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By K. S. Seljesater

Since the development of dispersion-hardenable lead-antimony alloys' in the laboratories of the Western Electric Co., Inc., studies have been made of the effect of various third constituents on these alloys.2 This paper covers a study of the effect of small amounts of arsenic. This element is important since it occurs as an impurity in many commercial antimonial leads and the results obtained are particularly interesting as indicating a class of addition agents which exert a stabilizing effect on dispersion-hardened alloys. Experimental In order to study the effect of increasing amounts of arsenic, a series of alloys were made up containing 1 per cent. antimony, and arsenic varying from 0. to 0.10 per cent., the balance Doe Run lead (99.997 per cent. lead). The alloys were cast into bars 1 1/2 in. wide by 3/4 in. thick and cold-rolled to about 1/8 in- thickness. Three samples of each composition were annealed 1 hr. at 225" C. and quenched in water. One set was aged at room temperature, one at 50" C., and the third at 100" C., the

Jan 1, 1929

By J. R. Campbell

It is the purpose of this paper to review somewhat in detail the literature on the subject that is extant, which ought to provoke considerable beneficial discussion. The value of 1 per cent. ash in blast-furnace practice was brought to the attention of the writer when he was connected with the U. S. Steel Corpn. Various groups of blast furnaces in the Pittsburgh district compiled a statement in which the average saving was about 6 c. per ton on pig iron for every 1 per cent. of ash reduction. Later, the same groups of furnaces, when materials had reached a higher price, compiled a statement in which there was a 30-cent saving on pig iron for every 4 per cent. ash reduction in the Pittsburgh district, which was equivalent to 7½ c. for every 1 per cent. ash reduction. In the Chicago district, on long freight hauls of coal and coke from the Pocahontas district, the set-up was 30 c. per ton of pig iron for every 2 per cent. ash reduction, or 15 c. for every 1 per cent. Some years ago, Ralph H. Sweetser, of the American Rolling Mill Co., presented a paper which provoked considerable comment.' To get the matter again before this body a résumé of his computations is given below: Comparison of Pig Iron Costs Ash in Coal. Ash in CoKe, Fixed Carbon in Coke peR Ton OB PIG Iron peN Day, Per Cent. PeR CeNt. Coke. Per Cent. PIG IRon. Lb. Tons 6.00 8.70 89.80 1800 365.0 7.00 10.15 88.35 1860 353.2 Coke with 8.70 Per Cent. Ash Ore mixture, 4144 It). = 1.85 tons @ $5.50.............................. $10.17 Coke, 1800 lb. = 0.90 ton @ $6.50..................................... 5.85 Limestone, 1000 lb. = 0.50 ton @ $1.50................................. 0.75 Labor, $365 per day on 365 tons....................................... 1.00 Supplies and service, $250 per day...................................... 0.68 Overhead, $200 per day............................................... 0.55 Reserve for relining and depreciation................................... 1.00 Total............................................................. $20.00

Jan 1, 1931

By Paul T. Allsman

1.1-1. Current Status. The history of surface mining is essentially that of mining coal, copper, and iron ores, and the nonmetallic minerals--clays, gypsum, phosphate rock, sand, gravel, and stone. Table 1.1-1 is a comparison of the estimated world production of crude metallic and nonmetallic ores and coal with that of the United States for 1964, in short tons. It includes an estimate of the quantity produced by surface mining methods. Two-thirds of the 8.8 billion tons are mined from surface, ranging from a third of the coal, half the metallic ores, and nearly all clays, stone, sand, and gravel. In the United States, production by surface methods parallels that of the world with two major exceptions: 74% of copper ore and 90% of usable iron ore are mined from the surface, whereas outside the country less than 40% of copper ore and only 50% of usable iron ore are mined from the surface. In surface mining of both metallic ores and coal, much waste is stripped, but in mining nonmetallics the ratio between waste and ore is small. Assuming that the stripping ratio for surface mining in the United States will apply equally to world production, it is estimated on a weighted basis that over 21 billion tons of waste, or a total of nearly 27 billion short tons of material, are mined annually in the world by surface methods. Data are not available to confirm the assumption for waste stripping outside the U.S., but the estimate expresses an order of magnitude that is worth noting. Table 1.1-2 shows the materials-crude ore, coal, and waste-mined by surface methods and the relation to total mined by both surface and under- ground methods in the United States during 1964. Eighty-two percent of all crude metallic and nonmetallic ores and coal was mined from the sur- face, including 82% of metallic, 96% of nonmetallic ores, and 35% of coal. Copper and iron ore comprise 83% of the metallic ores mined from the surface, and clays, gypsum, phosphate rock, sand, gravel, and stone

Jan 1, 1968

By George Burgess

NOTE BY THE EDITOR.-This resume of a Technologic Paper which is published in full by the U. S. Bureau of Standards, is brought before the membership of the Institute with the object of affording an opportunity for the open discussion of results of an investigation by the authors named above, of some phases of the manufacture of steel rails which are of great importance to manufacturer and consumer alike. This resume will be presented at the Pittsburg Meeting of the Institute. Discussion is invited, by either members or non-members of the Institute. The discussion may be sent in writing to the Editor, in advance of the meeting, or should preferably be presented in person at the meeting. To those who desire to discuss, copy of the full Technologic Paper (No. 38), as published by the Bureau of Stanaards, will be sent upon request addressed either to the Bureau of Standards, Washington, D. C., or to the Editor, American Institute of Mining Engineers, 29 West 39th Street, New York, N. Y. Summary of Researches and Conclusions The main objects of this investigation were to determine, from measurements taken at representative rail mills, the present American practice regarding the temperatures at which rails are rolled; to demonstrate the ease and accuracy with which such temperatures may be measured, without interfering in any way with the manufacture of the product; to find out what the "shrinkage clause" in rail specifications really means; and, finally, to determine some of the physical properties of rail steels, particularly those of interest in manufacture, some of which seem to be insufficiently known as yet. 1. Ingot Temperatures.-Measurements made in four mills show that ingots for rails are rolled at temperatures ranging from 1,075° to 1,150° C. (1,965° to 2,100° F.), and that, in a series of 20 to 40 ingots, the variation in temperature from one ingot to another is only 10° to 20° C. 2. Finishing Temperatures.-Rails are finished at temperatures ranging broadly from 880° to-1,050° C. (1,615 to 1,922° F.), but usually ranging within 50° C. of 935° C., and occasionally above 1,100° C.

Jan 9, 1914

By F. Habashi

Conclusive evidence is given showing that the dissolution of gold in cyanide solution follows mainly the over-all equation A similar equation can also be written for the dissolution of silver. Theoretical derivation of the velocity equation for this reaction has been obtained, which describes quantitatively the experimental facts. The equation is as follows: where [CN-] and [O2] = the concentrations (in moles/ml) of cyanide and dissolved oxygen, respectively. DcN- and Do2 = the diffusion coefficients of cyanide and dissolved oxygen; 1.83 x 10-5 and 2.76 x 10-5 cm2 sec-1, respectively. A = the surface area of the metal in contact with the aqueous phase, in cm S = the thickness of the boundary layer; varies between 2 and 9 x 10-3 cm, depending on the speed and method of agitation. Rate being expressed in g. equiv. sec-1. INTRODUCTION During recent years a number of papers have been published on the kinetics of dissolution of gold and silver in cyanide solution.1,6 The authors agree that the rate of dissolution is diffusion-controlled, but disagree regarding the mechanism of the reaction. Some authors'-3 support the Bodlander equations: The apparently conflicting results, when critically reviewed, are concordant and lead to the conclusion that the proper equation for the dissolution reaction should be only Eq. 1. This conclusion is based on the following facts: 1) For every 2.2 equivalents of metal dissolved one mole O2 was consumed. 2) For every 1 equivalent of metal dissolved, 2 moles of cyanide were consumed.' 3) Hydrogen peroxide is formed during the dissolution of gold or silver, and for every 2.2 equivalents of metal dissolved, one mole H2O2 was produced.1,2 The fact that Kudryk and Kellogg4 did not detect H2O2 during the dissolution of gold should not throw doubt on the many affirmative reports of other investigators. Lund2 has shown that depending on the crystalline nature of the silver surface, (e.g. method of etching the sample) H2O2 may undergo catalytic decomposition heterogeneously according to: Further Kudryk and Kellogg4 carried out their experiments in the presence of 0.5% KC1 to achieve a greater electrolytic conductivity, and Matsudaira7 has shown that H2O2 is catalytically decomposed (homogeneously) in the presence of C1- ion in alkaline solution. 4) Experiments carried out by Boonstra8 and Lund2 showed that the dissolution of gold and silver respectively in KCN + H2O2 in the absence of oxygen was a slow process. Thus the reaction takes place to a minor extent only. Further kinetic studies showed that: 1) The rate of dissolution depends on the surface area of the metal in contact with the liquid phase, thus indicating that the process of dissolution is a heterogeneous process. 2) The rate of dissolution depends on the speed of stirring,5,6 thus indicating that the process is controlled by a physical phenomenon.

Jan 1, 1967

By H. W. Schull

The Alligator Ridge gold deposits are located in northwest White Pine County at the southwest end of Alligator Ridge. Alligator Ridge is on the west side of Long Valley in the range of hills that separates Long and Newark valleys. The nearest population center is Ely, about 113 km (70 miles) to the southeast. Access to the Alligator Ridge gold mine from Ely is via US Highway 50 and the Long Valley Ruby Marsh Road. The latitude and longitude is approximately 39° 45' north and 116° 30' west. The Alligator Ridge gold deposits occur in the lower 61 m (200 ft) of the Mississippian-Devonian Pilot Shale formation, a laminated calcareous carbonaceous siltstone. The gold deposits contain carbonaceous, oxide, and jasperoid type gold ore. For more details on the geologic features of the deposits, see Schull, Sutherland, and Ilchik (in press). The discovery story of Alligator Ridge began in the spring of 1976 when the areas mapped as jasperoid on the 1:250,000 scale geologic map of White Pine County (Hose and Blake, 1976) were examined and sampled. Seventeen samples, about 23 kg (50 lb), were collected from an area about 0.65 km2 (0.25 sq mile). Of these 17 samples 10 reported atomic absorption values of less than 0. 1 ppm Au and 7 reported values in the 0.1 to 1.0 ppm Au range. In June of 1976 20 lode claims were staked to claim the area of the rock chip sampling. The claims' end and side center lines were used as a grid for soil sampling. Sample stations were 91 m (300 ft) apart and sample lines were 229 m (750 ft) apart. The samples were sieved to - 180 µm (- 80 mesh). Out of 99 soil samples, seven samples reported Au values in the 0.1 to 1.0 ppm range. Of the seven auriferous samples, five were located over the jasperoid outcrops previously sampled; one over unmineralized Pilot Shale a few hundred m (ft) downslope from some Au-bearing jasperoid; and one from a 15 x 15 m (50 x 50 ft) talus outcrop area of hitherto unnoticed mineralized Pilot Shale. Subsequent sampling, mapping and drilling would show that this 15 X 15 m (50 x 50 ft) outcrop area was the top of the Vantage 1 ore body, about 1.27 Mt (1,400,000 st) of over 3 g/0.9 t (0.11 oz/st) Au. Late in the summer of 1976 about 30 more 1.4 to 2.2 kg (3 to 5 lb) rock chip samples were collected in the course of preliminary geologic mapping. This mapping and sampling showed that the gold-bearing jasperoids occurred at the contact between the Pilot Shale and the underlying Devils Gate Limestone. Rock chip samples of 0.1 ppm Au or greater were restricted to the jasperoids and the mineralized outcrop area of Pilot Shale located by the soil sample survey. Rock chip samples from the area of mineralized Pilot Shale were in the 3 to 5 ppm Au range. Activity resumed in the summer of 1977 with a more detailed soil geochemical survey. The jasperoid areas were defined as large geochemical gold anomalies of values greater than 0.1 ppm and locally greater than 0.2 ppm Au. Soil samples from the area of mineralized Pilot Shale outlined a greater than 0.1 ppm gold anomaly 61 x 152 m (200 x 500 ft) in size. Further geologic mapping showed that elsewhere, even where bleached and silicified, the Pilot Shale areas were geochemically very low in gold (less than 0.1 ppm Au). In November, 1977 twelve rotary percussion drill holes were drilled to test the gold-bearing jasperoid areas and the mineralized Pilot Shale area. Drilling of the jasperoid areas showed that the jasperoids gave way at shallow depths [less than 24 m (80 ft)] to unmineralized unaltered Devils Gate Limestone. A five hole cross pattern was drilled to test the area of mineralized Pilot Shale. Four of the five holes re-

Jan 1, 1985

By Didyk M. Marko, Vásquez Ñ. Gregorio

INTRODUCTION The Salvador Division of Codelco Chile is an important copper molybdenum production center, located 1,200 km north of Santiago and 124 km east of the Pacific Ocean, in the foothills of the Andes Range. El Salvador Mine produces by block caving 8,600,000 t of ore with a content of 1 .l0% Cu and 0.02% Mo, from a copper porphyry ore body. The geology of the ore body shows a great heterogeneity as to rocks, grade distribution, ore height and undercutting levels. This has resulted in a variety of designs for the productive units (blocks) and in the mineral handling systems. Up to now, 4 systems have been employed: grizzly and intermediate haulage; slusher drifts with intermediate ore pass and haulage; slusher drifts with scraper loading directly into trains; and, LHD. On the other hand there is a marked dispersion in the final results from each block mined out. We report here the conclusions reached after analyzing the data gathered on the mining of 163 depleted blocks, representing a total production of 102,000,000 t, with an undercut area of 340,000 sq m. The final results of each block, characterized in terms of tonnage and fine cop- per recovery, were studied as a function of a number of design, geological, and operational parameters. BASIC CONCEPTS Indices of Efficiency In order to evaluate the final result of each block, or group of blocks, the following indices of efficiency were defined : Tonnage Recovery (TR) . It is the ratio between the ore tonnage extracted and the in situ tonnage estimate for each block or group of blocks. Grade efficiency (E) . It is the ratio between the Cu mill feed grade and the in situ Cu grade estimate for each block or group of blocks. Grade efficiency is numerically equal to the ratio between fine copper recovery (FR) and tonnage recovery (TR) . Figure 1 shows the final results for 163 mined out blocks, in terms of tonnage and fine copper recovery. Different grade efficiencies are also indicated. The marked dispersion in the final results of mined out blocks must be stressed since it greatly incides in mine planning. It must be kept in mind that an eventual concentration of low overall efficiency in the blocks during a given period (e.g., one year) will make it necessary to bring into production new blocks over the standard development schedule. It is worth noting that the final results in the 163 mined out blocks corresponding to a production of 102,000,000 t of ore with a mill feed grade of 1.445% were obtained from an ore estimate of 93,000,000 t and a grade of 1.671% Cu, which represents 110% tonnage recovery and 86% grade efficiency. Parameters Studied The following parameters are considered in this study: Geological Parameters. These parameters depend exclusively on the geological and geotechnical features of the ore body. IN SITU GRADE: The initial estimate of the block grade. Four copper grade categories were considered : (a) lower than 1.3% ; (b) 1.3 to 1.6%; (c) 1.6 to 2.0%; (d) higher than 2.0%. The overall initial in situ estimate of the ore body was 1.6% Cu. IN SITU TONNAGE: Initial tonnage estimate of the block. Four categories are considered: (a) below 500,000 t; (b) 500,000 to 1,000,000 t (c) 1,000,000 to 1,500,000 t; (d) over 1,500,000 t. CONTACTS SLOPE: Since up to 4 rock types can be found in a block, resulting in a corresponding number of contacts, a horizontal con- tact was defined as that which forms an angle of less than 45" with the horizontal and a vertical contact that which forms an angle of more than 45" with the horizontal. GRADE DISTRIBUTION: As in rock types, horizontal and vertical grades were defined. ORE HEIGHT: The distance measured between the undercut level and the top of ore. Four categories are considered: (a) less than 60 m; (b) 60 to 100 m; (c) 100 to 150 m; (d)over 150 m

Jan 1, 1981

By W. R. Davis

This paper presents a sunimary of the results of a study of the Dornick Hills-Springer sand reservoir of the Velma Pool. Stephens County, Oklahoma. The data are of interest because this type of reservoir is unusual in the mid-continent area. The reservoir is characterized by very steep dips and a relatively thick producing section. Its up-dip productive limit is defined by truncation of the productive sands, and the down-dip limit is defined by deterioration of sand quality rather than by the presence of an oil-water contact. The crude oil is of 29" API gravity, is relatively viscous and was saturated with dissolved gas at the original reservoir pressure. Consideratiqn of the physical characteristics of the reservoir indicates that natural water encroachment will be negligible, and that the producing mechanism will be a dissolved gas drive augmented by gravity drainage. Performance to date tends to confirm this. INTRODUCTION It is the purpose of this paper to present some of the features of the Dor-nick Hills - Springer sand reservoir which are unusual in mid-continent operations. The study was undertaken in an attempt to define the producing mechanism or mechanisms that will control reservoir performance, as an aid in determining well completion and production practices. GEOLOGY The reservoir under consideration is a series of sand bodies of Pennsylva-nian age found in the Lower Dornick Hills and/or Springer formations. For the purpose of this paper, the sand series is referred to as the Dornick Hills-Springer sand. This sand series lies on the northeast flank of the Velma anticline, and dips very sharply to the northeast in TWP 1S-5W, the average dip being about 45º. However, as shown in Fig. 1. the trend of the pool swings eastward in TWP 1S-4W with a coincident de- crease in dip angle. In this portion of the pool the average dip is about 25". Faults in Sections 30-1S-4W and 32-1S-4W are shown in Fig. 1. These faults show only minor displacements for the horizon contoured in Fig. 1, but they do show a pronounced displacement of the immediately overlying beds. Similar faults are thought to exist in TWP 1S-5W; however, since the detailed geological studies necessary to properly define these faults were not available to the author, no attempt to show them in Fig. 1 has been made. The up-dip productive limit throughout the pool is determined by the truncation of the Lower Dornick Hills-Springer sand. The down-dip productive limit does not appear to be defined by edge water, but by a gradual reduction in permeability of the sands. No dry hole has produced a significant volume of water from this sand section. There was a small gas cap in TWP 1S-5W. Most of the wells in TWP 1S-SW encountered an oil sand 70 ft to 150 ft thick, a shale 20 ft to 80 ft thick, and another oil sand 50 ft to 130 ft thick. A few scattered wells in this area encountered other and bodies below these

Jan 1, 1951

By W. R. Davis

This paper presents a sunimary of the results of a study of the Dornick Hills-Springer sand reservoir of the Velma Pool. Stephens County, Oklahoma. The data are of interest because this type of reservoir is unusual in the mid-continent area. The reservoir is characterized by very steep dips and a relatively thick producing section. Its up-dip productive limit is defined by truncation of the productive sands, and the down-dip limit is defined by deterioration of sand quality rather than by the presence of an oil-water contact. The crude oil is of 29" API gravity, is relatively viscous and was saturated with dissolved gas at the original reservoir pressure. Consideratiqn of the physical characteristics of the reservoir indicates that natural water encroachment will be negligible, and that the producing mechanism will be a dissolved gas drive augmented by gravity drainage. Performance to date tends to confirm this. INTRODUCTION It is the purpose of this paper to present some of the features of the Dor-nick Hills - Springer sand reservoir which are unusual in mid-continent operations. The study was undertaken in an attempt to define the producing mechanism or mechanisms that will control reservoir performance, as an aid in determining well completion and production practices. GEOLOGY The reservoir under consideration is a series of sand bodies of Pennsylva-nian age found in the Lower Dornick Hills and/or Springer formations. For the purpose of this paper, the sand series is referred to as the Dornick Hills-Springer sand. This sand series lies on the northeast flank of the Velma anticline, and dips very sharply to the northeast in TWP 1S-5W, the average dip being about 45º. However, as shown in Fig. 1. the trend of the pool swings eastward in TWP 1S-4W with a coincident de- crease in dip angle. In this portion of the pool the average dip is about 25". Faults in Sections 30-1S-4W and 32-1S-4W are shown in Fig. 1. These faults show only minor displacements for the horizon contoured in Fig. 1, but they do show a pronounced displacement of the immediately overlying beds. Similar faults are thought to exist in TWP 1S-5W; however, since the detailed geological studies necessary to properly define these faults were not available to the author, no attempt to show them in Fig. 1 has been made. The up-dip productive limit throughout the pool is determined by the truncation of the Lower Dornick Hills-Springer sand. The down-dip productive limit does not appear to be defined by edge water, but by a gradual reduction in permeability of the sands. No dry hole has produced a significant volume of water from this sand section. There was a small gas cap in TWP 1S-5W. Most of the wells in TWP 1S-SW encountered an oil sand 70 ft to 150 ft thick, a shale 20 ft to 80 ft thick, and another oil sand 50 ft to 130 ft thick. A few scattered wells in this area encountered other and bodies below these

Jan 1, 1951

By J. C. Fisher

THE following facts are reasonably well established concerning the influence of boron on the hardenability of steel:1-4 1—Boron does not appreciably influence the ther-

Jan 1, 1955

University of New Mexico, Albuquerque, N M Requests for publications should be addressed to the Librarian of the University Papers of the Geological Series of the University are bound in volumes: Vol 1, vol 2, pt 1, and vol 2, pt. 2, and vol. 3 are out of print, but separates of the following papers may be had at a uniform cost of 25 cents each. Vol. 1? No. 1, Geology of the environs of Albuquerque (1898); 2, Occurrence of copper and lead in the San Andreas and Caballo Mountains (1898), 3, Geological reconnaissance in western Socorro and Valencia counties (1898); Geology of the San Pedro and Albuquerque districts (1898). , Vol 2: No. 1, Geology of the Albuquerque Sheet (1900); 2, Miscellaneous economic papers by Herrick (1900); 3, Geological reconnaissance in western Socorro and Valencia counties (1900); 4, Geology of the White Sands of New Mexico (1900); 5, Identifi¬cation of an Ohio coal measure's horizon in New Mexico (1900); 6, Cyanide process in New Mexico (1900). Vol. 3. No 3, The oil situation in New Mexico (1920), 4, Geology of Sandia Mountains (1922); 5, Oil and gas m New Mexico in 1923.

Jan 1, 1933

By Harold M. Mooney

WHEN apparent resistivity data are taken with the symmetrical Wenner 4-electrode spread, a fixed center position is used and readings are taken for values of electrode separation. Basic data consist of apparent resistivity plotted against separation of adjacent electrodes . The interpreter attempts to infer geologic structure, such as the depth to discontinuities and the nature of subsurface earth materials. An earlier paper' described methods for interpreting resistivity data. All of these involve a severe assumption, namely that the earth in the region of interest consists of horizontal layers, electrically homogeneous and isotropic. The actual earth never satisfies this assumption exactly and may deviate from it so much that none of the above methods can be applied. Attempts in three directions have been made to modify the assumption so that it approaches known geologic complexity more closely. First, curves have been calculated for dipping discontinuities. Stern' and Aldredge hose a few widely separated dip angles. Trudu confined his attention to small dips. Berel'kovskiy and Zubanov computed gradient curves for widely separated dip angles. Unz as given the most complete solution, with a brief attempt to treat the three-layer case. Second, anisotropy can be taken into account. It seems geologically probable that layered materials have different vertical and horizontal conductivities. Cagniard, Maillet, and Pirson et up methods for finding an equivalent hypothetical isotropic medium. Standard interpretation methods can be applied to this, and the actual medium can then be deduced. Belluigi'" discounts the practical importance of anisotropy; Geneslay and Rouget do not agree with his conclusion. Third, the effect of variable resistivity in a layer can be considered. Keck and Colby" examine the mathematics of an exponential increase in a surface layer. Several authors, for example, Stevenson,'' consider a continuous variation of resistivity with depth. The present paper deals with a linear variation of resistivity in a surface layer. Geologically, surface variations should be expected. Unconsolidated materials such as glacial drift show marked irregularities over short distances. The effects of weathering change with depth. The moisture content of material above the water table may vary from a dry sand to a saturated clay, and both of these will be changed by rainfall. Figs. 1 to 6 present apparent resistivity curves to show the effect of a variable surface layer. In all cases resistivity varies linearly with depth down to a depth of one unit. Material beneath this depth has constant resistivity. Electrode separation is plotted in depth units. Insets on each figure show the corresponding cross-sections, plotting true resistivity against depth. To illustrate, consider curve A of Fig. 1. The true resistivity of material at the surface of the earth is taken as 0.4 units. Resistivity increases with depth, reaching a value of 1.0 at 0.5 depth units and 1.6 at 1.0 depth units. Below a depth of 1.0, all the material has very low resistivity (zero, for purposes of calculation). Curve A in the main part of Fig. 1 shows how apparent resistivity varies for this case as the electrode spacing is increased. To illustrate further, curve E of Fig. 4 corresponds to true resistivity of 1.2 units at the surface, 1.0 at a depth of 0.5 units, 0.8 at a depth of 1.0, and 1.5 for all depths greater than 1.0. Apparent resistivity curves have been plotted logarithmically so that the shape of the curves becomes independent of the units, giving the curves wide validity. A certain drift-covered area, for example, shows a gradual decrease of resistivity from 230 ohm-meters at the surface to 150 ohm-meters at the bedrock surface, 275 ft down; bedrock resistivity is 800 ohm-meters. Curve E of Fig. 5 indicates that true resistivity decreases from 1.2 units at the surface to 0.8 at a depth of 1.0, then increases abruptly to a constant value of 4.0 units. Since resistivity and depth ratios are the same, this can be used to predict the field curve. For Fig. 5, multiply true resistivity and apparent resistivity by

Jan 1, 1955

By P. Niessen, S. J. Bercovici

RECENTLY Aust et al.&apos; proposed a model of non-equilibrium grain boundary segregation based on a vacancy gradient induced uphill diffusion process of solute to grain boundaries. According to their proposal, solute atoms which are attracted to vacancies would, during the cooling of a sample, migrate with vacancies toward grain boundaries (vacancy sinks) and thereby enhance the solute concentration in boundary regions. For solute atoms which are not attracted to vacancies no such boundary solute enrichment would be expected. These concepts were given experimental support by boundary and bulk mi-crohardness measurements of various thermally treated dilute zinc and lead alloys and also doped aluminum oxide. The underlying assumption in their work was that solute enrichment in boundary regions is indicated by higher microhardness values. From thermodynamic considerations1&apos;2 it appears that the solid-liquid solute distribution coefficient (K), defined as the concentration of solute in the solid to the concentration of solute in the liquid at equilibrium, indicates whether attractive or repulsive forces exist between vacancies and solute atoms. K < 1 indicates a solute-vacancy attractions, K > 1 indicates no, or a repulsive, vacancy-solute interaction. This note is to report on observations of different grain boundary etching phenomena in high purity zinc alloys. These phenomena appear to depend on whether a K < 1 or K > 1 solute is present. In experiments designed to measure directly the solute enrichment in grain boundary regions, samples of zone refined zinc with 0.1 at. pct solute additions were cooled from 380°C at a rate of 3°C per min. A surface layer of approximately 50 was then removed from the samples by chemical polishing. The samples&apos; surfaces were then protected by a chromating treatment such that in a subsequent etch only grain boundary regions were dissolved. The dissolved regions were approximately semicircular troughs, 10 µ in width. The grain boundaries were usually in the middle, i.e., at the bottom, of this trough. For zinc containing the K < 1 solutes cadmium and aluminum, the grain boundary was always indicated by a ridge as shown in Fig. 1. For pure zinc, and zinc containing the K > 1 solutes copper and gold, a groove at the bottom of the trough indicated the boundary position, Fig. 2., To test if these different grain boundary etching behaviors were possibly connected with the segregation process discussed above, the Zn-Cd alloy was again heated to 380°C and quenched into brine. After surface protection and etching it was found that this specimen no longer exhibited the grain boundary ridges but rather grain boundary grooves as shown in Fig. 2. It would appear, therefore, that the boundary ridge is a consequence of a process occurring during the slow cooling of the samples. From absorption spectrophotometric analysis it was found that for slowly cooled samples containing K < 1 solutes, the solute/zinc ratios of the material from the boundary trough were consistently considerably higher than the solute/zinc ratios of the bulk alloys. This boundary region solute enrichment was not observed in quenched samples. It is concluded therefore, that the boundary ridge shown in Fig. 1 is directly or indirectly connected with nonequilibrium solute segregation.

Jan 1, 1970

By L. R. Velho, N. A. D. Parlee, J. Gani

THE unsteady-state type of gas-liquid metal diffusion cell1 has found considerable application.2"4 Cells of this general design can function as steady-state cells if allowed to operate long enough to reach steady state. This report tells of some success in the adaption of this type of cell for steady-state work, using the 0-Ag system in which the mass transfer of the gas "species" is basically controlled by diffusion in the liquid. The gas, e.g-., O2, is maintained in the cell tube&apos;-3 at 1 atm pressure, Fig. 1, by buret and sensitive oil manometer. The gas, after diffusing through the stagnant metal, escapes at the annular metal surface around the tube at a lower constant partial pressure, in this case that of O2 in slowly flowing air, 1 atm.

Jan 1, 1970

By Lucien Eaton

THE accuracy with which detachable bits can be sharpened and gauged by grinding has made possible refinements in design that cannot be attained by forging. This applies particularly to gauging, where accuracy is of the greatest importance. Gauge changes of 1/32 in., and even 1/16 in., are now feasible, whereas 1/16 -in. changes are about the limit in accuracy for forged bits. These small changes in gauge are feasible, however, only with certain fundamental principles of design. Within reasonable limits the depth to which a hole can be drilled is dependent primarily upon the loss in gauge, or diameter, of the bit. In drilling there is always a certain amount of " overdrilling "-the bouncing of the bit in the hole and its rotation wear away the walls-so that the actual diameter of the hole is greater than the diameter of the bit. The amount of this overdrilling depends upon the kind of rock being drilled, and is sometimes sufficient to permit the use of the same diameter of bit for the full depth of the hole. If, for instance, a 1 ¾ -in. diameter bit overdrills 1/32 in. in diameter, and loses 1/64 in. in gauge in drilling 2 ft. of hole, the bottom of the hole is still 1/64 in. larger than the original diameter of the bit, and a new 1 3/4-in. bit can be used to drill the next 2 ft. If, however, the overdrill is less than the loss in gauge, the next bit must be of smaller diameter. The first requisite in reducing loss of gauge in drilling is to have the points and corners of the wings of the bit cut circles of the same diameter, thus producing a reaming surface with continuous bearing on the wall of the hole. The greater the reaming surface, the less the gauge loss that is to be expected. If the reaming surface, however, is too large, it may impede the rotation of the bit through excessive friction. This principle was probably first recognized in the design of the Carr bit (Fig. 1)) a bit with a single cutting edge of wide angle, a definite reaming edge, and a large center hole. The same idea was followed in designing the double-taper cross bits, which are now standard almost everywhere Fig. 2 illustrates a typical double-taper cross bit. The taper from the point of the wing A to a little behind the corner B is 5° and from

Jan 1, 1939

Receipts General Funds: Initiation fees &apos; 6,008.50 Arrears of dues 1,772.69 Current dues 46,576.56 Advance dues 1,820.97 Sale of binding 10,554.43 Sale of advertising 6,003.42 Sale of Transactions 3,237.07 Sale of special editions &apos; 1,722.50 Sale of Bulletins and. pamphlets &apos; 2,263.95 Miscellaneous Receipts; Interest on investments and deposits 450.06 Sale of pins and fobs 312.80 For the collection of checks 161.84 From Finance Committee, Feb., 1916, Meeting 101.02. Subscriptions to the land fund 86.14 Arizona Meeting refund from N. Y. Central Railroad.. 52.48 Local Section refund from N. Y. Central Railroad.... 31.85 Miscellaneous 425.68 Total Receipts General Funds 81,611.96 Cash balance Dec. 31, 1913 1,063.39 $82,675.35

Jan 2, 1917

By Robert Mehl

THE system cadmium-mercury has attracted much attention, chiefly because of its importance in the study of electrochemical standard cells. The constitutional diagram, as it is accepted today, is essentially that FIG. 1.-BIJL&apos;S DIAGRAM FOR CADMIUM-MERCURY ALLOYS. developed by Bijl1 (Fig. 1), though there has been a great amount of work done on the system by other workers.2 A year ago the senior author published a paper on the crystal structure of the system, in which were described determinations of crystal structure

Jan 1, 1929

By C. A. Stickels

A heat of electrolytic iron, to whzch alunzinutn and nitrogen had been added, was hot-rolled, cold-rolled 90 pct, and recrystallized at temperatures from 500" to 700°C. Primary recrystallization textures appear to arise from competitive growth of two types of nuclei: 1) those having orientations belonging to the "usual" primary recrystallization texture found in riming steel, and 2) those with the {111} (110) ovientation. Development of a (111}(1 10) component in the primary recrystallization texture occurs only over a certain interval of isothermal recrystallizatzon temperatures when the material is supersaturated with respect to the precipitation of AlN. Lowering the degree of supersaturation depresses the temperature interval in which a (111)(110) component occurs. An elongated, &apos;pancake-shaped" recrystallized pain structure and a marked delay in the start of recrystallization were found in all specimens which were supersaturated with respect to A1N precipitation after cold work, regardless of their recrystallization texture. ONE of the consequences of killing low-carbon steel with aluminum is a significant change in recrystallization behavior. About 15 years ago, Solter and eatttiel showed that this behavior was largely controlled by aluminum and nitrogen in the steel. If complete precipitation of A1N was prevented before cold rolling, an increased "recrystallization temperature" was observed in subsequent. annealing, and the recrystal-lized grains were not equiaxed. Leslie et a1.2 studied this phenomenon in some detail and clearly demonstrated the relationship between A1N precipitation, recrystallization kinetics, and the development of "pancake-shaped" grains. It has also been known for some time that aluminum-killed steels, processed to produce elongated "pancake" grains, develop a (11 I}( 110) primary recrystallization texture. This texture has not been found in iron or low-carbon rimming steel as a primary texture4j5 but has been observed following grain growth in electrolytic iron.5 The present work was undertaken to study in more detail the effect of A1N supersaturation on recrystallization textures in iron. LITERATURE REVIEW The deformation texture in heavily rolled iron has been studied in detail by Bennewitz.~ The texture consists primarily of a partial fiber texture about a (110) axis in the rolling direction, designated here as fiber texture A. It includes the range of orienta- tions (111)[110] - (001)[ 110] - (11l)[110]. A weak secondary texture also is present.6 This is a duplex partial fiber texture about two (110) fiber axes located 60 deg from the rolling direction and 30 deg from the sheet normal. The range of this texture, designated here as fiber texture B, about the [101} fiber axis is (112)[110] - near (545)[252] - (211:1[011] *The range given here follows Bennewit~.~ A few pole figures from re-crystallized material indicate a broader range than this.&apos; However, the components which are strongest in the recrystallization texture are in this range.&apos;________________________________________________________ Primary recrystallization textures in unkilled steels can be accounted for by growth of members of fiber texture B present in the deformed metal.5 However, while members of fiber texture B dominate the primary texture, other orientations survive primary recrystallization as well. In particular, some {111}(110) members of fiber texture A must also grow during primary recrystallization, because a well-defined {1ll)( 110) texture develops during subsequent grain growth at 700°C.5 The unusual recrystallization behavior of deformed supersaturated solid solutions has been attributed to: 1) retention of the solute in solution,&apos; 2) formation of coherent, preprecipitation solute clusters prior to and during re~r~stallization,~ and 3) formation of a precipitate prior to and concurrent with recrystallization.&apos;~-&apos;~ When aluminum is supersaturated with iron, the difference in grain boundary mobility between general high-angle boundaries and certain special coincidence site boundaries is apparently eliminated.&apos; In aluminum-killed steels, precipitation of A1N can take place at ordinary subcritical recrystallization temperatures. The rate of precipitation increases with increasing aluminum or nitrogen contents.2&apos;13 There is some doubt, however, as to whether true precipitates form during the time at temperature needed to complete recrystallization. Leslie ef a1.2 found that precipitation in one steel was complete after about 100 min at 700GC, or after about 1000 min at 650GC, as measured by chemical analysis for AlN. Aoki et a1.,13 using internal friction for dissolved nitrogen, showed that a large fraction of the dissolved nitrogen was removed from solution within a few minutes annealing time at temperatures from 400" to 800°C. However , the rate of formation of AlN, as detected bv chemical analvsis. was much slower than the apparent rate of nitrogen removal. Hasebe,&apos;~~ using carbon extraction replicas, has identified A1N precipitates by electron diffraction in a 0.2 C steel, solution-treated at 1300°C and annealed 2 hr at 700°C. Borchers and kim,I6 also using a replication technique, observed precipitates after annealing treatments as short as 2 min at 640°C. However, Leslie et a1.&apos; state that no A1N precipitate can be seen while recrystallization is being inhibited in aluminum-killed steel.

Jan 1, 1967

By R. P. Carreker, R. L. Fullman, J. C. Fisher

CREEP tests are usually constant load tests. As deformation occurs, the cross section of the specimen changes so as to maintain approximately constant volume. Until necking starts, A _ l0 A0 - 1 where A is the cross-sectional area, 1 is the specimen length, and the subscript o indicates initial values. Thus, during a tensile creep test under constant load, the stress increases in proportion to the elongation. A A, 1, where S is the stress and F the load applied to the specimen.

Jan 1, 1954

By M. F. Trice

PYROPHYLLITE is a hydrous aluminum silicate (Al2Si4O10(OH)2)1 that occurs in both the foliated and the massive forms. The foliated variety resembles talc in that it has a greasy feel, a pearly luster, perfect basal cleavage, and usually is white, although in some deposits apple green, gray, brown, russet, and nearly black specimens may be found. The massive deposits of pyrophyllite yield pearl gray to light tan aggregates, some of which are so friable that they may be crushed with the hands into a fine gritty powder. Frequently specimens of the massive variety are studded with crystals of radial structure. The mineral has a specific gravity of about 2.7 and a hardness of 1 on Mohs&apos; scale.1 The chemical composition of commercial pyrophyllite reflects the mineralogy of the deposits, which, according to Stuckey,2 contain pyro-phyllite and quartz, together with chloritoid, sericite, and a few other minerals of negligible importance. The latter, he says, "are noticed in but small quantities to the extent they might occur as accessory constitu-ents of an igneous rock or as products of regional metamorphism or weathering. . . . Quartz is abundant everywhere except in the very best grades of pyrophyllite." Analyses of specimens from several deposits are listed in Table 1, showing that the silica content varies from 57 to 73 per cent and the alumina from 22 to 33 per cent. Oxides of iron and the alkaline metals are collectively present, with one exception, to an extent of about 1 to 1.5 per cent.

Jan 1, 1940