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By Lyman H. Hart

Some of the significant facts that will affect the supply and demand for metals during the next few decades are given in this presentation. This is important because the only hope for intelligent guidance of the national destiny depends upon presenting a consistent and accurate picture of basic problems. The metals problems now fomenting are only slightly less important than those precipitating the energy crisis. Admittedly, it is difficult to generate serious thinking on this subject when the world is enjoying easy access to all metals at modest prices. This is possible because of the current world capacity to over-produce, and United States credit standing, at least up to now, has permitted buying freely on world markets. Although both these vital conditions are subject to change, there is little doubt that if the U.S. financial house is kept in order, we can muddle along for a while on a generally expanding, metal-based economy. But should our credit standing be permitted to collapse, or should there be a tightening of world supplies of metals, the nation would immediately find itself in trouble. Unfortunately, clear signals indicate that both of these conditions are threatening. The metal age started many centuries ago, and with the exception of the noble metals, barely a dent was made in world reserves until the present century. Then, as a result of the pioneering ideas of D. C. Jackling and Paul Gemmel, and the willingness of the McNeals, Penroses and Guggenheims to back ideas with money, a virtual explosion of metal production was propagated throughout the world. It is important to examine the causes of this expansion to determine whether it is a trend that may be sustained or whether it is anomalous and therefore attributable to circumstances not likely to be repeated or continued. This period of plenty is a once only phenomenon. Vast metal resources have been sitting around here for eons, while the world was unaware of or unconcerned with their existence. But when population began to grow exponentially, and ways were discovered to use metals to develop machines to do the work of men and animals, the exploitation of raw materials increased at a staggering rate. And now the realization is dawning that minerals are a wasting asset and are beginning to present supply problems that should begin to be of concern. Two main factors have caused this condition; One is the population growth, and the other is the expanding demand for irreplaceable raw materials as nations become more developed. A group of systems analysts, at Massachusetts Institute of Technology in a book Limits of Growth, has warned that at present rates of population growth and resource consumption together with environmental constraints, a point will be reached within 100 years when the world population will become totally incapable of supporting itself. The authors conclude that this end result could be avoided if a planned equilibrium could be established at a lower rate of population growth and resource consumption. Fortunately, there is evidence that the population problem is moderating. The growth rate in the U.S. has decreased from 2.3% per year in 1900 to 1.1% in 1970. But, even at the lower rate, population will double in 70 years. The world rate is greater, and if it continues, world population will rise from 3.5 billion to nearly 10 billion in 50 years. All of these projections are questionable because the dangers of growing population are now being recognized. A marked decrease in growth rate has been noted in the more developed countries-in fact, it was zero in the U.S. in 1972. The other factor affecting consumption of raw materials is the increase in per capita use of most items as individuals become more affluent. Many economists believe that developing countries such as China and India represent a vast future potential market; but there are others, with whom the author agrees, who believe over-population tends to stagnate development. Nevertheless, it is a fact that vigorously developing societies increase consumption at a rate far above their population growth rate.

Jan 1, 1974

Discussion of the paper of Leonard Waldo, presented at the New York meeting, February, 1911, and printed in Bulletin No. 76, April, 1913, pp. 555 to 559. HERBERT HAAS, San Francisco, Cal. (communication to the Secretary*):-Referring to Leonard Waldo's paper on Temperature Conversion Tables, I believe the following formulas can be easily remembered, and by reference to the thermomter scales of Fahrenheit (F.), Celsius (C.), and Réaumur (R.), their derivation and relation are at once understood n° F. = 4/9 (n°-32) R. = (n°-32) C. n° C.=(9/5 n° + 32) F. = 4/5 n° R. n° R.=(9/4 n° + 32) F. = 5/4 n° C. As the freezing point of water is at 32° F., corresponding to 0° C. and 0° R., 32 must be deducted from the Fahrenheit degrees when these are converted into C. or R. degrees. Likewise, to C. or R. degrees converted into F. degrees 32 must be added. The fractions 4/9 and 5/9 and their reciprocals are obtained thus:

Jan 5, 1913

By W. A. Knoll

THE sinking of a new shaft at the Newport mine, Ironwood, Mich., was started in May 1931 and completed on Aug. 3, 1932. During this period, 2665 ft. of shaft in granite was completed, at, an average advance of 6 ft. per 24 hr. In November of 1930, experimental work was con-ducted oil 65 ft. of shaft through 22 ft. of overburden where ledge was encountered, and this work convinced the operators that all drilling could be clone with vertical holes with the blasting relief to a 4 ½ -in. hole drilled vertically in the center of the cut. The reason for not adopting the conventional V-cut was to get away from damaging the shaft steel. The blasting in a V-cut naturally tends to throw the broken material upward, but with the cut as adopted the relief was in a horizontal direction to a large extent. The scheme of mucking employed made it necessary to keep the steel shaft sets within 511 ft. of the bottom at all times. This could not be clone without serious damage with a V-cut when blasting in granite. The cut as drilled also broke 10.53 ft. per round of drilling, which was considerably more than the experience with a V-cut in the sane material. The extended cut reduced proportionately the delays for charging, blowing smoke and handling drilling equipment in the shaft. Another advantage was the ease in drilling vertical over inclined holes, and the preservation of the proper spacing of the drill holes at the bottom of the cut.

Jan 1, 1938

By Galen H., Clevenger

THE first and most important thing that affects the majority of lead producers in the Rocky Mountains and the western mining regions is that zinc in an ever-increasing degree is inseparably associated with the lead. The Sullivan mine in British Columbia is not only a great lead mine, but also a great zinc producer. This mine could not reach its full measure of productiveness, nor was its great value established, until a reasonably good physical method was developed for separating the lead from the zinc prior to the recovery of each metal by an appropriate process. The western metallurgists and also metallurgists in other parts of the world have long faced the lead-zinc problem. This for many years stimulated a great deal of experimental work looking toward the development of a physical or chemical method of separating lead and zinc in concentrated form from complex ores. Space will not permit of going into this work in detail, but it is sufficient to say that with but few exceptions the western lead smelter continued smelting charges many times too high in zinc for the most efficient operation of lead furnaces. The zinc was not only lost, as the zinc content of many a western slag dump bears mute testimony, but the cost of lead smelting was increased by reason of the presence of the zinc. The owner of the ore under these circumstances could be paid nothing for the zinc and, indeed, he was penalized for it

Jan 1, 1926

By Clyde E. Williams, V. N. Krivobok, C. H. Herty

THE extreme effect of the depression on the steel industry is well illustrated by the fact that the amount of iron ore shipped from the Lake Superior district was the lowest in 47 years. Something over 3 1/2 million tons of ore was shipped in 1932, compared with over 65 million tons in 1929. It is not surprising then that practically no advancement in iron ore beneficiation has been made. Improved research and development work, however, was continued at the Michigan College of Mines in the survey of the lean iron ores of Michigan with reference to the possibility of their concentration; at the University of Minnesota in its large-scale jigging studies of Lake :Superior ores; and at Battelle Memorial Institute in the development of improved jigging of iron ores. The successful operation of the new Hadsel mill at a gold property in California gave such low grinding costs on a quartz ore as to give much encouragement in the iron ore industry for a cheapening of the cost of grinding lean ores for concentration.

Jan 1, 1933

By John W. Langley

In the summer of 1888 it was the fortune of the writer to present the subject of the desirability of establishing a set of samples of steel, which should be analyzed with extreme care, in order that they might become standards to which scientific and commercial analyses of iron and steel could be subsequently referred; also, that greater uniformity in the results of analyses might be brought about, since these standards would bear towards analytical methods somewhat the position which the original units of weight and length, the gramme and meter, or the pound and yard, preserved in Paris, London, and Washington, do to the mechanical arts. The plan met with hearty co-operation, with the result that com mittees were appointed in Sweden, Germany, France, England, and America, to receive the material and to see that the necessary au alyses were executed. In England, the British Association for the Advancement of Science appointed a committee, which subsequently published a plan

Jan 1, 1891

By E. Buehler

THE floating zone refining technique'-4 has provided a new experimental approach to the purification of many reactive materials especially those with high melting points. The method avoids the use of a crucible and at the same time it provides a favorable environment at the freezing interface for the growth of single crystals. The physical properties of refractory metals such as molybdenum and tungsten have long been of interest from a fundamental point of view and because of their widespread use in vacuum electronics. Single crystals are often desirable for study. Prior to the introduction of zone melting, single crystals of these metals were usually grown in the solid state at temperatures slightly below the melting point by strain annealing3 or by the application of a moving temperature gradient. Such techniques were mostly successful on wires, ribbons, or rods having small cross-sectional areas. Recently, larger samples of tungsten, molydbenum, tantalum, columbium, and rhenium were melted by a floating zone process which employed electron bombardment heating.10 Single crystals were obtained in many instances

Jan 1, 1959

By James R. Garvey

Coal reserves of the United States are enormous. Considered on the basis of proven reserves, and compared with reserves of other fuels, coal constitutes 88% of the proven recoverable energy resources of the nation. That includes oil, natural gas, oil shale and uranium used in burner reactors. Considered on the basis of "ultimate recoverable reserves", that is, those resources which can be extracted with improved technology and at somewhat higher costs and including those resources of all fuels not yet found but believed to be waiting discovery, coal still accounts for 75 % of the national total. In the U.S. hydrocarbon economy about 76% of the total energy needs are supplied by petroleum and natural gas and about 20% by coal. Domestic resources of oil and gas are running dangerously low, and assuming a continued hydrocarbon economy, the 20% supplied by coal may increase dramatically. Nearly a quarter of the petroleum consumed in the U.S. is imported. Natural gas is in shorter supply than petroleum and the problems of its importation seem at least as difficult as those of oil.

Jan 6, 1972

By Donald O. Rausch, Adolph Soderberg

4.1-1. Introduction. Open pit mine planning must be correlated to all phases of a mining operation. The factors that must be considered in planning an open pit mine are numerous and must reflect the characteristics and surrounding conditions of a particular orebody. Therefore, only an outline of the subject can be presented here to aid the planning engineer in pointing out procedures that are generally applicable to pit design. The presentation is directed toward the planning of an open pit copper mine, but the procedures discussed should apply in most cases to any type of metallic or nonmetallic orebody. In planning an open pit mine the pertinent elements that must be included are: assays, geology, tonnage and areal extent of ore reserves, topography, mining equipment, economic factors of operating costs, capital expenditures, profit, types of ore, pit limits, cut-off grade, stripping ratio, rate of production, pit slopes, bench heights, road grades, ore metallurgical characteristics, hydrological conditions, property lines, and marketing considerations. 4.1-2. Sampling. Sampling must be adequate to classify the various materials to be removed from an open pit mine. Drillhole assays are plotted on sections and plan maps with ore and waste blocked out by selected bench intervals. These maps are prepared to depict the geological and other conditions existing on each mine level. The confidence factor to be applied to sampling and the resultant assays values must be given careful consideration. In uniformly disseminated de- posits, the confidence factor will approach unity. For widely fluctuating values, the average assay values should be discounted somewhat, the degree depending upon such factors as drillhole spacing. Such samples, at best, represent a very few pounds for hundreds to thousands of tons of ore

Jan 1, 1968

By S. H. Ash

SAFETY records have improved in all branches of the mineral industry. While annual production was rising from $2 billion in 1910 to nearly $12 billion in 1950, fatalities decreased from 3539 in 1911 to 974 in 1951. And, although safety progressed under hand methods, it progresses more effectively under modern, mechanized methods. But powered devices such as electric or diesel units can in themselves create new hazards unless proper safeguards are instituted. Health Use of diesels underground is increasing. Diesel equipment has been used in construction of 34 tunnels in 16 states in the last 17 years. The 77 noncoal underground mines in 24 states using diesels underground range from mines with large open stopes to those with narrow veins, so that the ventilating problems represent a good cross-section of U. S. mining. The U. S. Bureau of Mines has received no report that anyone has suffered ill effects from the proper use of properly maintained equipment. The necessity for correct maintenance and use of satisfactory exhaust conditioners cannot be emphasized too strongly.

Jan 2, 1954

By William B. Plank

Mineral engineering student enrollment in U. S. and Canadian schools for 1955-1956 is 11,408, an increase of 11 pct more than last year. The undergraduate and graduate engineering students in both countries number 251,765, or 13 pct more than last year. These figures have been released by the American Society for Engineering Education. The largest increase in the undergraduate mineral engineering group in the U. S. is in ceramic engineering, with 676 students, or 31 pct more than last year. The next largest increase is in mining which shows a gain of 13.5 pct, compared with a gain of 13.1 pct in the total undergraduate enrollment in all engineering schools approved by the Engineers' Council for Professional Development. This figure is 190,355, which approaches the prewar figure. Metallurgical undergraduates increased 13 pct; geological, 5 pct; and petroleum, 2.5 pct. This year the figures for geophysics are combined with those for geological engineering, making a total of 1610 undergraduates, of which the geophysicists comprise about one seventh.

Apr 1, 1956

By Paul F. Kerr, Raymond F. Robinson

Uranium mineralization occurs in the footwall of the Sunshine vein from the 2900 to the 3700 level. Veinlets of uraninite associated with pyrite and jasper have been so extensively divided and recemented that units more than a few feet in length are seldom observed. The wall rock is St. Regis quartzite of the Belt series. The age of the uraninite, on the basis of isotopic analyses, is 750 * 50, which agrees with geological data suggesting that phases of the Sunshine mineralization are pre Cambrian. THE Sunshine mine in the Coeur d'Alene district, Idaho, is well known for its silver-bearing veins but prior to the summer of 1949 had not been recognized as a possible source of uranium. At that time, during a geiger counter reconnaissance by T. E. Gillingham, R. F. Robinson, and E. E. Thurlow, high radioactivity was noted and radioactive specimens were collected from the footwall of the Sunshine vein.' The detection led to the identification of uraninite-bearing veins, since explored jointly by the Atomic Energy Commission and the Sunshine Mining Co. After the occurrence was noted, the geology of the uranium deposit was studied by the Sunshine staff, and a laboratory examination of the ores was conducted at Columbia University. Several types of laboratory work were undertaken. Differential thermal curves were made of selected siderite samples and results from many more were secured through the work of Mitcham.2 X-ray diffraction and X-ray fluorescence analyses were employed on uraninite, jasper, and siderite. Chemical analyses were made through the cooperation of the Division of Raw Materials of the Atomic Energy Commission. General Geological Features Several silver-bearing veins cut the overturned north limb of the Big Creek anticline as mapped by Shenon and McConne1,³ while the Osburn fault, a long-recognized regional feature about a mile away, marks the north boundary of the Silver Belt. The Sunshine vein, Fig. 1, has a south dip more or less parallel to the 60" axial plane of the fold and cuts rocks of the Belt. Series, starting with the Wallace formation near the surface, continuing downward through the St. Regis formation, and probably extending into the Revett quartzite which lies below the bottom or 3700-ft level. The limb of the anticline is locally modified by secondary folds, one being prominently exposed in the uranian area along the Jewel1 crosscut near the Sunshine vein. Crumpling of the limb resulted from compression which formed the anticline and probably preceded the faults in which the vein deposits accumulated. Evidence of drag along these faults points to reverse movement in the uranium-bearing area and elsewhere. This is true of major faults in the mine workings, and the majority of faults which can be mapped, as pointed out by Robinson.' The St. Regis formation, as measured in the mine, appears to have an initial thickness of some 2000 ft, but the apparent thickness due to thickening during folding is some 3400 ft. Along the Sunshine vein the purple and green rocks characteristic of the Wallace formation in the nearby Military Gulch section p. 37 of ref. 5) have been completely bleached because of introduced sericite. Hydrothermal solutions acting on the wall rock have substituted for the original color a pale greenish cast, although no pronounced mineralogical change has resulted, as Mitcham has observed.' The silver and the uranium depositions appear to belong to distinct epochs resulting from several periods of emplacement. Likewise, multiple periods of deformation account for the faulting. Uraninite is generally associated with silicification, while silver . mineralization accompanies carbonate veins. Rarely, uraninite may be found in a matrix of siderite. Ordinarily uraninite formed prior to ar-gentian tetrahedrite. Where clusters of veins form a stockwork, uraninite-jasper veins often favor one trend while tetrahedrite-siderite veins favor another. During deformation, brecciation of the St. Regis quartzite provided openings between broken rock fragments for precipitation from vein-forming solutions. Fractures due to major breaks were filled during the first stages of vein formation, while later deformation displaced the first veins and provided new channels along which further mineralizing solutions proceeded. The uraninite veins, as the first formed, have suffered fracturing, displacement, and segmentation. Uranian vein segments uncut by faults and more than a few feet in length are rare or nonexistent. Siderite veins are more massive and often extend without a break for tens and even hundreds of feet. In general they show much less segmentation. While the siderite is usually later, there is an overlap in the periods of deposition, some earlier siderite veins being extensively segmented in much the same way uraninite veins have been broken. Vein silica is more extensively distributed than the uranium and iron mineralization it carries. Along the vein course concentrations of uraninite frequently fade away and barren white quartz continues, the transition often occurring within a few feet along strike or down dip. An example appears on the 3700-ft level where a uraninite vein, see Fig. 2a,

Jan 1, 1954

By R. J. Mellen

In a chloridizing roasting and cyanida-tion plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42-in. rolls, the latter in closed circuit with two Tyler type 400 screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼-in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the rushed Ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1947

By R. J. Mellen

In a chloridizing roasting and cyanida-tion plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42-in. rolls, the latter in closed circuit with two Tyler type 400 screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼-in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the rushed Ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1947

Reports on process control computer projects at the Kamioka Plant are summarized from the Japanese technical literature (see references). Fluidized Bed Roaster and Sulfuric Acid plant1 A small process control computer (Yokogawa CCS-200 with an OKITAC 5090P 8K central processor) was installed in the Kamioka smelter in the summer of 1964. A year-and-one-half installation project followed, which successfully brought the roaster and acid converters under computer control. The process plant consisted of two 130 RPD fluosolids roasters and two Monsan to acid units. The roasters are charged with lead-zinc concentrates with SO2 from the roaster processed to sulfuric acid. The computer system provides the following functions: 1) log data from the plant--the input is taken from 50 analog signals, 130 high-low digital alarm points, and 50 analog high- low alarm points; 2) set-point control on 18 control loops; 3) direct digital control (DDC) on eight control loops for which conventional controllers were difficult to adapt due to abnormal loop characteristics or delay times; 4) execution of a safe shut-down procedure

Jan 1, 1969

By R. J. Mellen

IN a chloridizing roasting and cyanidation plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42in. rolls, the latter in closed circuit with two Tyler 400-mesh screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼ -in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes ,were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the crushed ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1945

By L. A. Willey, William L. Fink

TWO of the aluminum-alloy systems previously studied—the aluminum-titanium and the aluminum-chromium—exhibit a peritectic reaction at the extreme aluminum end of the diagram. Preliminary work indicated that the aluminum-zirconium and certain other aluminum systems were similar in this respect. The present paper, the twentieth of a series from the Aluminum Research Laboratories, reports the results of a recent investigation of equilibrium relations in aluminum-zirconium alloys of high purity, which confirm the preliminary observations. Previous Investigations There has been only one investigation of the aluminum end of the aluminum-zirconium system. Sykesl prepared and studied alloys containing 0 to 70 per cent Zr. The lowering of the freezing point of aluminum by the addition of zirconium was found to be very little, if any (less than 2" C.). He placed the solid solubility of zirconium in aluminum at room temperature between 0.57 and 0.97 per cent. A few other investigators have attempted to determine the composition of the intermetallic compounds occurring in this system. Honigschmid2 obtained the compound ZrAla as a by-product during the preparation of ZrSi2. He made no attempt to prepare other aluminum-zirconium compounds. Weiss and Newmann3 prepared an alloy by heating aluminum with a double fluoride of potassium and zirconium. When this alloy was treated consecutively with caustic soda and hydrochloric acid, there was a residue, which contained approximately 72 per cent Zr and 28 per cent A1 (Zr3AI4). Marden and Rich4 prepared an alloy in the same way, treated it with caustic soda, and obtained a residue of silver-white crystals, which contained 70 per cent Zr and 30 per cent A1 (Zr3A14). Alloys Materials Used.—Electrolytically refined aluminum was employed for thc preparation of all alloys used in this investigation. The chemical

Jan 1, 1939

By William L. Fink, L. A. Willey

TWO of the aluminum-alloy systems previously studied—the aluminum-titanium and the aluminum-chromium—exhibit a peritectic reaction at the extreme aluminum end of the diagram. Preliminary work indicated that the aluminum-zirconium and certain other aluminum systems were similar in this respect. The present paper, the twentieth of a series from the Aluminum Research Laboratories, reports the results of a recent investigation of equilibrium relations in aluminum-zirconium alloys of high purity, which confirm the preliminary observations. Previous Investigations There has been only one investigation of the aluminum end of the aluminum-zirconium system. Sykesl prepared and studied alloys containing 0 to 70 per cent Zr. The lowering of the freezing point of aluminum by the addition of zirconium was found to be very little, if any (less than 2" C.). He placed the solid solubility of zirconium in aluminum at room temperature between 0.57 and 0.97 per cent. A few other investigators have attempted to determine the composition of the intermetallic compounds occurring in this system. Honigschmid2 obtained the compound ZrAla as a by-product during the preparation of ZrSi2. He made no attempt to prepare other aluminum-zirconium compounds. Weiss and Newmann3 prepared an alloy by heating aluminum with a double fluoride of potassium and zirconium. When this alloy was treated consecutively with caustic soda and hydrochloric acid, there was a residue, which contained approximately 72 per cent Zr and 28 per cent A1 (Zr3AI4). Marden and Rich4 prepared an alloy in the same way, treated it with caustic soda, and obtained a residue of silver-white crystals, which contained 70 per cent Zr and 30 per cent A1 (Zr3A14). Alloys Materials Used.—Electrolytically refined aluminum was employed for thc preparation of all alloys used in this investigation. The chemical

Jan 1, 1939

By CHARLES HOPPER

In preparing the Steep Rock Lake iron ore body for mining, it was necessary to drain Steep Rock Lake. Using diamond drills, a cut 1800 ft long, 100 ft wide, and maximum depth of 95 ft amounting to 300,000 cu yd of material, was excavated in Greenstone to provide a drainage channel. This was done under the general contractor C. A. Pitts, of Toronto, who in turn employed a diamond drill contractor, Boyles Bros., to do the drilling. The method of excavation was selected be- cause a clean square cut and grade could be maintained in the bottom of the channel. Bootlegged holes would have left blocks of rock in the corners which would have created a situation which would have become progressively worse as the open cut advanced from one side of the ridge to the other.

Jan 1, 1949

By Douglas Bunting

IN THE anthracite fields of Pennsylvania, mine fires occur with more or less regularity and their existence is an ever-present hazard in coal mining. In all probability 90 per cent. of the mine fires can be ascribed directly or indirectly to the ordinary miner's open lamp. Other causes may be smoking, electrical installations, gas explosions, gas feeders, and the communication of fire to the outcrops from ash dumps, culm banks, timber and brush fires. Fires in pump rooms, engine rooms, and barns have been largely eliminated as a result of the Act of .Assembly, approved June 15, 1911, which provides that all inside buildings, including engine rooms, pump rooms, barns etc. shall be constructed of incombustible materials. Smoking of pipes, cigars, and cigarettes should be prohibited in the mines and too much care cannot be exercised in preventing fires by enforcing all rules designed for that purpose. Mine fires that involve underground fighting are invariably dangerous, and all fires, irrespective of the method pursued in extinguishing them, involve property loss and expense. Each fire generally presents a more or less original problem, dependent on the particular conditions. Generally mine fires may be divided into two classes: Those occurring in inaccessible workings and those occurring in accessible workings. The methods pursued in extinguishing or controlling mine fires are as follows: (1) direct attack, (2) smothering by sealing, (3) flushing with silt or other solids, (4) flooding with water, (5) cutting off, (6) surrounding with incombustible materials, (7) digging out by stripping from the surface. Most fires are extinguished in their incipiency by direct attack and failure to extinguish by this method is due to late discovery, lack of water, and facilities for fighting, or improper procedure in the initial attack. Should the method first pursued be unsuccessful, careful consideration should be given in determining next the method to be adopted; then the work should be carried on as rapidly as possible and under the supervision f the most competent persons. Materials and supplies should be fur-

Jan 9, 1921