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By D. S. Tomalin, D. F. Stein

The effect of reducing oxygen to low concentrations on the fracture of high-purity iron single crystals has been examined at 78° and 20°K. It is found that iron single crystals grown by the strain-anneal method usually contain a few occluded grain boundaries which may become embittled in the presence of oxygen, thereby nucleating cleavage fractures. High purity with respect to interstitial elements was found to inhibit twinning and evidence is presented for an orientation dependence of the resolved yield stress. Deformation occurred by slip rather than twinning at both temperatures of testing with elongations of as much as 9 pct at 20°K. THE fracture properties of iron single crystals have been observedl-3 to be a function of temperature, orientation, and purity. Allen, Hopkins, and McLennan1 demonstrated that at 78°K iron single crystals became increasingly brittle as the tensile axis approached the (001) pole of the stereographic unit triangle. Iron crystals with the tension axis near a (001) pole were completely brittle and orientations near the (011)-(111) boundary were very ductile, achieving a 100 pct reduction in area prior to fracture. Later work of Biggs and pratt2 and of Edmondson3 demonstrated that by reducing the carbon content of the single crystals the transition between brittle and ductile failure at 78°K could be shifted to orientations nearer the (001) pole. Ed-mondson went further and pointed out that any mechanism which tended to increase the yield strength of the iron (i.e., carbon addition, pre-strain) also increased the tendency for brittle behavior at 78°K. Thus by a reduction of carbon content Biggs and Pratt were able to obtain ductile behavior to within 20 deg of the (001) pole, and Ed-mondson was able to obtain ductile behavior to within 26 deg of the (001) pole. Edmondson's material had a total interstitial content (carbon, oxygen, and nitrogen) of approximately 60 ppm and, although Biggs and Pratt reported no analysis, indications were that their iron contained comparable impurities. Stein, Low, and seybolt4 purified iron single crystals to a total carbon, oxygen, and nitrogen content of approximately 20 ppm. They observed a lowering of the yield stress with this increased purity, and thus one might have expected an observation of increased ductility. Although they tested specimens of orientations which the previous workers had indicated should be ductile, the crystals failed at a 78°K in a brittle manner with little elongation. Stein noted,' however, that about 90 pet of the failures could be traced in origin to occluded grains. Allen et al.1 and Edmondson3 do not report examination of their cleavage surfaces, but Biggs and pratt2 reported that microscopic examination of the cleavage surfaces of many of their specimens revealed the presence of small occluded grains. Honda and cohen6 and Keh7 have also observed the initiation of fracture at occluded grains in iron single crystals. Therefore, the additional complication of occluded grains must be considered in studying the properties of iron single crystals and the origin of fractures determined if the ductility exhibited by the crystals is to be considered meaningful. Various investigators8-12 have studied the effects of impurities on grain boundaries in high-purity polycrystalline iron. Rees and Hopkins10 showed that the addition of oxygen to low-carbon (0.002 pet) iron weakened the grain boundaries, causing a shift from transcrystalline to intercrystalline fracture and a progressive decrease on the brittle-fracture stress with increasing oxygen content. In addition, it has been shown by Low and Feustel11 that the addition of carbon to polycrystalline iron containing oxygen eliminated grain boundary brittleness. Thus, oxygen can embrittle grain boundaries in high-purity polycrystalline iron, but the addition of an appropriate amount of carbon can eliminate the oxygen-induced brittleness. The oxygen content (19 ppm) of the iron "single crystals" employed by Stein el al. was large enough to suspect impurity effects at the occluded grain boundaries. Allen et al.1 Biggs and pratt,2 and Edmondson3 may have masked any occluded grain problem in their specimens considering the relatively high carbon levels they employed. stein13 demonstrated that the addition of as little as 0.9 ppm C to previously purified (less than 5 x 10-3 ppm C) "single crystals" would increase the elongation from a few percent to more than 20 pet at 78°K and the fracture was not associated with grade boundary initiation.

Jan 1, 1965

A filled stope is one in which the support for walls and men and, at times, for the back of ore, is furnished by waste rock or sand tailings. The filling may be rock sorted out in the stope or from the walls, develop-ent, or the surface. The orebody is mined in sections, each practically an open or timbered stope, filled wholly or in part before adjacent ground is attacked. As a rule, the filled-stope method is limited to overhand flat-back, stepped-face, or rill stopes. The examples given are Copper Range, Iron Cap, Verde district, Zaruma district, Lucky Tiger, and Silver King Coalition.

Jan 1, 1925

By Nigel B. Pearson

INTRODUCTION Chingola Division of Nchanga Consolidated Copper Mines Limited is situated in the Copper- belt of Zambia, approximately 30 kilometres south of the border with Zaire. It is the country's leading producer of copper and its principal foreign exchange earner, and has been, in its time, the largest single copper producer in the British Commonwealth and the second largest in the world. There are two 'primary' copper sources at the Division, namely Underground producing approximately 100 000 tonnes of contained copper annually, and the Open Pits, producing approximately 200 000 tonnes annually. Following the commissioning of the Tailings Leach Plant in 1974, treating final tailings, metallurgical recovery increased from approximately 70% to in excess of 80%. Extensive deposits of cobalt are present and following the rise in the market price in 1979 and 1980, plans were made to extract this separately. Cobalt production commenced in 1980 from Underground and will increase substantially when the Open Pit cobalt reserves become available in approximately 2 years time. This paper covers the development of the production method of the principal Underground Orebody, known as the Lower Orebody, and the control systems applied to it. HISTORY The Lower Orebody, formerly known as the Nchanga West Orebody, was discovered by extensive Diamond Drilling operations in the late 1920's. Development of the Shaft complex and the orebody commenced and continued until 1931 when flooding from development on the 480 foot horizon necessitated closure. The mine remained closed until 1937 and production commenced from the Lower Orebody in 1939. Various expansion schemes of the Metallurgical Plants and the Underground workings were completed over the years with the result that Underground production increased steadily to in excess of 150 000 tonnes copper annually in the 1960's. In recent years production has levelled out at slightly less than 100 000 tonnes annually. This decrease is as a result of producing from a lower grade reserve, available within a more limited production front, and the application of production constraints in the light of experience gained. GEOLOGY Within the Nchanga Mining Licence Area of Chingola Division, there are a total of 14 known copper orebodies in the basal part of a thick succession of sediments of Pre-Cambrian age. The sediments rest on a Basement Complex of granites, gneisses and schists. Three superimposed stratiform orebodies, known locally as the Lower, Intermediate and Upper Orebodies, occur over a vertical interval of approximately 150m. Current annual copper production is approximately 300 000 tonnes. The ore originates from four different mining operations, one Underground mine and three Open Pits. Approximately 9 million tonnes of ore grading between 3 and 3.5% copper are milled annually. The two largest sources of ore, the Under- ground mine and the Nchanga Open Pit are on the gently dipping south limb of the Nchanga Syncline (Figure 1). In the Nchanga Open Pit the major source of ore is the Upper Orebody with lesser tonnages from the Intermediate and Lower Orebodies. In the Underground mine the Lower Orebody is the only orebody being mined, underground mining of the Upper Orebody being suspended for economic reasons in 1978. The typical stratigraphic column is illust- rated in the legend on Figure 1. The footwall rocks comprise either Nchanga Red Granite or the Arkose formation, both of which are hard competent rocks, offering generally good mining conditions for footwall development. Copper mineralisation of the Lower Orebody extends from the upper part of the Arkose to the top of the Lower Banded Shales. Overlying the Arkose is the Transition zone (0-18111 in thickness). The Transition zone comprise a lower member, Transitional Arkose, consisting of incoherent kaolinised sandstone, and an upper member, Transitional Lower Banded Shales consisting of normally soft bleached and silicified shale. The soft nature of the Transitional rocks is

Jan 1, 1981

E. T. Dumble, Houston, Texas (communication to the Sec-retary?):—In addition to the list of publications mentioned by Mr. Schorr and those by Prof. Hofman, I call attention to the following references:— Sludi sulle Lignite, by Capacci. Turin (1890). (An excellent work.) Report of Brown Coal Industry in Germany. J. Cosmo New-berry. Department of Mines, Victoria (1892).

Jan 1, 1905

By H. M. Morse

This report on oil and gas development in Mississippi covers the period from Jan. I, 1944, to Dec. 31, 1944: Salt Domes During the year six salt domes were drilled for sulphur, but no commercial sulphur deposits were found. One salt dome, the Bruinsburg, in Claiborne County, has produced gas- at 935 to 940 ft' in Moody's Branch marl. The Stringer salt dome in Smith County and the Ruth salt dome in Lincoln County both had a slight showing of heavy oil. Other salt domes have been dry. In addition to the Bruinsburg dome, the Allen salt dome, in Copiah County, and the Richton salt dome, in Perry County, were discovered in 1944; also the Sumrall dome, in Covington County. New Oil Fields Baxtemille Oil Field.—The Baxterville oil field was discoveredby the Gulf Refining Co, in sec. 7, T. I N., R. 16 W., Lamar County. At the end of the year this field had one oil well, with two active operations. Production for 1944 was 1047 bbl. of oil. GwinvilLe Field.—Gwinville field was discovered by S. W. Richardson, in sec. 24, T. 9 N., R. 19 W., Jefferson Davis County. This discovery well was a gas-distillate well, the oil being crystal clear. At the end of 1944 there were two gas wells and one oil well in the field. The oil well is approximately 2 1/2-miles from the gas- distillate wells. Gas production was 55,801 M cu. ft. and the distillate production was 2644 barrels. Heidelberg Oil Field.— The Gulf Refining Co. discovered the Heidelberg field in sec. 30, T. I N., R. 13 E., Jasper County, in January 1944. This field has by far exceeded the other 1943 and 1944 discoveries, having approximately 56 oil wells, 7 dry holes, and 10 active wells at the end of the year. Production to Jan. I, 1945 was 1,441,192 bbl. of oil. Mallalieu Oil Field.—The california Company discovered the Mallalieu oil field in set. 10, T. 7 N., R. 8 E,, Licoln. County. At the end of the year there was one well in the field with three active operations. Production was 32,303 bbl. of oil. Old Oil Fields Brookhaven Oil Field, in Lincoln County, has one oil well, which produced 2590 bbl. of oil during 1944. CarY Oil Field, Sharkey County1 has only one producing well, which produced 11,818 bbl- of oil during the Year- In Cranfield Oil Field, dams County, II oil wells and two gas-distillate wells were drilled in 1944 and there were five active wells on Dee. 31, 1944- The total production for the Year was 5251422 bbl-of oil. The gravity of the oil is 40" to 54.9'- In Eucutta Oil Field1 Wayne County1 34 oil wells were completed during 1944. There were five dry holes and six active wells at the end of the year.. The field produced 454,332 bbl. of all d the average gravity is 25" Flora Oil Field, Madison County, pro-

Jan 1, 1945

By S. H. Williston

IT not often that bore hole surveys can be checked by actual. civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bis-bee, Ariz., drilled a 1245 ft, 8 in. diarn, churn drillhole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location. The noise of the drilling tools was reasonably clear, but the direction of sound was uncertain. Preliminary tests with available equipment were not successful in locating the bottom of the hole. Because of the mineralized character of the area and the fact there was some casing in the hole, any magnetic method of well surveying would give results of doubtful value. Sperry-Sun gyroscopic well-surveying instruments were finally used to locate the bottom of the hole. These instruments consist of a gyroscope to determine azimuth and either a pendulum or bubble inclino-meter. A multiple shot camera photographs both instruments on a single film and superimposes the photograph of a watch. Coordination of depth with time at the surface makes it possible to select the corresponding picture for any depth. After making several runs of the empty instrument housing from the top of the hole to the bottom to make sure there were no obstructions in the hole, three surveys on wire line were completed during the afternoon. The three surveys, in which readings were taken at different points in the hole on each survey, were computed and gave the following locations of the bottom of the hole in relation to the surface collar: survey No. 1—24.92 N, 30.30 W; survey No. 2—24.24 N, 31.11 W; survey No. 3—26.54 N, 27.72 W. Then the data from the three surveys were combined into a single set of calculations which gave a location for the bottom of the hole: combined surveys—24.27 N, 30.16 W. (Fig. 1.) Immediately upon the determination of the coordinates at the bottom of the hole, a drift on the 1300 ft level was started toward the indicated loca- tion some 38 ft northwest of the coordinates of the surface location, The bottom of the hole was located within the drift round in which it was expected, and the transit survey run to the actual location of the hole indicated N 27.18, W 29.71. This shows a discrepancy between the well survey and the transit survey of 0.45 ft in the westerly direction and 2.91 ft in the northerly direction. All surveys, both gyroscopic and transit, fell well within the width of an ordinary drift. While this is satisfactory for almost any and all mining requirements, a theoretical examination was made as to reasons for the discrepancy. A study of the course of the hole indicates that considerable right turn or spiral existed, and in all probability the surveying instrument was pulled out of alignment while traversing the turn by approximately 0.05 ft at the top and another 0.05 ft in the opposite direction at the bottom of the instrument. If such an allowance were to be made, the survey calculations would almost exactly correspond with those determined by transit. This sort of discrepancy would be minimized by the use of stabilizing guides. It is unfortunate that physical laws probably effectively prevent the use of gyroscopic instruments in EX and AX diamond drill holes. The directive power of a gyroscope falls off inversely at some rate between the third power and the sixth power of the diameter. Present instruments can be run in casing 53/4 in. ID or over and might be adapted to somewhat smaller diameters, but the difficulty of reducing these diameters to 11/4 in. or 2 in. is almost insurmountable at the present time. Acknowledgment The author wishes to express his appreciation to the Phelps-Dodge Corporation for permission to publish this article, and to the Operating and Engineering Departments for their cooperation on the survey; also to Donald Hering, of the Sperry-Sun Well Surveying Co., who actually made the survey and calculated the results.

Jan 1, 1951

By S. H. Williston

IT not often that bore hole surveys can be checked by actual. civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bis-bee, Ariz., drilled a 1245 ft, 8 in. diarn, churn drillhole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location. The noise of the drilling tools was reasonably clear, but the direction of sound was uncertain. Preliminary tests with available equipment were not successful in locating the bottom of the hole. Because of the mineralized character of the area and the fact there was some casing in the hole, any magnetic method of well surveying would give results of doubtful value. Sperry-Sun gyroscopic well-surveying instruments were finally used to locate the bottom of the hole. These instruments consist of a gyroscope to determine azimuth and either a pendulum or bubble inclino-meter. A multiple shot camera photographs both instruments on a single film and superimposes the photograph of a watch. Coordination of depth with time at the surface makes it possible to select the corresponding picture for any depth. After making several runs of the empty instrument housing from the top of the hole to the bottom to make sure there were no obstructions in the hole, three surveys on wire line were completed during the afternoon. The three surveys, in which readings were taken at different points in the hole on each survey, were computed and gave the following locations of the bottom of the hole in relation to the surface collar: survey No. 1—24.92 N, 30.30 W; survey No. 2—24.24 N, 31.11 W; survey No. 3—26.54 N, 27.72 W. Then the data from the three surveys were combined into a single set of calculations which gave a location for the bottom of the hole: combined surveys—24.27 N, 30.16 W. (Fig. 1.) Immediately upon the determination of the coordinates at the bottom of the hole, a drift on the 1300 ft level was started toward the indicated loca- tion some 38 ft northwest of the coordinates of the surface location, The bottom of the hole was located within the drift round in which it was expected, and the transit survey run to the actual location of the hole indicated N 27.18, W 29.71. This shows a discrepancy between the well survey and the transit survey of 0.45 ft in the westerly direction and 2.91 ft in the northerly direction. All surveys, both gyroscopic and transit, fell well within the width of an ordinary drift. While this is satisfactory for almost any and all mining requirements, a theoretical examination was made as to reasons for the discrepancy. A study of the course of the hole indicates that considerable right turn or spiral existed, and in all probability the surveying instrument was pulled out of alignment while traversing the turn by approximately 0.05 ft at the top and another 0.05 ft in the opposite direction at the bottom of the instrument. If such an allowance were to be made, the survey calculations would almost exactly correspond with those determined by transit. This sort of discrepancy would be minimized by the use of stabilizing guides. It is unfortunate that physical laws probably effectively prevent the use of gyroscopic instruments in EX and AX diamond drill holes. The directive power of a gyroscope falls off inversely at some rate between the third power and the sixth power of the diameter. Present instruments can be run in casing 53/4 in. ID or over and might be adapted to somewhat smaller diameters, but the difficulty of reducing these diameters to 11/4 in. or 2 in. is almost insurmountable at the present time. Acknowledgment The author wishes to express his appreciation to the Phelps-Dodge Corporation for permission to publish this article, and to the Operating and Engineering Departments for their cooperation on the survey; also to Donald Hering, of the Sperry-Sun Well Surveying Co., who actually made the survey and calculated the results.

Jan 1, 1951

By H. Bloom, H. E. Hawkes

STREAMS and rivers are the principal channels into which the weathering products of rocks and their contained ores are funneled. The inorganic load of a stream system is a crude sample of all the earth materials in the drainage basin of the stream, and the heavy mineral assemblage of sediments is one of the oldest and most successful guides to certain kinds of bedrock ore in unexplored terrain. Early application of this method was made possible by the ease of separation and identification of heavy minerals in the field with the simplest equipment. Water analysis for traces of base metals became a practical possibility about eight years ago with the development of rapid chemical tests that could be carried out at the field site.' , Since that time the water method has been widely used in mineral reconnaissance, and a substantial number of discoveries have been reported. Here again, the techniques developed for this work were relatively simple and could be performed easily at the field site. Sediment analysis by chemical methods, as distinguished from mineralogical, has received less at- tention. Some experimental work was done by Lovering and others," but to the authors' knowledge, chemical analysis of sediment samples has not been used in routine exploration programs until very recently. In spite of the obvious promise of this method, it was not pursued vigorously because of the lack of a simple and portable chemical test that could be used conveniently at the field site or in a field camp. Procedures that had been developed for trace analysis of soils all called for a rigorous chemical attack with hot acids—tests more conveniently applied in processing large volumes of routine soil samples at an established field headquarters. Recently an extremely rapid and simple test was developed by Bloom.4,5 The extractant used in this procedure is a cold solution of ammonium citrate. Elements measured are the heavy metals, principally zinc, lead, and copper. This method differs from the established procedures for soil analysis in that it measures only a small fraction, about 5 pct, of the total metal content of the sample. It has been found, however, that the significant pattern of metals in the sediments of streams below a base-metal deposit can be brought out as well and sometimes better by this simple cold-extraction procedure than by the more tedious methods requiring hot acids. To understand fully the significance of the relation between the readily extractable and the total

Jan 1, 1957

By Samuel G. Vera

INTRODUCTION The Climax Mine is a property of Climax Molybdenum Company, a division of AMAX Inc. Climax is located in the central part of the Colorado Mineral Belt on the west slope of the Tenmile Range, Lake County, Colorado at 3,425' m. above sea level. Climax Mine, the largest single producer of molybdenum in the free world, has been a cornerstone of mining in the State of Colorado for many years. Before molybdenum was known, the early hunters of gold and silver laid their hands on the outcrops of the Climax deposit and left it as low quality lead or graphite without much value. Today, 44,000 TPD of ore is extracted from the heart of Bartlett Mountain. The molybdenum, separated from the ore, is vital for mankind, being used for everything from fertilizer to fine components of aerospace ships, because of its valuable physical and chemical properties. The operations at Climax have never remained static. Constant demand for the "gray gold", and technological advancements, have converted Climax into an "open book of technology", where learning the still-hidden secrets of mining is a continuing process. (All figures related to tons are metric tons) BRIEF HISTORY The mine produced its first molybdenum ore in 1918. It was shut down in 1919, resumed production in 1924, and has been in continuous production since that time. During World War 11, a peak of 20,000 tons per day was reached, then tonnage dropped from 1945 to 1950. Since 1951, production has been in- creasing, reaching a peak of 46,000 tons per day during the month of June, 1980. Early mining used shrinkage stoping. Block caving using the grizzly system was introduced in the early 1930's. Caving methods were improved through empirical studies of rock behavior and by trial and error, evolving a very efficient mining method by the late 1940's. This form of continuous retreat panel caving, with small modifications, is still used today, and has made the Climax Mine one of the biggest underground mines in the world. The Climax Mine, synonymous with molybdenum, was essentially an underground mine until 1973 when the open pit was introduced. Since then, both mining methods have operated successfully. Current production is 60% from underground, and the remaining from the open pit. Climax has been in continuous operation for 56 years, and as of December 31, 1980, more than 386 million tons of ore have been mined. The orebody is not fully defined, and at the present mining rate, the known life span of the mine will be to the year 2020. GEOLOGY The Climax deposit is a large low-grade porphyry molybdenum deposit somewhat similar to porphyry copper deposits, with a well- mineralized quartz molybdenum stock work sys- tem. It lies within the pre-Cambrian granites and schists and the Tertiary Climax porphyry. The age of mineralization is thought to be mid-Tertiary and is characterized by fracture filling and replacement through silicification. The main economic mineral is molybdenum, with pyrite, cassiterite and wolframite as by-products. Fluorite and topaz also occur, but are not recovered at this

Jan 1, 1981

By H. S. Wang, Y. P. Gupta

Diffusion of zinc and germanium in liquid silver is measured in the temperature range 975" to 1400°C by use of the capillary-reservoir technique. The ex-perinzental results are discussed in terms of the fluctuation and the "hole" theories of liquid diffusion, assunzing the solute as a point charge equal to the difference in the valences of the solute and solvent atoms. The additional coulombic interaction term due to the excess solute charge is evaluated in terms of the Thomas-Fenni approximation. The theoretical and experimental results agree with both the fluctuation and "hole" theories of liquid diffusion. No perceptible effect on solute diffusion in liquid silver of the size difference between the solute and solvent atoms/ions is apparent. DIFFUSION in liquid silver of electropositive solutes indium, tin, antimony, and homovalent gold has been previously reported by Leak and Swalin1 and Gupta,' respectively. These solutes have nearly the same atomic size but differ principally from one another in excess valence (+2, +3, +4, respectively). The observed differences in the temperature dependence and the rate of solute diffusion were explained in terms of an added coulombic interaction between the solute and the solvent atoms. In the case of diffusion of gold in liquid silver, the data were found to be consistent with an effective valence for gold of -0.85, as proposed by Le Claire.3 Recently, Wang and Gupta4 have reported a study of diffusion of ruthenium, iron, and cobalt in liquid silver. The results of diffusion of these solutes and their temperature dependence were found consistent with the hole theories and the fluctuation theories of liquid diffusion, if ruthenium, iron, and cobalt were diffusing in liquid silver with relative valences -2, -1.3, and -0.5, respectively, instead of -3, -3, and -2 as suggested by the relative position of these elements in the Periodic Table of Elements. Furthermore, the predicted valences for these solutes were observed to be consistent with the sparsely available information on the magnetic properties of some of the dilute alloys of copper and silver with these solutes. In theoretical calculations of solute diffusivity, the solute was assumed as a point excess charge equal to the difference in the valences of the solute and solvent atoms. Both the assumption of a point excess solute charge and the use of the Thomas-Fermi approximation to calculate the potential of a solute charge are less precise for the case of electronegative solutes. Therefore, it was not possible to infer from the data on ruthenium, iron, and cobalt an effect, if any, on solute diffusion of the relative differences in either the Period or the size of the solute and solvent atoms or the solute atoms themselves. A study of the diffusion of Period 4 elements, specifically zinc and germanium was, therefore, undertaken since zinc and germanium, respectively, have relative valences of +1 and +3 in liquid silver and the data obtained on diffusion of these solutes could be compared with the available data on diffusion of Period 5 electropositive solutes, indium, tin, and antimony in liquid silver. From such a comparison, it was hoped to learn something about the effect, if any, of the ion size or the structure of the ion core on impurity diffusion in liquid silver. EXPERIMENT The diffusion of tracer zinc and germanium in liquid silver was studied by the capillary reservoir technique, modified adequately for use at high temperatures. In principle, capillaries containing liquid silver

Jan 1, 1970

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By C. H. Mathewson

MICROSCOPIC. rnetallography has been exploited quite well enough to bring about a very general understanding that the typical metal or alloy is composed of minute crystalline particles blended into a coherent microstructural mosaic. One does not have to be a specialist in metallography to realize that the properties of such an aggregate are essentially a summation in appropriate form of individual effects derived from the shape, size, placement or orientation and cohesive characteristics of the component particles. It is clearly of great importance to consider the various forms of discontinuity which may occur at the boundaries between these crystalline particles. When a number of 'particles each possessing the same orderly arrangement of atoms are brought together into a close-fitting system of purely haphazard contacts, not unlike a handful of snowflakes compacted into a snowball, the particles are said to possess random orientation. There are no generally accepted views concerning the arrangement of the atoms or the constitutional reaction between atoms where one crystal meets another, but these contact regions are characterized by strength rather than weakness and it is customary to require the presence of many rather than few boundaries in preparing metal for useful service. Under certain circumstances all of the particles in a metal may be nearly alike in orientation. Other general tendencies of orientation may be associated with particular forms of mechanical and thermal treatment. In contrast with these cases of fortuitous or statistical diversity of orientation, we often find adjacent particles united along a plane which possesses a grouping of atoms belonging equally well to both structures. This, of course, determines a fixed relationship between the two orientations and it is always possible to derive one from the other by some form of rotation or reflection prescribed by the symmetry of the crystal structure under consideration. Particles united in this manner are known as twin crystals although the term refers to the form of association rather than the number of individuals concerned.

Jan 1, 1928

By William P. Shinn

It is almost exactly seven years since the last previous meeting of the Institute in this city. In a paper on " Pittsburgh, its Resources and Sorroundings," read at that meeting, I showed that Allegheny County had produced in 1878, " one-twelfth of all the pig-iron,.... two-ninths of all the rolled iron (except rails), over one-eighth of all the Bessemer and open-hearth steel, two-thirds of all the crucible blister and puddled steel, one-fourth of all the glass, one-eighth of all the coal, and one-sixth of all the pig-lead produced in the United States;" and the object of this paper is to show that, great as the record then was, it has been greatly surpassed since.* Population. In my former paper I expressed the opinion that the census of 1880 would show the population of Allegheny County to be in excess of 300,000. The census report made it 355,869. The following shows the growth of the county in population from 1790 to 1880: 1790,...........10,309 1800,........... 15,087 1810,...........25,317 1820,...........34,921 1830,...........50,552 1840.............81,235 1850............138,290 1860............178,881 1870,...........262,204 1880,...........355,869 This makes Allegheny County, in 1880, rank sixth in population of the counties of the United States, whereas in 1870 its rank was ninth. The population of Pittsburgh in 1880 was 156,359, being twelfth

Jan 1, 1886

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By G. T. Jackson

The Alaska Gastineau Mining Co.'s mine is located at Perseverance, about 4 mi. east of Juheau, Alaska. Its property consists of a group of claims, the lode system traversing these claims for a distance of 11,000 ft. (3352 m.). It is shown in Fig. 1. The ore deposits occur in what is known as the Juneau gold belt. This gold belt can be traced for about 50 mi. northwest of Juneau and 40 mi. southeast. It consists of a single band, several hundred feet wide, in which stringers and veins of quartz, carrying gold, occur in a slate formation near its contact with some altered volcanic rock known as greenstone. The Perseverance orebody consists of stringer lodes, having a strike northwest and southeast, dipping about 60° to the northeast. These

Jan 1, 1920

By H. H. Stoek

Danville, the county seat of Montour county, Pa., was one of the earliest and best-known centers of the iron intlustry in the State. It is situated on the north bank of the north branch of the Sosquehanna river, twelve miles east of the confluence of the north and west branches of that river at Northurnberlttnd. The ore-belt, extending for a distance of twelve miles along the north bank, has been developed in a number of places, but nowhere so extensively as in the immediate vicinity of Danville. Hence, a description of the mines in that locality will suffice, as typical of the workings of the entire section. Geology. The Montour ridge consists of an anticlinal fold extending from the source of Wapwollopen creek on the east to the west branch of the Susquehanna, which it reaches at the mouth of Chilisquaque creek. According to the reports of the First Geological Survey of Pennsylvania, the length of the ridge is nearly twenty-seven miles, with an average breadth of three-fourths of a mile, and a maximum elevation of six hundred feet in the vicinity of Danville, where it also attains its greatest width. Just north of Danville the Mahoning creek cuts through it, exposing the iron-ore on the hill-side, where advantage has been taken of this to work the beds on the southern slope of the ridge by means of drifts. The sections given in Fig. 1 are taken from the reports of the First Geological Survey of Pennsylvania, made in 1858, by H. D Rogers, whose description of the Moatour ridge was so complete and accurate that the entire report was reprinted as an appendix to Vol. G 7 of the Second Geological Survey. The approximate crosssection shows that near Danville the top of the ridge is of Clinton vol, xx.—24

Jan 1, 1892

By C. R. Fast, George C. Howard

Laboratory and field testing of various squeeze cementing techniques and materials revealed that many improvements could be made in squeeze cementing operations. The use of a slow-pumping squeeze cementing procedure permitted the control of the quantity of cement displaced into a formation and aided in obtaining a high final squeeze pressure. Field testing indicated the desirability of obtaining high final squeeze pressures, the need for improved formation breakdown fluids, and the necessity of controlling the pressure differential during testing after a squeeze job. INTRODUCTION During the past few years, squeeze cementing operations have been investigated, both from the theoretical as well as the practical point of view, in an effort to arrive at some solution to the many and varied problems that arise with this type of operation. A review of well files indicated that in many instances, wells were being squeeze cemented four and five times before a satisfactory pressure buildup was obtained. There seemed to be considerable variation in opinion as to the proper method of squeeze cementing, as well as disagreement as to the proper breakdown agent, the quantity of cement that should be used, type retainer and length of WOC* time that was necessary. Consequently, it was deemed advisable to approach this problem theoretically, field testing each idea developed in hopes of eventually arriving at a squeeze cementing procedure that would give reasonable assurance of a successful job after one squeeze cementing operation. This investigation was. therefore, divided into two broad headings: (1) Preliminary studies, which include both theoretical and laboratory studies, and (2) Field investigations to determine the proper squeeze cementing methods. PRELIMINARY STUDIES Overburden Pressure Studies Squeeze cementing problems are so intimately related to the earth's overburden pressure that solutions to them are very difficult, if not impossible, without an understanding of the nature and magnitude of the overburden pressures encountered at the depths of producing wells. It was thus felt that the first approach to research on squeeze cementing should be made through a study of these pressures. Since there has been rather loose usage of a number of the terms related to overburden preasure by different writers, the terms to be subsequently used are defined here. Theoretical overburden pressure is the calculated pressure at any given depth exerted by the theoretical weight of overlying formations as calculated from the average rock density. Fluid displacement pressure is that pressure at which fluid can be injected into a formation after it has been broken down. It may be more or less than the theoretical overburden pressure. Formation breakdown pressure is that pressure at which a formation parts or breaks and is usually thought of as fluid displacement pressure plus rock strength. Breakdown agent is that fluid with which the formation is fractured. The first step in the study of the earth's overburden pressures involved a review of pertinent technical literature. The general opinion expressed was that the overburden pressure gradient of the earth's formations is roughly 1.0 psi per foot of depth, and that application of sufficient pressure to stratified formations. will cause them to part at or parallel to the bedding planes. The second step in the study of overburden pressures involved a study of available well data in an effort to obtain some definite information as to the actual values experienced in the field for the above mentioned terms. Data from 115 cement squeeze jobs in the Gulf Coast area and 46 cement squeeze jobs and acid jobs in the West Texas-New Mexico area were closely analyzed. The analysis of these data showed that the average formation breakdown

Jan 1, 1950

By Samuel Shallcross

THE principal function of the Bell limestone mine, of the American Lime & Stone Co., at Bellefonte, Pa., is to supply raw material to the company's modern rotary kiln plant at Bellefonte. Because mining costs are naturally higher than many open quarries with more favorably lying strata, it cannot compete, except to a limited extent, in the markets for road material and fluxing stone. In the broad sense, mining does have an advantage over the average open quarry when the end products are designed for and controlled by process and final inspection for the exacting demands of the industrial and chemical markets. Approximately 80 per cent of the output of the Bell mine is sold in the industrial and chemical markets. And as higher standards are set each year for construction and agricultural work, the remainder of the mine output goes to these two classes. Mining produces more fines than open quarries, and these fines are difficult to market; on the other hand, owing to absence of contaminating clay and impure rock, a greater percentage of the output of a mine is marketable. Because of the great production of fines, rotary kilns for lime burning are necessary when the stone must be mined. In 1921 the American Lime & Stone Co. operated 19 quarries and 72 lime shaft kilns spread over an area of 40 miles between Bellefonte and Altoona, Pa. At that time the line-production capacity of the 72 kilns amounted to approximately 700 tons per day. After 15 years the company operates three rotary kilns and six shaft kilns, with a combined output of approximately 500 tons of lime per day, concentrated in its single Bellefonte plant (Fig. 1). Its only stone-producing plants are now the limestone mine at Bellefonte and a construction-stone open quarry at Union Furnace, near Tyrone, Pennsylvania.

Jan 1, 1936

By George B. Langford

ON THE QUESTION of postwar controls there are today three schools of though ; some advocate state control of everything the socialists ; second are those who advocate the removal of all governmental control over industry - the freedom of enterprise ideal ; and third are those who advocate a middle course - sufficient control to stabilize economic and industrial life, and to prevent violent fluctuations such as we saw in the interwar years. I advocate this third course. I firmly believe un free enterprise, but not in freedom of enterprise, or laissez faire, and there is a great and significant difference between these two.

Jan 1, 1944