7 .57 .5 .4 .29 .20 Galena,. ..7.5 .62 .55 .48 .39 54 .28 .20 .14 Pyrites,. ..5.0 .49 .43 .38 .30 .26 .22 .15 .10 Quartz,. ..2.6 -30 .27 24 .19 .17 .14 .10 .07 Coal,. ..1.3 .13 .12 .10 .08 .07 .06 .04 .03 With the purpose of extending this valuable table, to cover a larger number of minerals and sizes of each mineral, and of making it more easy of application in deciding upon a sieve-scale for any group of minerals, we have prepared the curves of velocity of fall of bodies in water shown in Plates I., XI., 111. and IV., using Rittinger's formula V= 2.44d(8 —1) where V= velocity in millimeters per second. " D = diameter of particles in meters. " 8 = specific gravity of the mineral. These curves embrace a large number of specific gravities, and for each specific gravity the relation of the diameter of the mineral particle to the velocity of its fall in water is given for all diameters and velocities used in practice. The second diagram is an enlargement of the lower left-hand corner of the first, the third bears the same relation to the second, and in like manner the fourth to the third. Plate I, therefore, covers the work of the coarsest jigging 2 inches in diameter; and under

Jan 1, 1890

By R. F. Shurtz

The author proposes a method used with some success on a magnesite deposit at Gabbs, Nev. He believes this procedure to be more sound than the blind practice of assigning uniform quality to large, soinetimes vast blocks or ore on the basis of analyses made of a small number of samples. When an ore deposit is evaluated on the basis of core sampling, questions always arise as to how much weight should be given the various sample grades and how the deposit should be divided into specific volumes assigned to corresponding grades. For a tabular deposit it is customary to use the so-called zone-of-influence. This zone is usually defined as the polygon in the plane of the deposit having corners halfway between a given core drill intersection and each of the nearest neighboring intersections. The volume of the ore in this polygon is assigned the analysis obtained from the drill core taken at the intersection. In massive or irregular deposits, volumes of various shapes may be defined with respect to the locations of drill core samples. These volumes may be regarded as three-dimensional zones-of-influence; the theory involved in assigning the values of the central analyses to the ore in the volumes is the same as that used in the case of a tabular deposit. The zone-of-influence method was used in evaluating one of the magnesite deposits at Gabbs, Nev.'-"

Jan 1, 1960

By Alan K. Burton

The individuality of off-highway trucks is largely determined by a number of truck components that include beds, frames, and suspensions. With the exception of tires, these features are generally standard equipment and the mine operator gets "what comes with the truck." With the advent of larger trucks, more complicated haul profiles, and sophisticated production control systems, a new truck support-component industry is rapidly appearing. The impact of the new options for off-highway trucks can be significant in terms of future fleet performance. Innovations include "Mesabi" radiator cores that reportedly offer greater engine temperature control; load indicators that insure proper truck loading at the shovel; and digital communication systems that signal truck status in the haulage cycle. An understanding of tires, brakes, and newly introduced support components is necessary before an optimum choice of truck can be made

Jan 11, 1975

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By AIME AIME

BLANDFORD. C. BURGESS, the new Chairman of the Industrial Minerals Division, took the advice of Horace Greeley in reverse-he turned his hack on San Francisco and the Golden Gate, and after a few side excursions toward Alaska and Arizona, hit the trail eastward, not for the lonesome pine, but for Spruce Pine, N. C. Born Aug. 10, 1893, in San Francisco. Burgess found his first job, shortly after finishing high school at 19, as axeman for a surveying party of engineers working for the Southern Pacific Co. in Oregon. Later he moved on up with them into Alaska. He liked the engineering profession, and decided to equip himself further so he entered the University of California in the fall of 1913. and for the next two years devoted himself to acquiring an education in civil engineering.

Jan 1, 1945

By Francis S. Peabody

This paper will not be a technical paper, because, although I have been in the business of mining and selling coal for 30 odd years, I am neither a mining engineer nor a practical miner. If I digress from time to time from my thesis, I must be forgiven by the expert engineers and highly technical gentlemen for whose proceedings this article is written. Waste of the wonderful store of power that the Creator placed at our disposal appears to have been the theory and the basis of our methods of extracting this power—waste from the time the coal is mined to the time it is consumed. Let us consider the prime causes of this wastage which has been the heritage of the coal mining industry. In the early days—about 150 years after " Stone Cole" was discovered in Illinois by Joliet and Marquette, and Father Hennepin noted a "cole" mine on his map—the mode of mining prevalent among farmer landowners, on whose property coal was discovered near the surface or outcropping on the hillsides, was to enter the seam by a shallow shaft or drift into the hillside and remove the coal by wedging it down until a large room was left. The roof without support would fall and our primitive coal operator would then sink another shaft or drive another drift. In these early operations only the large lumps were taken and all small pieces and screenings were left to be covered by the fall of the roof—in fact this condition prevailed in Illinois up to about 1885. Gradually the method of mining changed from the open chamber, to a single entry with rooms turned to the right and left. Later, as the demand for coal increased and the necessity for larger mines became evident, the "room and pillar" system grew to be regular practice. The room and pillar system of mining consisted of starting away from the shaft bottom or drift mouth, with a pair of entries usually 12 to 15 ft. wide and about 30 ft. center to center, one entry being used for the air intake and the other for the return air and haulage; from each entry rooms or chambers were turned, narrow at the mouth and widening

Jan 1, 1918

By D. W. Frommer

Processing nonmagnetic taconites by selective flocculation-desliming and flotation requires large volumes of water. If impounded without treatment, these off-process waters require excessively large areas for containment. To discharge the waste water into natural waterways would contribute to stream pollution and likely would not be permitted. In U.S. Bureau of Mines experiments conducted in the Twin Cities Metallurgy Research Center's 900-lb per hr pilot plant, approximately 85% of water requirements for the flotation-based treatment of a Michigan nonmagnetic taconite were met by reclaimed water. Water reclamation of the off-process streams from flotation was accomplished by controlled additions of lime, sodium carbonate, and a synthetic flocculant to reduce turbidities to 51000 ppm equivalent SiO*, while maintaining a Ca(II) content of =16 ppm in the finished effluent. Flotation concentrates of good quality were obtained using the reclaimed water. The cost of chemicals used in water reclamation was approximately equal to the savings in flotation reagents attributed to recycling of the water. Water quality is perhaps as important to flotation as are the reagents used. The character of water is extremely variable, depending on whether the source is a well, lake, or stream, upon the season and temperature, upon prior use, and upon the character of the watershed. All of these factors influence the water hardness and the quantity of other dissolved inorganic salts, turbidity, dissolved and suspended organic matter, dissolved gases, and pH. Frequently, the differences in water quality can measurably influence flotation selectivity, often to the point of spelling success or failure. Water hardness is particularly troublesome in flotation systems employing fatty acids, but other unrecognized constituents may also contribute to peculiarities in flotation behavior. Furthermore, a given water source may be entirely satisfactory in one flotation system, but entirely inappropriate in another. In recent years, society has given more attention than formerly to water use, even in areas where water is plentiful. However, both the demands of a growing population, with increased per capita needs, and also those of industry must be met. As a result, riparian rights must be negotiated with the appropriate government agency or agencies so that consumption of water is often allocated or otherwise controlled. Furthermore, the disposal of off-process industrial and domestic water is coming under the increasing scrutiny of these same governmental units. In these respects, the mineral industry is no exception, so that conservation, water reclamation, and reuse may be expected to assume increasing importance. In 1932, it was stated that, "water reclamation is generally more expensive than the economy in water and reagents resulting from its use. It is employed only if it is urgent to save water."' The economic aspects of this statement may still be true, but the unrestricted use of water is becoming less and less an option of the user. The purpose of this paper is to discuss an investigation conducted by the U.S. Bureau of Mines (USBM), in which water reclamation and treatment were undertaken to develop procedures for, and to assess the effects of, water reuse on the flotation treatment of nonmagnetic taconites. This investigation is a logical extension of previous work described by the author and associates at the Twin Cities Metallurgy Research Center employing selective flocculation-desliming and anionic flotation of silica from low-grade, nonmagnetic iron ores2,8 The requirements for the selective flocculation-de-sliming and the anionic flotation of silica processes determined the direction of the investigation. Previous studies had indicated that pH levels of about 11.0 and 11.8 were required for selective flocculation and flota-tlon, respectively. The calcium content of the water was believed to be of importance in both of these operations, and since Minneapolis tap water with a Ca(I1) content of about 16 ppm had been successfully used, an attempt was made to reclaim the process water at an equal level of dissolved calcium. Additionally, the objective was to nullify or effectively limit the effects of dispersants and fatty acid residuals from prior stages of processing. Last but not least, the system of water treatment had to have the capability of reducing turbidities to workable levels. With about half of the effluent being derived from the selective flocculation-desliming step and carrying about 25,000 ppm of highly dispersed, suspended fines, this last objective appeared formidable at the outset. Procedures for water reclamation were derived, in part, from well-known mineral dressing practices, from past observations and investigations, and from concepts contained in various pertinent publications on water treatment."c These procedures involved: 1) flocculation with lime and poly electrolyte-type flocculants, 2) lime-soda-ash softening, 3) chemical precipitation, and 4) mineral surface adsorption. Control was exercised at various stages of water reclamation by frequent measurements of Ca(11), pH, and turbidity.

Jan 1, 1971

By Charles Will Wright

THE economy as well as the living standards of a country depends largely upon adequate supplies of raw materials at reasonable prices. Geological and climatic conditions responsible for the occurrence of commercial minerals and the growth of raw agricultural and forest products are so varied in different parts of the world that no country can be self-sufficient. The world is rapidly shrinking because of improvement in transportation facilities, but populations are steadily increasing and so are the needs for a greater production of raw materials and for their more equitable distribution. Too many barriers between nations prevent free access to the main sources of supply of the raw materials vital to neighboring nations and it is just such barriers that are often the cause of friction and even war. The International Trade Organization is doing a commendable job in helping to. break down barriers and to reduce friction, and we should now find ways to facilitate access to these sources of supply and encourage more efficient and greater production of the raw materials so urgently needed throughout the world. This cannot be done by individual selfish governmental agencies, influenced as they are by politics, nor by socialistic governments attempting to nationalize industries. It can be done efficiently only by the free individual through private enterprise properly protected against adverse governmental action. This protection can be supplied by the World Bank.

Jan 1, 1948

By L. W. Blau

RESEARCH work in exploration and production was further reduced during 1943 owing, partly, to difficulties in the acquisition of apparatus and, principally, to the exodus of research men to governmental agencies conducting research and to the armed forces.

Jan 1, 1944

By J. A. SINGMASTERN

THE New Jersey Zinc Co.'s vertical retort plants are believed to have been in continuous operation through the whole year. At Palmerton metal purer than that made from the same ore in the old plant has been produced and has been able to compete successfully with very pure electrolytic metal in die castings and other special uses where extreme purity is worth money. A new plant using this process was put in production by the Grasselli Chemical Co. at Meadowbrook, W. Va., early in the summer. It is said to be working very satisfactorily. It is understood that this plant will not produce high-grade metal. Two patents issued to Bunce and Mahler (Nos. 1,749,126 and 1,749,127) describe a method of operating the vertical retort so that practically lead-free metal can be obtained from ores containing consider- able lead. The zinc vapors and other gases leaving the reaction zone are filtered through a column of material maintained at a temperature high enough to prevent much zinc from condensing. The filtering medium is either periodically or continuously changed.

Jan 1, 1931

By Herbert Clark Hoover, Lou Henry Hoover

PREVIOUSLY I have dealt with the methods of separating silver from copper. There now remains the portion which treats of solidified juices ; and whereas they might be considered as alien to things metallic, nevertheless, the reasons why they should not be separated from it I have explained in the second book. Solidified juices are either prepared from waters in which nature or art has infused them, or they are produced from the liquid juices themselves, or from stony minerals. Sagacious people, at first observing the waters of some lakes to be naturally full of juices which thickened on being dried up by the heat of the sun and thus became solidified juices, drew such waters into other places, or diverted them into low-lying places adjoining hills, so that the heat of the sun should likewise cause them to condense. Subsequently, because they observed that in this wise the solidified juices could be made only in summer, and then not in all countries, but only in hot and temperate regions in which it seldom rains in summer, they boiled them in vessels over a fire until they began to thicken. In this manner, at all times of the year, in all regions, even the coldest, solidified juices could be obtained from solutions of such juices, whether made by nature or by art. Afterward, when they saw juices drip from some roasted stones, they cooked these in pots in order to obtain solidified juices in this wise also. It is worth the trouble to learn the pro- portions and the methods by which these are made. I will therefore began with salt, which is made from water either salty by nature, or by the labour of man, or else from a solution of salt, or from lye, likewise salty. Water which is salty by nature, is condensed and converted into salt in salt-pits by the heat of the sun, or else by the heat of a fire in pans or pts or trenches. That which is made salty by art, is also condensed by fire and changed into salt. There should be as many salt-pits dug as the circumstance of the place permits, but there should not be more made than can be used, although we ought to make as much salt as we can sell. The depth of salt-pits should be moderate, and the bottom should be level, so that all the water is evaporated from the salt by the heat of the sun. The salt-pits should first be encrusted with salt, so that they may not suck up the water. The method of pouring or leading sea-water into salt-pits is very old, and is still in use in many places. The method is not less old, but less common, to pour well-water into salt-pits, as was done in Babylon, for which Pliny is the authority, and in Cappadocia, where they used not only well-water, but also spring-water. In all hot countries salt-water and lake-water are conducted, poured or carried into salt-pits, and, being dried by the heat of the sun, are converted into

Jan 1, 1950

By D. L. Ritter, N. J. Grant, B. C. Giessen

Forty-six alloys covering the complete concentration range of the Nb-Rh system were examined by metallographic and X-ray methods; solubility limits of terminal ad intermediate phases and transformation and solidus temperatures were determined. There are nine intermediate phases; two eutectic, five peritectic, two eutectoid, and three peritectnid reactions were encountered. The experimental methods and the constitution diagram are described in Part I. Part 11 deals with the crys -tallography and relation of the intermediate phases. The Nb-Rh system was selected for study for three reasons: first, it was hoped that the anticipated similarity of the Nb-Rh phase diagram to the Ta-Rh1 and Ta-Ir2 diagrams treated earlier (see also Ref. 3) and to the Nb-1r4 diagram treated simultaneously would allow a crystallographic and comparative evaluation of the intermediate phases in these related systems; second, it was likely that the previous experimental methods could be carried over without significant change; and third, it was expected that the interesting mechanical properties of the a1 phase of the Ta-Rh and Ta-Ir systems would also be encountered in similar phases of the new systems, and that these phases could be investigated more closely. The phase diagram had not been treated fully in the literature; only brief surveys of some intermediate phases were found.5-7 It was attempted to cover the details of the phase boundaries, Part I, as well as to present the crystallography of the intermediate phases, Part 11. EXPERIMENTAL METHODS Specimen Preparation and Treatment. The following high-purity starting materials were used. 1) Niobium powder, 99.8 pct pure, mesh size -60 +200, from Wah Chang Corp., Albany, Ore. The analysis furnished by the supplier showed the following major impurities: Ta, <500 ppm; Zr, <500 ppm; 0, 300 ppm; Fe, 280 ppm; Ti, <I50 ppm; W, Si, <100 ppm; C, N, 30 ppm; all other impurities, <30 ppm. 2) Rhodium powder, ground sponge, 99.9+ pct pure, mesh size -80, from J. Bishop Co., Malvern,

Jan 1, 1964

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Jan 1, 1938

By Joseph A. Holmes

(Pittsburg Meeting, March, 1910.) AT the Mining Experiment Station of the U. S. Geological Survey, in Pittsburg, an investigation of the process of combustion is being carried on in a specially-designed furnace having an unusually long combustion-chamber. This work is conducted by a committee consisting of H. Kreisinger, mechanical engineer; Dr. J. C. W. Frazer, chemist; and Dr. J. K. Clement, physicist. The problem, essentially one of physical chemistry, is extremely interesting to all who are concerned in the burning of coal or the construction of furnaces. The main object of these experiments is to ascertain the relation between the amount of the volatile combustible driven from the coal and the combustion-spaces necessary to burn it completely. Our best steam-coals vary from 15 to 45 per cent. of fixed combustible. The greater part of the latter is burned on the grate; but the volatile combustible leaves the freshly-charged coal and must be burned in the combustion-space. If this space is not large enough, the volatile combustible will leave the furnace only partly burned, and the result will be a considerable heat-loss and a smoky stack. Strictly speaking, the factor which determines the completeness of combustion of the volatile matter, after it has been mixed with a certain amount of air, is the length of time the mixture is allowed to remain in the combustion-space; but this length of time depends on the extent of the space itself. Let us suppose, for instance, that, when a given coal is burned at a certain standard rate, the volume of the volatile combustible driven off per second is 10 cu. ft. Adding, say, 10 cu. ft. of air, the volume of the resulting burning mixture is increased to 20 cu. ft. If the combustion-space is 20 cu. ft., the burning mixture will stay in it,

May 1, 1910

By AIME AIME

DURING my senior year at college a professor said to his class that a student who failed to obtain a passing grade in that certain subject could not graduate with his class and that his diploma would be withheld until the failure had been removed. In that way for four years a diploma had been made to appear to us as the objective in attending college. We grew up to think that a diploma was the means of providing employment; consequently, a diploma could not be jeopardized and, heeding the professor&apos;s remarks, we resolved to put forth special efforts and make good. During one of our class periods this same professor digressed briefly from his prepared lecture in metallurgy and advised the boys to join a national technical society. He spoke with earnestness and sincerity. After class we students discussed those remarks. We concluded that the professor had not convinced us. We were already members of technical societies at school and saw no need of joining another. In fact, even up to that moment a society had never indicated to any of us that joining it had any relation to landing a job, and we were fully aware that joining another society meant paying out more money, for fees and dues, which money we then needed to buy food, clothes, and perhaps a new slide rule. Furthermore, we knew that joining a society was not part of the curriculum and, therefore, the diploma could not be withheld if we did not follow the advice. Our conclusion was that we would not join but would concentrate on our work that assured getting our diplomas.

Jan 1, 1936