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By L. A. Willey

IN 1953, a method for filtering a liquid phase from solid phases while in equilibrium at elevated temperatures was devised at the Alcoa Aluminum Research Laboratories. Since then, it has been applied to the investigation of liquid-solid equilibrium in metallic systems. In application, the method simplifies the centrifuge filt.ration procedure as developed by Scheil and described by Gebhardt Fig. 1 shows a sketch of the filtering apparatus located within a thermal analysis furnace. It consists of a stainless steel filter crucible A, into the bottom of which a porous carbon filter disk B, 1 3/4 in. diam by 1/4 in. thick, is sealed to the shoulder and rim with carbonaceous cement. The base of the filter crucible, machined to an angle, fits into a tapered seat at the mouth of a stainless steel vacuum chamber C. A connecting tube from the vacuum chamber to a vacuum line also serves as a means for supporting the apparatus from the base of the furnace. An alundum crucible D of 25 ml capacity is placed with- in the vacuum chamber below the filter for collecting the filtrate. Two thermocouples E, in stainless steel protection tubes, which extend from 1/16 to in. above the center of the filter, are mounted on the rim of the filter crucible. One thermocouple is used in differential connection with the furnace control thermocouple to regulate rates of heating and cooling. The other serves as a means for recording the temperature of the alloy prior to filtering. A suitable mold coating is used on exposed surfaces of stainless steel parts to guard against attack upon these parts as well as to avoid contamination of the alloy samples. A mold wash prepared from precipitated CaCO, is suitable for aluminum alloys. The porous carbon for the filter disks and the carbonaceous cement, grade 200 (graphite base), were obtained from the National Carbon Co. Porous car-bon, grade Nos. 10 to 60, has been employed. Grade No. 40, however, appeared to give the optimum advantages in operation. In operation, an alloy sample is remelted and heated to about 50°C above its liquidus temperature. At the same time, the filter apparatus is preheated to about the same temperature. The alloy then is cast into the filter crucible.

Jan 1, 1957

By M. A. Scheil, S. L. Hoyt

About three years ago eight standard steels were prepared for the cooperative investigation of methods for oxygen analysis, sponsored by the Iron and Steel Division of the American Institute of Mining and Metallurgical Engineers and by the National Bureau of Standards. A preliminary report of the results was made at the annual meeting of the A.I.M.E. in February, 1936. † This report showed the reliability of vacuum-fusion methods and further brought out certain advantages of the fractional vacuum-fusion method. The results obtained by the authors with the latter method will be discussed here, with some comments on the procedure for those who are interested in determinations of gases in steel. Some additional comments are offered on the use of the polarizing microscope for identifying the inclusions in the steel to check and amplify the findings of the fractional vacuum-fusion analysis. For details of the fractional method, see the paper by L. Reeve1. Degassing the Crucible One of the sources of error in vacuum-fusion methods has been the blank gases evolved from the graphite crucible after preliminary out-gassing. We have used 1750" C. as the outgassing temperature. This gives a base pressure at that temperature of less than 0.030 mm. after one hour and of 0.004 mm. on dropping the temperature to 1575" C. At lower extraction temperatures the base pressure is 0.001 mm. at 1050" C., 0.002 mm. at 1170" and 0.003 mm. at 1320' C. Because of the size of our induction coil we are limited to this temperature, although it 'would be advantageous to use 1950" to 2000" C. Hamilton2, after outgassing at 1850" C. for 100 min., was able to reach a base pressure of 0.001 mm. for a temperature of 1700' to 1725" C. This would give a blank corresponding to less than 0.00015 per cent 02 and 0.00005 per cent H2. Chipman and Marshall outgas at 2000" C. for one hour by using a special construction

Jan 1, 1937

By H. G. Doll

A new logging method, called induction logging, is described; it measures the conductivity, or resistivity, of the strata traversed by a bore hole. The apparatus, which is briefly described, comprises a coil system coupled with the ground by induction, so that there is no need for direct contact with the mud, or with the formations. For that reason, the method is particularly useful in oil base mud, and it was first applied to that case. A discussion concerning the respective contribution of each region of ground in the vicinity of the coil system is given. Through the concept of a geometrical factor, it is demonstrated that the total signal, which measures the apparent conductivity, is the summation of the partial products — geometrical factor by conductivity — for the different regions of the ground under consideration. It is thus possible to establish, for each type of coil arangement, investigation characteristic as well as departure curves and correction charts for beds of various thicknesses. This should greatly facilitate the quantative interpretation of the field logs, even in cases where such interpretation would have been difficult on electric logs taken with electrodes. It is also shown that focussing combinations of coils can be used, which give, directly on the recorded log, a value close to the true conductivity, even for fairly thin beds, thus eliminating in most cases the need for corrections. These focussing systems also give a sharper definition of the boundaries between successive strain.

Jan 1, 1949

By Kenneth Cottingham

The number of wells completed in Ohio during I940 was about 20 per cent more than during the preceding year, the comparative totals being 1020 and 1228. Of the 1940 completions, 327 were oil wells, 491 gas wells, and 410 dry. The completions for the past five years, divided into oil wells, gas wells, and dry holes, are shown in Table 2. Wells drilled deeper and input wells in water-flood areas are not included in the tabulation. The annual production of oil has decreased steadily in Ohio from the maximum of 23,941,000 bbl. produced in 1896. Table 3 shows the quantities of oil produced in the last five years. In 1940, the total was 3,052,000 bbl., produced by about 25,500 wells. The number of wells and the approximate quantities by grades are shown in Table I. Table i ,—Tabulation by Grade Grade 1 2 Barrels __________________________________________________________________________________________ Pennsylvania ............. . 1,198,000 Lima.................... . 325,000 Corning................... 3,730 1,159,000 Lodi...................... T,040 350,000 Cleveland................. 30 1:1,000 Total................... 25,500 3,052,000 The prices paid for Ohio crude oil in 1940, together with dates of changes in posting, are shown in Table 4. The production of natural gas in 0hio has declined from a maximum of 79,510,032 M cu. ft. in I9I5 S to 36,469,000 M cu. ft. in 1939, the most recent year for which a figure is available. Consumption reached its peak in Ohio in the year 1916, when it was 169,480,01r M cu. ft., declining to 114,720,ooo M CU. ft. in 1939. The produc-tion and consumption of natural gas in Ohio for the past five years are shown in Table 5. TAble 2.—Completions in Ohio during Past Five Years Numbs of Wells Drilled during Year YearI, ] Oil Gas Dry Total 1936............. 263 570 324 1,157 1937............. 317 503 45I 1,271 1938............. 189 433 288 gro 1939............. I70 501 349 1,020 19391940............. 327 491 410 1,228 Total.......... 1,266 2,498 1,822 5,586 Table 6 shows the number of oil wells and the number of gas wells completed by sands during 1940, with the average initial production per well for each sand. Table 7 shows the total number of wells drilled—oil, gas and dry—by sands. Development during 1940 Szth-Trenton.—For several years, considerahle interest has been shown in a horizon lying some 500 to 700 ft. below the top of the Trenton, known to the driller in western Ohio as the "green sand," and in eastern Ohio as the St. Peter. In 1939, ten 'lgreen sand" tests were drilled, all in western Ohio and all dry- In 19401 eight wells were drilled to this horizon, one of which was a small oil well (one barrel) in Men county- Of the remaining seven—all of which were dry-—one in Franklin County, in central Ohio, was drilled to a depth of 2903 ft. .Another was in Medina

Jan 1, 1941

By J. D. Wright

INTRODUCTION The Endako deposit was first explored in 1926, but development waited until the early nineteen sixties. It first came under the control of Placer Development in 1963, and early in 1964 a decision to bring it into production was reached. The deposit is located near the geographic center of British Columbia, and is served by an east-west communication system including a paved highway, mainline railway and good telephone, teletype, and computer facilities. The plant area is shown in Figure 1. The operation is known as the Endako Mines Division of Canex Placer Limited of Vancouver. First production was achieved in June, 1965, at a nominal rate of 10,000 tons per day, but it soon was closer to 15,000 tons per day. In 1967, two new grinding and flotation, sections brought capacity to the present rate of approximately 27,000 tons per day. With mill heads of around 0.15 percent MoS2, production is about 14,000,000 pounds of molybdenum per year. A flowsheet of the operation is given in Figure 2. A legend for this flowsheet is given in Table 1. Published ore reserves are estimated at 214,500,000 tons at an average grade of 0.143 percent MoS2. PRIMARY CRUSHING The 85 or 100-ton haulage trucks dump over a grizzly into a 42 inch x 65 inch C.A.C. (Canadian Allis Chalmers) gyratory crusher. The

Jan 1, 1976

By C. R. Wohrman

A melt of pure electrolytic iron with about 0.4 per cent. sulfur and 7 per cent. manganese was prepared in connection with a study of inclusions in iron. The alloy darkened rapidly when etched with a 10 per cent, alcoholic solution of nitric acid, and microscopic examination revealed the presence of a semimartensitic, semi-Widmanstättian structure (Figs. 1 and 2) closely resembling the structures found by Benedicks in his artificially prepared "meteoric iron."' The structures found by Benedicks were readily explained by the presence in his alloy of the two well-known iron-nickel solid solutions, taenite and kamacite, and of their eutectoid (?) aggregate, plessite. By analogy, one would infer the existence of similar phases in the iron-manganese alloy just described, such an inference being strengthened by the similarity displayed in many ways by nickel and manganese in their effects on iron. Iron-manganese Alloys are Regarded as Homogeneous Heterogeneity of iron-manganese alloys has not been recognized heretofore, judging from the data of standard metallographic books and those of original papers. In fact, iron and manganese are quite generally regarded as completely miscible in the solid state. Dr. Rosenhain,2 for example, writes: Alloys of group (a), of which those of gold-silver, iron-manganese, and copper-nickel are examples, when allowed to crystallize sufficiently slowly to attain their condition of final equilibrium, present a microstructure exactly similar to that of a pure metal—the homogeneous character of the molten solution is in that case so completely maintained that the microscope cannot detect the presence of any second constituent. Guillet and Portevin3 refer to the iron-manganese system as an example of the "case of metals completely miscible in the solid state."

Albin, B. R Billings, Mont. '28 Amster, Nathan L 25 Broad St., New York, N. Y. '28 Bache, Jules S 42 Broadway, New York, N. Y. '28 Barrett, C. P 5338 Harper Ave., Chicago, 111. '28 Bayless, William S Russ Bldg., San Francisco, Cal. '28 Bean, Louis, H 60 E. 42nd St., New York, N. Y. '28 Beatty, Ross J Room 604, 6 N. Clark St., Chicago, III. '28 Benwell, Oswald F Hold Mags. '28 Beymer, Albert S., V.P., Keystone National Bank Pittsburgh, Pa. '28 Blumenthal, Hugo 44 Pine St., New York, N. Y. '28 Boettcher. Charles Ideal Cement Bldg., Denver, Colo. '28 Bowlus, Thomas H., Banker, Pres., Allen County State Bank Iola, Ran. '28 Brown, Colvin B Chamber of Commerce, Washington, D. C. '28 Burt, Maxwell Struthers Southern Pines, N. C. '28

Jan 1, 1934

By I. W. L. Finlay, H. R. Ogden, R. Jaffee, D. J. Maykuth

Aluminum has been found to be soluble in a titanium to about 26 pct, and to raise the temperature range of transformation from a to 8. Two intermediate phoses exist in the system, a new face-centered tetragonal phase, designated as 7, which occurs between 34 and 46 pct Al, and TiAI Metallographic and X-ray diffraction data were used to determine the diagram. ALUMINUM is one of the useful alloying additions for titanium-base alloys. It is an important constituent in two commercial alloys released in 1950, one containing 4 pct A1 and 4 pct Mn, and the other containing 3 pct A1 and 5 pct Cr as the metallic alloy content. It is probable that, in the future, other titanium alloys containing aluminum will be released. In the course of alloy development work, some of the features of the constitution of titanium-rich Ti-A1 alloys have been developed. These are described in the present paper. The only available information on the Ti-A1 system concerns the aluminum-rich end of the diagram. In summarizing the early work, Hansen' indicated that there is little solubility of titanium in aluminum, and that a two-phase region exists up to the intermediate compound, TiA1a (37.2 wt pct Ti), which has a congruent melting point at 1355°C. Fink' has reported that a peritectic reaction occurs at 665°C between the melt and TiA1³ to form the terminal solid solution of titanium in aluminum. The solubility of titanium in aluminum has been established at 0.24 pct at 510°C and is estimated to be between 1 and 1.5 pct at the peritectic temperature. No information has been published on the diagram above TiAl,, other than a statement- hat aluminum raises the a to ß transformation temperature of titanium. Materials and Preparation of Alloys Iodide-type titanium and high-purity aluminum were used in this investigation. Typical analyses of these materials are given in Table 1. Alloys were prepared by arc melting on a water-cooled copper hearth, using the techniques described previously for contamination-free melting.' Weight losses during melting were very low, ranging from 0 to 0.02 g in a 10 to 15 g charge, indicating that little or no loss of aluminum occurred during melting. This permitted the use of nominal compositions for these alloys. Alloy ingots containing from 0 to 7.5 pct A1 were rolled in air at 850°C to 0.040 in. sheet. Alloys containing from 7.5 to 16 pct A1 could only be rolled about 25 pct reduction at temperatures up to and including 1200°C. Alloys containing above 16 pct A1

Jan 1, 1952

By Herluf P. Nielsen, Robert I. Jaffee

IN recent years much attention has been given to high-melting alloy systems, It has been of interest in this respect to investigate the alloys of platinum and tungsten, which were known from the literature to have very high melting points up to about 50 pct tungsten, but about which little had been published on the high-tungsten side of the system. Powder metallurgical methods were used throughout this investigation for the preparation of the alloys, utilizing electrical sintering techniques such as are used for tungsten and molybdenum. In the course of the work it was found desirable to determine fusion temperatures of the alloys, and it was possible from these and other data to draw a tentative constitutional diagram (solidus, mostly) applicable to the high-temperature part of the system. No attempt was made to determine solid solubility vs. temperature data, nor to investigate the lower temperature transformations in the solid state. Data on the microstructure, hardness, and a note on oxidation resistance of the sintered alloys were obtained. PAST WORK Müller1 determined liquidus temperatures of platinum-rich alloys to 50 pct tungsten, using a radiation method employing a photoelectric cell, and found that the melting points increased progressively from 1771°C for platinum to 2350°C for the 50 pct alloy. Microscopic evidence by Müller that complete solid solutions exist to 50 pct tungsten was not checked by Kremer2 who, according to Vines and Wise,3 found that "although the 8 pct tungsten alloy was a solid solution, the 26 pct tungsten alloy contained a eutectic or an intermediate phase." Hultgren and jaffee4 using X ray methods found continuous solid solutions to at least 33.6 pct tungsten (35 at. pct) after annealing in vacuum for 6 days at 1000°C. Vines and Wise3 report work communicated to them by E. C. Rhodes on wrought alloys to 7 pct tungsten, which showed that tungsten hardened platinum extremely rapidly, and was one of the most potent hardeners of the binary additions to platinum. Considerable quantities of 4 pct tungsten platinum wire have been produced for grids in radar tubes and for aircraft spark plugs. Some of the alloy for the latter application is made by sintering, the remainder by smelting and casting. MATERIALS The platinum powder used in this work was kindly supplied by the research and Development Laboratory of the Owens-Corning Fiberglas Corp., Newark, Ohio, It was prepared from scrap platinum. After dissolving in aqua regia and evaporating almost to dryness, the solution was worked up several times with concentrated HCI. the resultant platinic chloride was diluted considerably, and metallic zinc was added, which precipitated the platinum in a very finely divided form. After dissolving the

Jan 1, 1948

By John J. Collins

[CONTENTS PAGE Purpose and scopeI Core samplingI Sludge sampling.....7 Combining core and sludge assays13 Deviation and surveying of drill holes17 Conclusions24 Acknowledgments25 Bibliography25] PURPOSE AND SCOPE The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described 25,26,31,32.50,65 but more information is needed on the results shown by actual mining operations. CORE SAMPLING The accuracy of diamond-drill-hole sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says, in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change, often indicated by a color change in the sludge water, should be a signal for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By R. I. Jaffee, H. P. Nielsen

In recent years much attention has been given to high melting alloy systems. It has been of interest in this respect to investigate the alloys of platinum and tungsten, which were known from the literature to have very high melting points up to about 50 pct tungsten, but about which little had been published on the high-tungsten side of the system. Powder metallurgical methods were used throughout this investigation for the preparation of the alloys, utilizing electrical sintering techniques such as are used for tungsten and molybdenum. In the course of the work it was found desirable to determine fusion temperatures of the alloys, and it was possible from these and other data to draw a tentative constitutional diagram (solidus, mostly) applicable to the high-temperature part of the system. No attempt was made to determine solid solubility vs. temperature data, nor to investigate the lower temperature transformations in the solid state. Data on the microstructure, hardness, and a note on oxidation resistance of the sintered alloys were obtained. Past Work Miillerl determined liquidus temperatures of platinum-rich alloys to 50 pct tungsten, using a radiation method employing a photoelectric cell, and found that the melting points increased progressively from I771 C for platinum to 2350°C for the 50 pct alloy. Microscopic evidence by Miiller that complete solid solutions exist to 50 pct tungsten was not checked by Kremer2 who, according to Vines and Wise,3 found that "although the 8 pct tungsten alloy was a solid solution, the 26 pct tungsten alloy contained a eutectic or an intermediate phase." Hultgren and Jaffee4 using X ray methods found continuous solid solutions to at least 33.6 pct tungsten (35 at. pct) after annealing in vacuum for 6 days at I000 C. Vines and WiseS report work communicated to them by E. C. Rhodes on wrought alloys to 7 pct tungsten, which showed that tungsten hardened platinum extremely rapidly, and was one of the most potent hardeners of the binary additions to platinum. Considerable quantities of 4 pct tungsten platinum wire have been produced for grids in radar tubes and for aircraft spark plugs. Some of the alloy for the latter application is made by sintering, the remainder by smelting and casting. Materials The platinum powder used in this work was kindly supplied by the Research and Development Laboratory of the Owens-Corning Fiberglas Corp., Newark, Ohio. It was prepared from scrap platinum. After dissolving in aqua regia and evaporating almost to dryness, the solution was worked up several times with concentrated HC1. The resultant platinic chloride was diluted considerably, and metallic zinc was added, which precipitated the platinum in a very finely divided form. After dissolving the

Jan 1, 1949

By Earnshaw Cook, J. A. Fellows, Howard S. Avery

Heat-resistant alloys of the higher nickel and chromium ranges have been empirically developed through the practical experience of the past two decades to a position of significant industrial importance. Few quantitative metallurgical data have been published, however, that could be related directly to engineering specifications and design. Inadequate information led to the inauguration, in 1934, of a research directed toward a better understanding of the properties and performance of these steels at temperatures between 1400° and 2000°F. Investigation was concentrated upon the 26 per cent Cr:I2 per cent Ni type, which owes its extensive application to an excellent combination of economy, strength and surface stability at elevated temperatures. The requirements of the oil-refining industry have been prominent in influencing the scope and evolution of the program. Creep strength at the maximum allowable deformation ranges has been employed to evaluate the merits of these alloys. Permissible elongation rates vary from 1.0 per cent in one million hours for turbine blades to I .o per cent in ten thousand hours for tube supports and heating-furnace equipment. Throughout this paper, the limiting creep stress (L.C.S.) for a given temperature has been defined as the stress that will produce a uniform elongation rate (stage 11) of one per cent in ten thousand hours as extrapolated from rates per hour for tests of less than two thousand hours duration. Certain current specifications require a minimum L.C.S. of 1600 lb. per sq. in. at I800°F., of 3600 lb. per sq. in. at 1400°F., and include a tensile acceptance test at room temperature, after aging for 24 hr. at 1400°F., as a measure of embrittlement in service. An elongation of 4 per cent is considered satisfactory for general use, while a minimum of 9 per cent is prescribed on occasion. Many differences of opinion concerning the properties and performance of these alloys appear to result from their sensitivity to unappreciated variations of chemical analyses which, for example, may change their L.C.S. from 3500 to 350 lb. per sq. in. at 1800°F. Within broad chemical limits, it is possible to balance the six or more important elements present to attain desirable properties with a number of different combinations. One method for adjusting the composition has been described under the sponsorship of the Alloy Casting Institute,' in which, by maintaining a substantially austenitic* alloy, optimum creep strength and minimum tendency to embrittle after aging are expected. Unless, however, the limits are undesirably close, it has not been possible to ensure a moderate range of strength and of ductility by any chemical specification alone. There is a definite need

Jan 1, 1942

By J. A. Fellows, Earnshaw Cook, Howard S. Avery

Heat-resistant alloys of the higher nickel and chromium ranges have been empirically developed through the practical experience of the past two decades to a position of significant industrial importance. Few quantitative metallurgical data have been published, however, that could be related directly to engineering specifications and design. Inadequate information led to the inauguration, in 1934, of a research directed toward a better understanding of the properties and performance of these steels at temperatures between 1400° and 2000°F. Investigation was concentrated upon the 26 per cent Cr:I2 per cent Ni type, which owes its extensive application to an excellent combination of economy, strength and surface stability at elevated temperatures. The requirements of the oil-refining industry have been prominent in influencing the scope and evolution of the program. Creep strength at the maximum allowable deformation ranges has been employed to evaluate the merits of these alloys. Permissible elongation rates vary from 1.0 per cent in one million hours for turbine blades to I .o per cent in ten thousand hours for tube supports and heating-furnace equipment. Throughout this paper, the limiting creep stress (L.C.S.) for a given temperature has been defined as the stress that will produce a uniform elongation rate (stage 11) of one per cent in ten thousand hours as extrapolated from rates per hour for tests of less than two thousand hours duration. Certain current specifications require a minimum L.C.S. of 1600 lb. per sq. in. at I800°F., of 3600 lb. per sq. in. at 1400°F., and include a tensile acceptance test at room temperature, after aging for 24 hr. at 1400°F., as a measure of embrittlement in service. An elongation of 4 per cent is considered satisfactory for general use, while a minimum of 9 per cent is prescribed on occasion. Many differences of opinion concerning the properties and performance of these alloys appear to result from their sensitivity to unappreciated variations of chemical analyses which, for example, may change their L.C.S. from 3500 to 350 lb. per sq. in. at 1800°F. Within broad chemical limits, it is possible to balance the six or more important elements present to attain desirable properties with a number of different combinations. One method for adjusting the composition has been described under the sponsorship of the Alloy Casting Institute,' in which, by maintaining a substantially austenitic* alloy, optimum creep strength and minimum tendency to embrittle after aging are expected. Unless, however, the limits are undesirably close, it has not been possible to ensure a moderate range of strength and of ductility by any chemical specification alone. There is a definite need

Jan 1, 1942

By M. M. Neale

The pottery department has been aptly described as the "heart of the zinc plant," md every operator of a horizontal-retort plant is acquainted with the fact that good retort practice is the keystone of the smelting process. Since retorts and condensers are clay products, it is obvious that improvements in other branches of the ceramic industry should be observed for possible adoption in the pottery process for zinc smelting. Deaeration of clay bodies is one of the improvements that have become fully, established in many brick and pottery plants. Our investigation of a number of these installations always yielded the same answer—improved practice. Improvement in workability, elimination of laminations, less porosity, greater wet and dry strength, were among the claims put forth for this process. All of these features are exceedingly desirable characteristics in the manufacture of a zinc retort. A preliminary test, which involved truck transportation of deaired slugs over a relatively long distance, did show the value of many of these characteristics in forming the retort. Results from the standpoint of , longer retort life in the smelting furnace were not outstanding. The idea lay dormant for several years but was revived again in 1940, when it was decided to make whatever changes were necessary in the equipment for pottery manufacture to fully demonstrate the process. A number of the manu- facturers of machinery for clay work were asked to submit proposals, and the plan finally adopted was extremely simple and economical. Machine FOR DeaeRRATING Selection of the ''Straight-line'' principle for conversion of the "Wettengel" pug mill to a ''vacuum-extrusion" machine preserved some usable parts of the original equipment, such as the drive mechanism, and main shaft. The converted machine consists of a feed hopper, a compression and shredding zone, a vacuum chamber, and an extrusion barrel, as shown in Fig. I. In operation the clay mix travels from the feed hopper into the compression zone, where it is compressed and finally forced around the tapered cone into the vacuum chamber. As it enters this chamber the clay body is shredded into thin plates, for exFosure to the vacuum. The shredder consists of a hub mounted on the main shaft, to which are bolted segmental sections, each of which carries shredder knives. The compression zone and extrusion barrel function as seals for the vacuum chamber and the shredded clay mix is agitated thoroughly during its exposure to the vacuum. A vacuum pump, equipped with a 3-hp. motor, is attached to this chamber, and maintains 25 to 26 in. vacuum during operation of the mill. Changes in Practice Prior to the adoption of the deaeration process, our standard pottery practice called for double mixing in the combination 7

Jan 1, 1944

By John J. Collins

Page Purpose and scope..................... 521 Core sampling......................... 521 Sludge sampling....................... 527 Combining core and sludge assays ...... 533 Deviation and surveying of drill holes. . . . 537 Conclusions........................... 544 Acknowledgments...................... 545 Bibliography.......................... 545 Purpose and Scope The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described,25,26,31,32,50,65 but more informa- tion is needed on the results shown by actual mining operations. Core Sampling The accuracy of diamond-drill-hole Sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says; in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change; often indicated by a color change in the sludge water, should be a a for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

Prospectors were digging silver in the Ray mine area in 1873, and by 1880 high-grade copper ore was feeding a 30-ton copper furnace. In 1910, D. C. Jackling and his associates organized the Ray Consolidated Copper Company, known in mining circles as Ray Con, to exploit the Ray orebody. For 50 years, all mining was underground and the Ray mine was known the world over for innovations in mining methods. It was the first to mine important tonnage (8,000 tpd) by block caving methods. The mill at Hayden, 26 miles (41.6 km) from the mine, kept pace with the new mining methods developed at Ray and in the early 1920's flotation replaced gravity concentration by the vanner machines. First mining by open pit methods began in 1950 and in January 1955 all underground mining was discontinued. The new primary crushing plant, installed in early 1967, featured design simplicity with a low capital cost. The 54 in. gyratory crusher was installed within a total height of 6 1 ft (18.65 m) from the truck dump road bed to the floor beneath the feeder-conveyor under the surge pocket. The crusher is fed by 120-ton trucks dumping from one side only into a 200-ton hopper equipped with a horizontal splitter bar to protect the crusher spider. The crushed ore pocket, of 120 -ton live capacity underneath the crusher, discharges to a 0 to 200 fpm (61.2 mpm), variable-speed, 72 in. by 80 ft long (1.83 m by 24.4 m) sloped belt feeder, probably a first in the industry. Wood-picking stations are provided on either side of the belt so that old mine timbers and tramp iron can be removed from the crushed ore and the wood can be dropped into Dempster containers located below the belt. Counter-weighted finger gates control the bed depth on the belt at 14 in. (356 mm) and a low- level radiation unit automatically maintains a minimum of a 3 ft bed (900 mm) of material on the belt directly under the surge pocket. The conveyor is skirted for its full length and is approximately 60 in. (1.52 m) wide inside its metal skirt liners. Ten impact idlers at 12 in. (305 mm) spacing are provided under the loading pocket and remaining idlers are spaced at 18 in. (457 mm). The belt is 1 ¼ in. (31. 8 mm) thick; and, while it has occasionally been damaged by tramp material, belt life averages 11 million tons. When the belt is changed, the coiled new belt is lowered into position onto one of two belt stands mounted over the feeder, the old belt is cut, and one end is attached temporarily to the new belt. The other end of

Jan 1, 1978

By J. F Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By W. David Menzie

An inventory of the mineral resources of a region is a most desirable basis for planning their systematic development. In the absence of appropriate samples upon which to base a reliable estimate of this inventory, investigators have resorted to a number of indirect approaches; the unit regional value concept forms the basis for one such approach. The produced resources of a region are a function of the inventory multiplied by the degree of intensity (some fraction between zero and one) of development of the resources. A convenient means of expressing this amount is to accumulate the weight and value of resource and to prorate the totals over the area represented by the unit sampled; i.e., weight and value per unit area or tonnes/km.2 and U.S. 1967 $/km.2. When this approach is used for the U.S.A., South Africa and Rhodesia the average value of (log.) U.S. 67$/km.2 for the states of the U.S.A. and South Africa are similar ($55,176 U.S.$67/km.2 and $52,538/km.2 respectively); whereas, Rhodesia is relatively underdeveloped (10,301 U.S.67$/km.2). On the other hand if the individual commodities are used as a basis for comparison, the U.S.A. shows the greatest diversity (67 commodities out of a possible total of 75) with South Africa next (50 commodities) and Rhodesia last with only 26 commodities. Since it is known that area and value are not simply related perhaps diversity is directly related to area. Again the kinds of commodities which have contributed the greatest value to the totals differs, with fuels (petroleum, coal and natural gas) predominating in the U.S.A. totals; whereas, South Africa and Rhodesia obtain their greatest contributions from gold, diamonds (S. Africa) coal and asbestos. In each case the greatest value of resources is produced from sedimentary rocks. As a somewhat different kind of example, consider a single resource (petroleum, the most valuable in the U.S.A.) over a large area (number of sampling units) the U.S.A. by state. Furthermore, assume that the model frequency distribution of the amount of petroleum (tonnes/km.2) is log normal. Then the present production fails to match the model first because 17 states (34% out of 50) have not produced petroleum. Of the remaining 33 states, Louisiana has produced the most (15,177 tonnes/km.2) and the frequency histogram is still skew. A spatial analysis of the distribution of the amount of petroleum by state shows that the samples (i.e., states) are not random; the non-producers and producers tend to cluster. It remains to be seen if this is indeed a true representation of the incidence of petroleum or is a function of the exploration philosophy currently in vogue.

Jan 1, 1977

By L. B. Herring, Harold Decker

The area under discussion includes districts 2 and 4, so designated by the Texas Railroad Commission,‡ and comprises the following 26 counties: Bee, Brooks, Calhoun, Cameron, De Witt, Duval, Goliad, Gonzales, Hidalgo, Jackson, Jim Hogg, Jim Wells, Karnes, Kenedy, Kleburg, Lavaca, Live Oak, Nueces, Refu-gio, San Patricia, Starr, Victoria, Webb, Willacy, Wilson and Zapata. Drilling and Production The total number of wells drilled in this district during 1944 is shown in Table 5. This indicates a 41 per cent increase in all drilling over 1943. In spite of every handicap created by the war, the oil industry more than met the crisis in its drilling program in this district. Last year we thought it would be impossible to add to the total number of wells drilled in the previous yea;, but in spite of the lack of manpower and materials this large increase was shown. During 1944 this district produced 124,201,978 bbl. of oil from 11,607 producing wells, whereas in 1943 it produced 88,421,693 bbl. from 11,211 producing wells. This is an increase of 35,780,285 bbl. of oil. Some decline has been shown in the older fields arid undoubtedly the majority of wells in the district are producing at too high a rate. However, this district has been able to meet any demands made on it by the Petroleum Administration for War for increased production. Discoveries and Extensions There was a slight decline over the preceding year in the number of wildcats drilled, but the number of discoveries increased. In most instances subsequent development has not been sufficient to arrive at preliminary valuations but it appears that the over-all oil discoveries were below normal and that the gas discovcrics were relatively more important. The Canolis and Tijerina pools, Jim Wells County, and the Dan Sullivan pool, Brooks County, appear to be the outstanding oil discoveries of the year. No important oil reserve was discovered or developed in the Wilcox horizons, although considerable wildcatting was done. Important extensions were made at Agua Dulce and Stratton in Nueces and Teleberg Counties. Production was increased appreciably at Willimar, Willacy County, and Seeligson, Jim Wells County, in sands and areas which were apparently productive at the beginning of the year. Refining Most of the major work in the changing and completing of refineries was finished this past year. Some of the most modern refineries in the industry are now located in the Corpus Christi area. Gas and Condensate No new recycling projects were commenced during the year, nor were there any important additions to the old operations. No major gas reserve was developed, although several new discoveries suggest potentid possibilities.

Jan 1, 1945