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By R. S. Dean

THE first attempt to establish an equilibrium diagram of the lead-antimony series was made by Roland-Gosselin1 in 1896. This investigation classified the system as a purely eutectiferous one, with the eutectic at 12.5 per cent. antimony and 228° C. In the next year, Stead2 placed the eutectic temperature at 247° C. The error in temperature made by Roland-Gosselin, however, has been perpetuated in the literature with a remarkable degree of persistency; this value was accepted by Campbell (1902), and Ewen (1910),3 and will be found in the last edition of Desch's3 Metallography (1922). Gonterman's investigation, made in 1907, was the most complete one available at the time the present investigation was started. He concluded' that the system was a simple eutectiferous one and placed the eutectic at 13 per cent. antimony and 247°C. Other investigations by Tammann (1907),5 Garbow (1909),6 Rudolfi (1910),7 Loebe (1911),8 and Leroux (1913)9 added little, if any-thing, to Gonterman's diagram. A number of investigations have been made of physical properties of the alloys. The electrical conductivity of the series was studied by Matthiessen (1860)10 and by A. W. Smith (1921);" the thermoforce by Rudolfi (1910),12 and Pelabon (1923);13 the specific heat by Durrer

Jan 2, 1926

By Henry A. Stiff, J. P. Hammond, A. B. Westerman, H. C. 195-000-000-014 Cross, and Lawrence E. Davis

The phases in Cr-Fe-Mo alloys have been investigated with homo-genization, aging temperature, composition range, and alloy addition as variables. Metallography, three X-ray methods, and hardness were used as methods of study. The behavior of s, M23,C8, and Z phase are reported for cornpasition range 60 pct Cr-15 to 25 pct Fe-15 to 25 pct Mo-<0.005 to 0.36 pct C with aging at 1400º to 2000°F. With 2 pct Ti, Tic and TiC-TiC are formed; with nitrogen Cr2N. THE recently developed class of chromium-base heat-resistant alloys shows appreciably higher strength than commercially used high temperature alloys. However, further developmental work is required to impart needed shock resistance and some degree of room-temperature ductility to these alloys. Extensive exploratory work on chromium-base alloys was begun in a program initiated by the War Metallurgy Committee of the National Defense Research Council at the Climax Molybdenum Co. in the early part of 1942.&apos; Alloys were sought for gas-turbine blades for use at 1600°F. A minimum of 5 pct elongation in the stress-rupture test was a requirement. From the Climax work, ternary alloys of chromium, iron, and molybdenum appeared to show the greatest promise as materials for gas-turbine blades. The composition line in the ternary diagram joining the 60 pct Cr-15 pct Fe-25 pct Mo and 60 pct Cr-25 pct Fe-15 pct Mo alloys was indicated as representing the most useful combination of strength and ductility.&apos; Strength increased while ductility decreased as molybdenum was raised in this composition range.&apos; The 60 pct Cr-15 pct Fe-25 pct Mo type alloy was thought to have the most suitable properties for gas-turbine blades.&apos; Concurrent with the study reported here, other investigations were being made on Cr-Fe-Mo alloys. The liquidus temperatures on a series of low carbon, ternary alloys were being determined, and isothermal sections drawn at 2370°, 2010°, and 1650°F (1300°, 1100°, and 900°C).2 Also, various methods of preparation and some mechanical and physical properties of chromium-base alloys, particularly the 60 pct Cr-15 pct Fe-25 pct Mo type, were being investigated., At the inception of the present program, only a limited study had been made of etchants for developing the microstructures of chromium-base alloys; X-ray analysis of the microconstituents had not been made. A review of the literature revealed that no phase-diagram work had been reported on the Cr-Fe-Mo system. Scope of Work The work on chromium-base alloys includes the following: 1—The development of etching methods for differentiating between the microconstituents; 2—The identification of microconstituents by X-ray diffraction methods; and 3—A comprehensive metallographic and hardness study after various heat treatments. The phases were studied by three standard X-ray techniques: 1—The block-sample focusing-camera method; 2—The powder-diffraction method on elec-trolytically separated residues; and 3—The powder-diffraction method on cold-worked aggregate samples. The results by the first two methods were correlated with the metallographic data. The third method was used largely to study conditions approaching equilibrium, since cold working prior to heat treating accelerated precipitation. Seven types of alloys were investigated: 50 pct Cr-50 pct Fe, 40 pct Cr-40 pct Fe-20 pct Mo, 55 pct

Jan 1, 1953

By Allan Martin, R. A. Swalin

Diffusion rates of manganese, aluminum, titanium, and tungsten in nickel were measured at temperatures between 1100° and 1300°C. Activation energies, Q, and values of the frequency factor, Do, were calculated for each system. Values of Q were found to be within 15 pet of that for self-diffusion of nickel in agreement with present theories of diffusion. ONE of the most interesting problems in diffusion concerns factors governing the order of diffusion coefficients of various solutes in a common solvent metal at low solute concentrations. An early suggestion by Rhines and Mehl&apos; was that diffusion coefficients might each approach that of the self-diffusion coefficient of the solvent itself. Later Thomas and Birchenall&apos; pointed out correlations between the order of the diffusion coefficients and the freezing point depressions of the solvent by the various solutes. Other correlations have been sought. To settle this question it seems likely that much more data will be needed and, in particular, data collected at low solute concentrations. This type can be obtained by special analytical techniques as well as by use of radioactive tracers. The following study was made of diffusion of manganese, aluminum, titanium, and tungsten in a common solvent, nickel, at relatively low solute concentrations: 1—Production of alloys of nickel with manganese, aluminum, titanium. and tungsten at concentrations of the order of 1 atomic pet. 2—Use of pressure-welded diffusion couples in cylindrical form. 3—Diffusion anneals in the temperature range of 1100" to 1300°C. 4—Lathe sectioning of the diffused couples. 5—Spectrophotometric analyses of lathe turnings for the solute being studied. 6—Calculation of the diffusion coefficient for each of the solutes as a function of temperature. 7—Interpretation of results with current theories of diffusion. Experimental Work Materials—Two types of nickel were used in this research. The grade designated as Vacuum melted Mond nickel was used for all diffusion systems except the Mn-Ni system for which the Vacuum melted electrolytic nickel was used. Analyses of the two grades are given in Table I. The manganese used in making alloys was 99.9 + pet pure, aluminum 99.99 pet pure, titanium 99.9 pet pure, and tungsten of C.P. grade powder. Alloys used in this investigation were prepared by melting in vacuo or under a low pressure of argon. After melting, all heats except those of nickel-manganese were poured into a cylindrical copper mold 3 in. long and 1 in. ID. In the case of nickel-manganese, the alloys were solidified in the melting crucible. For nickel-aluminum heats alumina crucibles were used and for all other heats, zirconia crucibles.

Jan 1, 1957

By Cyril Stanley Smith, Earl W. Palmer

A complete discussion of the literature on the subject of the influence of copper on iron and steel will be published elsewhere.&apos; The present paper is concerned especially with the precipitation-hardening of copper steels, and brief mention will be made here only of the work of Kin-near,2,3 Nehl4 and Buchholtz and Koster,5 who alone have studied this phase of the subject. Kinnear found that normalized low-carbon steel castings containing about 1.0 per cent copper on drawing at temperatures in the neighborhood of 530" C. (1000" F.) increased in tensile strength and yield point, with little decrease in ductility. The maximum effect was obtained with about 0.9 to 1.0 per cent copper, and when the copper exceeded 3 per cent no improvement occurred on reheating. Kinnear did not study the effect of time of reheating—an all important factor where precipitation is concerned. Neh1,6 who apparently was unaware of Kinnear&apos;s work, recognized the change as due to precipitation and studied the interdependence of time and temperature of reheating, the temperature of the initial solution heat treatment, and the speed of cooling necessary to retain the bulk of the copper in supersaturated solution and to give an increase in hardness and tensile strength on reheating. Nehl found that if the final rolling or forging temperatures exceeded 700" C. (1300" F.) it was not necessary even to normalize the steel, but full precipitation occurred on merely reheating the forged or rolled shapes at 500" C. (930" F.). Buchholtz and Koster7 studied in detail the precipitation-hardening of a mild steel containing 1.0 per cent copper and 0.5 per cent chromium. From data on the electrical and magnetic properties of copper steels, they drew the portion of the equilibrium diagram in Fig. 1, assuming the

Jan 1, 1933

Jan 1, 1966

By G. L. Oldright

THE development of the reverberatory furnace for smelting copper ores up to 1912 was described by E. P. Mathewson1 with details concerning the great changes in dimensions of the furnace. Hayward2 tabulated similar data, giving developments up to about 1924, in his revision of Hofman&apos;s Metallurgy of Copper. DEVELOPMENT OF PRESENT SMELTING PRACTICE Dimensions of Furnace Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may be considered that the improvements made were due largely to increases in the size of the furnace used. Doubt-less some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By S. F. Shaw

The Seminole field was first drilled in 1913. During the next 10 years other attempts were made to discover oil in this field, but without encouragement until March, 1926, when the Indian Territory Illuminating Oil Co. completed NO. 1 Dyer in the Hunton lime for 1100 bbl. On July 16, 1926, the Independent-Garland interests discovered oil in the Wilcox formation when No. 1 T. Fixico came in at 4055 ft. and on Sept. 9, 1926, production reached a peak of 8040 bbl. On Aug. 1, 1927, the daily production of the pools making up the Seminole district amounted to slightly over 500,000 bbl., practically all of which was coming from the Wilcox formation. The Hunton formation has not been very productive of oil, but has maintained a fairly high gas pressure; this formation has been encountered at from 3800 to 4200 ft., below which there is about 100 ft. of Sylvan shale. The shale is underlain by 75 ft. of Viola limestone in a few places in which formation oil has been encountered. Below the Viola lies the upper Simpson formation in which is found the Wilcox sand. The depth of the Wilcox in the Seminole field varies from 3900 to 4500 ft., and initial production has reached as much as 12,000 bbl. The gravity of the oil is about 41° Be. The general practice is to set 6 5/8 -in. casing above the Wilcox and drill in, though in several cases 5 3/16-in. casing has been set for the oil string, and in a few instances, casing as small as 4 3/4 in. has been employed. Production Statistics Production for the Seminole district for the various months since its discovery is shown in Table 1.

Jan 1, 1928

By W. W. Stephen, J. P. Walsted, W. J. Kroll

FOR the production of carbides, various furnace types are available, especially those using arc, resistance, and high-frequency heating. Selection of a specific means of heating depends primarily on the material to be treated and the physical properties of the carbide produced. In the present case, zirconium carbide had to be prepared on an industrial scale as a raw material for the production of anhydrous zirconium chloride. Considering that a rather expensive pure oxide was to be used, the arc-furnace treatment recommended for zircon sands in a previous publication&apos; was ruled out because of the considerable volatilization and dust losses caused by the blast of the arc. For this reason, either high-frequency or resistance heating seemed to offer more promise. Since there was not enough capacity of the former available, resistance heating was chosen. It was first thought that the Acheson silicon carbide furnace would be suitable for the present purpose, but the voltage in such a furnace, in which the current passes through the batch, varies from 220 to 75 v from the start to the end of a run. This variation is so great that a special tap transformer would have been required. Trouble was also expected by local melting of the carbide. Pure zirconium carbide melts at about 3527°C, but the melting point is brought down to 2427°C, according to Agte,2 when an excess of 6 pct C is present in the carbide. This we found confirmed by experiments in a high-frequency furnace. Excess carbon is needed in the batch to obtain a complete reduction. Fusion of the charge would cause great difficulties in an Acheson-type furnace because of the good electrical conductivity of the carbide as compared with that of the loose batch. Also, fused carbide is much more difficult to chlorinate than the spongy product that can be made in the radiation furnace described below. It was apparent that, to obtain a good-qual-ity zirconium carbide, the heat input would have to be well-controlled. The hairpin-resistor principle seemed to offer possibilities in this regard, and a furnace of this type was therefore developed. The advantages of the hairpin-resistor radiation principle have been discussed in previous publications, and a split-tube graphite-resistor furnace," now increasingly used in various laboratories, as well as a centrifugal quartz melting furnace4 of this type, has demonstrated the usefulness of this heating method. The hairpin-heater element has the following definite advantages over a straight resistor of the type used, for instance, by Georges:5 Its resistance is four times greater; it can expand freely; it is sturdier because of the larger diameter, and it has a larger radiation surface; there are no hot contacts that might wear out or overheat; only one clamp is used which permits assembling all electrical leads at one side of the furnace, making the other sides easily accessible to the operator. The shorter element and its larger diameter permit greater concentration of heat. The furnace developed is shown in fig. 1. The box (I), made of 2 1/2-in. graphite plates, has inside dimensions of 23x17x16 in. It contains the briquet-ted batch (2). The box is embedded in lampblack (3) up to the cover plate. The cover plate contains an opening for the gas escape (4) and for the observation hole (5), which permits measuring the temperature with an optical pyrometer. The cover plate is embedded in charcoal (6). The lampblack is contained in the insulating brick lining (7), held in the 1/4-in. sheet steel box (8). The graphite box is set on two rows of triangular graphite bars (9). The hairpin-heater element (10), the dimensions of which are given below the main drawing, extends horizontally in the graphite chamber and radiates freely on the batch. A graphite tube (11) keeps the lampblack from falling into the slot. The split electrode, which in reality is turned 90" against the drawing, is so arranged that the slot is vertical. The water-cooled packing gland (12) is insulated by an airgap from the heater element. A thin pipe (13)

Jan 1, 1951

Discussion of the paper of BENJAMIN F. TILLSON, presented at the NEW York meeting, February, 1915, and printed in Bulletin No. 9S, February, 1915, pp. 505 to 528. T. E. STURTEVANT, * New York, N. Y.-Mr. Tillson&apos;s paper on hammer drills is of much interest to me. Early in 1909 I was invited to make tests of our drills at the Franklin mine of the New Jersey Zinc Co. The first drill that we tested gave very poor results. This was a drill having a hammer piston 1¾ in. in diameter and 3 ¼ -in. stroke, the hammer piston weighing about 6 1b., the 1 ¾ in. diameter being exposed to pressure at either end. Although this drill has given excellent results as a sinking drill, its performance as a stoping drill was poor. This was one of the surprises of the limestone as spoken of by Mr. Tillson. This led me to further investigation, and a second drill was developed, with a much lighter piston and a somewhat longer stroke, that gave better results. In the design of this drill, no attention was paid to the cost of construction, our whole attention being given to the increase in drilling speed and reduction of the quantity of air used. The first drill of this type had a piston weighing about 2 ½ lb. and a stroke of 3 ¼ in. The piston was made with a hammer-bar extension, and the cylinder was made with two diameters of the bore, one in which the hammer-bar extension moved and the larger one in which the piston head moved. The air was taken in at either end of the larger bore, thus giving a greater area to the

Jan 5, 1915

By David B. Reger

The expansion of the previously discovered gas pools and an intensified search for additional pools in the Oriskany and deeper sands were the principal petroleum activities in West Virginia during 1941. One small oil pool was discovered and another was considerably extended. One small gas pool was discovered, several isolated well discoveries were expanded into definite pools, and various recognized pools were substantially increased. The proved oil territory of the state was increased about 600 acres and the proved gas territory about 45,000 acres. The new discoveries of oil did not equal withdrawals but gas reserves added were greater than gas withdrawals. Oil activity, as in the past several years, was slight because of unfavorable crude prices. Most operators prefer to leave oil in the ground rather than to drill for it without a better return. Gas activity increased greatly because of better demand, even though the price of gas did not improve. The account of operations, as gathered from trade journals and other reporting services, shows that 838 new wells were drilled, resulting in 114 new oil wells with 2606 bbl. of daily new production; 558 new gas wells with 949,226,000 cu. ft. of daily open flow; and 166 dry holes. Also, 86 old wells were drilled to deeper sands with 9 bbl. and 56,030,000 cu. ft. of added production. On the new wells the oil average was 22.86 bbl. per well per day; and the gas average was 1,701,000 cu. ft. per well per day. On new wells the ratio of dry holes to completions was 13.61 per cent. On deeper drilling to other sands the ratio of failures was 15.12 per cent. From an exploratory standpoint one strictly new oil pool (Higginbotham Run) was opened in Kanawha County and one strictly new gas pool (Lost Run) in Taylor County. One entirely isolated wildcat gave commercial Oriskany-sand gas production and may open a pool. Various other wildcats greatly extended the boundaries of known Oriskany-sand gas pools and various others may have opened new pools in the shallow sands. On the other hand, various Oriskany tests in Harrison, Monongalia, Roane and Wood Counties were dry and all five tests to the White Clinton sand (Sil.) in Boone, Harrison, Kanawha and Wood Counties failed to find commercial production. The principal oil wells completed, by counties, were: Calhoun, 24; Clay, 7; Gilmer, 6; Hancock, 9; Kanawha, 9; Pleasants, 10; and Ritchie, 14. Production for the year was estimated by the Oil Weekly as 3,421,000 bbl., compared with U. S. Bureau of Mines final figures of 3,444,000 bbl. in 1940. The leading counties in gas-well completions were: Boone, 32; Braxton, 19; Cabell, 16; Calhoun, 46; Clay, 20; Gilmer, 37; Jackson, 94; Kanawha, 82; Lincoln, 17; Putnam, 46; Ritchie, 33; Upshur, 20; and Wayne, 33. Production for the year is estimated by the author as 210,000,000,000 cu. ft., as compared to the U. S. Bureau of Mines&apos; final figures of 188,751,000,000 cu. ft.* in 1940.

Jan 1, 1942

By David B. Reger

The expansion of the previously discovered gas pools and an intensified search for additional pools in the Oriskany and deeper sands were the principal petroleum activities in West Virginia during 1941. One small oil pool was discovered and another was considerably extended. One small gas pool was discovered, several isolated well discoveries were expanded into definite pools, and various recognized pools were substantially increased. The proved oil territory of the state was increased about 600 acres and the proved gas territory about 45,000 acres. The new discoveries of oil did not equal withdrawals but gas reserves added were greater than gas withdrawals. Oil activity, as in the past several years, was slight because of unfavorable crude prices. Most operators prefer to leave oil in the ground rather than to drill for it without a better return. Gas activity increased greatly because of better demand, even though the price of gas did not improve. The account of operations, as gathered from trade journals and other reporting services, shows that 838 new wells were drilled, resulting in 114 new oil wells with 2606 bbl. of daily new production; 558 new gas wells with 949,226,000 cu. ft. of daily open flow; and 166 dry holes. Also, 86 old wells were drilled to deeper sands with 9 bbl. and 56,030,000 cu. ft. of added production. On the new wells the oil average was 22.86 bbl. per well per day; and the gas average was 1,701,000 cu. ft. per well per day. On new wells the ratio of dry holes to completions was 13.61 per cent. On deeper drilling to other sands the ratio of failures was 15.12 per cent. From an exploratory standpoint one strictly new oil pool (Higginbotham Run) was opened in Kanawha County and one strictly new gas pool (Lost Run) in Taylor County. One entirely isolated wildcat gave commercial Oriskany-sand gas production and may open a pool. Various other wildcats greatly extended the boundaries of known Oriskany-sand gas pools and various others may have opened new pools in the shallow sands. On the other hand, various Oriskany tests in Harrison, Monongalia, Roane and Wood Counties were dry and all five tests to the White Clinton sand (Sil.) in Boone, Harrison, Kanawha and Wood Counties failed to find commercial production. The principal oil wells completed, by counties, were: Calhoun, 24; Clay, 7; Gilmer, 6; Hancock, 9; Kanawha, 9; Pleasants, 10; and Ritchie, 14. Production for the year was estimated by the Oil Weekly as 3,421,000 bbl., compared with U. S. Bureau of Mines final figures of 3,444,000 bbl. in 1940. The leading counties in gas-well completions were: Boone, 32; Braxton, 19; Cabell, 16; Calhoun, 46; Clay, 20; Gilmer, 37; Jackson, 94; Kanawha, 82; Lincoln, 17; Putnam, 46; Ritchie, 33; Upshur, 20; and Wayne, 33. Production for the year is estimated by the author as 210,000,000,000 cu. ft., as compared to the U. S. Bureau of Mines&apos; final figures of 188,751,000,000 cu. ft.* in 1940.

Jan 1, 1942

By J. P. Zannaras

If we assume that testing of the model started at time t1, that time t2 was the instant at which the elastic limit of the material was passed at the points of the maximum stress, and that at time t3 the failure of the model was completed as shown in the typical photograph (Fig. I), then actually the photograph shows the final results of what took place between time t2 and t3. By the principle of similitude all conclusions drawn from the behavior of the model were applicable and valid for actual mine roofs up to the time t2 after time t2 motions and friction came into play during the destruction of the model. The events between t2 and t3 and the photograph itself are not representative of the events expected in an actual mine roof unless corrections imposed by the principle of similitude are made. It can therefore be stated that all conclusions and observations of the author from the photographs are applicable only to his small-scale experiments and not to actual mine roofs. Without investigation of the true stresses or the combined maximum stress induced in the beds it has been assumed by the author that failure occurred solely due to bending stress caused by loading of the beds. (The author states that D was taken to be the theoretical strain given by elementary beam theory.) Fig. 1 shows that the beds were clamped together. This clamping was necessary to prevent slipping of the beds due to the horizontal shear. However, this clamping caused an uncontrolled and undertermined com-pressive stress S at the end of the beds, and this com-pressive stress caused a true tensile stress ?S (? PoiS-son ratio). This true tensile stress combined with the tensile stress produced at the end of the beds due to loading (fixed beam uniformly loaded), and the horizontal shear which is maximum at the end appears to be the most probable maximum stress produced in the beds. It is therefore evident that the author in performing this experiment has introduced stresses in the model not existing or dissimilar to actual mine roofs, and therefore his conclusions may be applicable to his small-scale experiments but inapplicable to actual mine roofs. Louis A. Panek (author&apos;s reply)—The discusser&apos;s difficulty is related to mechanics of materials rather than similitude. The effect of time did not enter into the tests in any way, being simply excluded from consideration in this study, because all quantitative results entering into the design equations are based on measured (between time t1 and time t2) bending strains that are well below the breaking strain (i.e., in the elastic range) for the model material. Time does not become a significant factor until the rock stress reaches 80 to 90 pct of the breaking stress." The objective of this investigation was not to predict that bolted roof will fail after time t, but was instead to develop a safe roof, that is, roof in which the bending strain is much less than (say less than 50 pct of) the breaking strain. The models tested to failure had no direct bearing on the design equations. Although for some rocks time may influence the magnitude of stress or strain at which fracture occurs, breaking stress or breaking strain values were not employed for any purpose in this investigation. Tests to failure were made only to obtain additional evidence regarding the behavior of the bolted laminae&apos; as evinced by the locations of the cracks. As stated in the last paragraph of the paper, if one wishes to predict the time or stress conditions for which the roof will fail, additional information is required. The reason for clamping the laminae is to make them behave liked clamped beams, which closely approximate the clamped-plate condition of the actual mine roof beds (Ref. 5, p. 1; Ref. 6). The writer has previously demonstrated that the bending stresses induced by centrifugal loading in a clamped model beam are in agreement with those predicted by theory;&apos; hence clamping has no significant effect on the bending ,trains measured during a test. Moreover, actual ,in, roof beds are likewise subject to a clamping effect due to weight of superincumbent strata carried by the pillars or ribs on each side of the opening. It is emphasized that this study was restricted to evaluating only the reinforcing effect produced by the bolts. Procedures employed in the experiments and in the data analysis were such as to deliberately exclude the effects of other factors. The objections offered are therefore invalid,

Jan 1, 1957

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A 1/2x4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four 5/s in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of 3/s yard capacity, both manufactured by Blaw Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1950

By Henry N. McCarl

Mineral aggregates are those natural and manufactured industrial mineral and rock materials that provide bulk and strength in port- land cement concrete, bituminous concrete mixes, and plaster or stucco surfaces. They may also provide special characteristics such as thermal and acoustical insulation, weight, surface textures, abrasion resistance, and impermeability to various concrete products and mixes. Most aggregates are relatively inexpensive materials, but despite their low unit value, the tonnages consumed each year in the United States and throughout the world make mineral aggregates by far the leading industrial mineral and rock materials both in terms of volume (tonnage) produced and value consumed. The major use of mineral aggregates is in concrete. The low cost, high bulk aggregates tend to keep concrete construction costs at levels competitive with other building materials. Other major uses of aggregates include highway and railroad base or ballast materials, graded fill, and various industrial uses such as metallurgical and chemical fluxes and raw materials-these latter uses are often considered nonaggregate uses of mineral aggre- gate materials. The value of mineral aggregate materials at the production plant varies from less than a dollar per ton to several dollars per ton for most crushed stone and sand and gravel, $2 to $15 per ton for slag and expanded clay or shale aggregate, to as much as $60 to $100 per ton for expanded perlite or vermicu- lite. The more specialized the use, the higher the price-although there are naturally some exceptions to this rule of thumb. The first four parts of "Construction Materials" deal with the utilization of four major mineral aggregate types: Crushed Stone, Light- weight Aggregates, Sand and Gravel, and Slag. While the focus will be on the use of these industrial minerals and rocks there will also be some coverage of sources, mining and production, especially in those cases where they are

Jan 1, 1975

By Alfonso W. Ovalle, Héctor R. Albornoz

INTRODUCTION The El Teniente Mine of Codelco-Chi le has extracted 588 mil lion tons of ore in its 77 year history for a total copper output of 8.800.000 tons. The mine production has been obtained mainly by traditional block-caving where grizzlemen using a hand held lever send ore through long-raises to load trains. At present the mine is about to introduce mechanized block-caving using LHD equipment to extract the primary rock. This paper describes this method which will be tested starting next month by caving an 11.000 square meter block. THE PRIMARY ROCK PROBLEM The secondary enrichment sulfides of the El Teniente ore body have characteristics which makes a standard unmechanized block-caving practice very successful. In the 60&apos;s parts of blocks encountered primary ow in the Teniente 1 and Sub-13 levels which proved to be extremely hard to extract. These foretold the obstacles that arised in the central part of the Teniente 4 level. Preparation of a completely primary 8.100 square meter block started in 1970. Its undercutting was completed by the end of 1972 and by 1973 the problem of mining the primary rock was clearly visualized. The numbers below describe the difficulties in mining this block: -Long-hole drilling rate is 90 m (270 ft) per shift in secondary rock and 37 m (110 ft) in primary. - Undercutting which is completed in 9 shifts in secondary took 180 shifts in the primary block. - Average grizzlymen productivity in secondary is 250 tons per shift while in primary it is only 25-30 tons, both at 343 mm (13 1/2 in. ) grizzly opening. - Powder consumption for secondary blasting increased from 8 grams per ton in secondary to 84 in primary. - The protection pillar between the under- cutting and the production levels was 60% broken in primary while 20% is usual for secondary. - Extraction advanced 5 cm per day in primary while a 50 cm per day rate is the aver- age in secondary. This is the main reason that after 8 years this primary block has a 20% extraction. Secondary blocks take 1.5 to 3 years to complete their extraction. - 80% of the boulders in primary rock are larger than 343 m (13 1/2 in.) with 10-20 % of them larger than 3 meters. -Direct cost of production (preparation, undercutting, extraction and haulage), using the traditional block-caving design for secondary rock, increased 3.5 to 4 times when used in primary rock. The following indicators of Table 1 show the rock characteristics differences.

Jan 1, 1981

By AIME AIME

THE fall meeting at El Paso this year (Oct. 13¬15) will be of unusual interest due to the international atmosphere imparted by the many engineers from Mexico, who are making arrangements to attend this meeting. The meeting will be held jointly with the Western Division of the American Mining Congress. Cooperating in the program will be the representative society of Mexican engineers known as Centro National de Ingenieros of Mexico City, who are plan¬ning on sending many delegates and well known Mexican engineers to the convention. The cooperation of the West Texas Geological Society and the American Association of Petroleum Geologists has also been secured. Brent N. Rickard, manager of the El Paso Smelting Works of the American Smelting & Refining Co. at El Paso, Tex., is general chairman in charge of arrangements. He is receiving splendid cooperation from the Mining Department of the El Paso Chamber of Commerce and from the business people of El Paso, who feel highly honored in having these mining societies meet with them this year.

Jan 1, 1930

By H. E. Mussey

WHEN the American Institute of Mining Engineers visited the Birmingham district in May, 1888, the four Ensley furnaces (Fig. 1) then FIG. 1.-BLAST-FURNACE DEVELOPMENT IN ALABAMA. completed were referred to as monumental.1 Their dimensions are given in Table 1. Mr. Gordon stated, with considerable satisfaction, that on the preceding day one furnace had produced 165 tons. The blast was

Jan 10, 1924

By T. J. Koppenaal, M. E. Fine

A yield point effect attributed to short-range ordevi?g (SRO) occurs in Cu base Al. At at 296°K varies with heat treatment, decreasing as the annealing ternperature is raised .from 433Oto 598°K. Davies aid Cahn&apos; observed a corresponding decrease in SRO. (523 °K anneal, measured at 77°K) is approximateZy proportional to the variations of A7 with strain rate and testing temperature are also Consistent with the idea that is associated with SRO. A preliminary investigation of the tensile properties of a Cu-Al single crystals showed the presence of a rather strong yield-point effect (drop in flow stress after initial yielding). The object of this research was to investigate its origin and behavior. a Cu-Al alloys are particularly interesting because diffuse X-ray scattering measurements by Davies and Cahn,&apos; Houska and Averbach,&apos; and Borie established the presence of short-range order. The degree of local order may be changed with heat treatment.&apos; cottrel14 suggested that the presence of local order might result in a yield-point effect, and thus the possibility exists here for experimentally ascertaining the importance of short-range order with respect to yield points in these alloys. Since a 12 pct difference exists between the atomic sizes, elastic or Cottrell locking5 must also be considered. Further, Howie and swanne have shown that the stacking fault energy of copper is reduced by aluminum additions. The width of extended dislocations should thereby increase. hus the conditions appear attractive for Suzuki locking.&apos; Finally, the possibility of stress-induced order at dislocations, schoeck locking,&apos; must also be examined, EXPERIMENTAL PROCEDURE Alloys up to 14 at. pct Al were prepared by induction melting high-purity (99.999 pct) Cu and (99.996 pct) Al in a graphite boat under a dynamic vacuum of 5 X 10o mm of Hg. After homogenizing the ingots at 900°C for at least 24 hr under vacuum, they were rolled with intermediate anneals to strips 1.55 mm thick. Single crystals 10 in. long were grown by lowering strips, contained in a split graphite mold sealed in fused quartz at 5 X lo-&apos; mm of Hg, through a single coil induction heater at a constant rate of 1/2 in. per hr. Tensile specimens 1.25 in. long were cut from the single crystal strips and reduced cross sections about 0.7 in. long and 3.0 to 3.5 mm2 in area were introduced by filing and abrading.8 To remove the worked portion about 10 pct of the cross-sectional area was removed by etching. Back-reflection Laue photographs of a filed and etched specimen were taken before and after annealing at 900°c for 24 hr. Small, well defined Laue spots were obtained with no visual difference in the two photographs. Further, specimens with and without the reduced sections began yielding at about the same stress. Hence, for our purposes, filing and abrading did not affect the structure of the specimens. Each single crystal specimen was annealed at 900° c in vacuum for at least 24 hr and furnace cooled; while cooling through the range of 250o to 200°c, the rate was about 55°C per hr. Orientations were determined by the usual back-reflection Laue

Jan 1, 1962

By Ernest N. Patty, Sherwin F. Kelly

The present case history deals with the examination of an almost forgotten nickel prospect, near Winesap, Chelan County, Washington. Although the final results yielded no ore body of commercial importance, the integration of surface and underground exploration, geophysical survey and diamond drilling endow the story with interest. The program is significant because it is illustrative of how the most disappointing expenditures in mining exploration, the "not-finding" costs, could have been reduced had geophysical methods been applied at an appropriate stage in the prospecting work. The Chelan nickel deposit was discovered about 1900, on a sage-brush-covered hillside with a southerly slope, adjacent to Winesap Canyon; the paved highway now extending from Wenatchee northward, up the Columbia River and into Okanogan Valley, is only a mile away. After the deposit was found, three short prospect tunnels were driven into it, and the prospect then remained dormant for 40 years. No. x tunnel was 80 ft. long, running north into the hillside. For the first 35 ft. this adit exposed a typical, iron-stained gossan with streaks and stains of malachite, and small seams of water-soluble nickel sulphate. Beyond this, it encountered a relatively fresh peridotite, well mineralized with blebs of pentlandite. Systematic sampling yielded an average assay of 1.5 per cent nickel and 0.3 per cent copper. Down the hillside to the south and 35 ft. lower in elevation, No. 2 tunnel was driven under No. I and in the same direction, traversing 50 ft. of gneiss before it reached the peridotite. At this point the latter formation showed only scattered blebs of sulphides, not of commercial grade. To the west, on the same contour line as No. I tunnel, and 800 ft. around the hillside from it, No. 3 tunnel was driven 50 ft. into the hill to explore a strong gossan outcrop that assayed 0.5 per cent nickel (Fig. I). From the geological exposures on the hillside and in the tunnels, it appears that the main mass of the hill is formed of an ancient and altered gneiss complex. Into this, a sill-like injection of amphibolite (altered peridotite) was intruded, which, at least in some places, contains commercial quantities of nickel sulphide. Its structure is obscure, but it appears to dip into the hill at an angle of about 15°. In the vicinity of No. 3 tunnel a large dike of quartz diorite cuts the peridotite or amphibolite; about 140 ft. north of the portal of the tunnel, a dike of andesite 00 ft. wide cuts boldly up the hillside to break the continuity of the peridotite in that direction. The geological evidence available was not adequate to decide whether the peridotite forms a continuous sill, or occurs as isolated bodies in the gneiss. Not all of this peridotite carries commercial amounts of pentlandite, but if the mineral-

Jan 1, 1946

By Ernest N. Patty, Sherwin F. Kelly

The present case history deals with the examination of an almost forgotten nickel prospect, near Winesap, Chelan County, Washington. Although the final results yielded no ore body of commercial importance, the integration of surface and underground exploration, geophysical survey and diamond drilling endow the story with interest. The program is significant because it is illustrative of how the most disappointing expenditures in mining exploration, the "not-finding" costs, could have been reduced had geophysical methods been applied at an appropriate stage in the prospecting work. The Chelan nickel deposit was discovered about 1900, on a sage-brush-covered hillside with a southerly slope, adjacent to Winesap Canyon; the paved highway now extending from Wenatchee northward, up the Columbia River and into Okanogan Valley, is only a mile away. After the deposit was found, three short prospect tunnels were driven into it, and the prospect then remained dormant for 40 years. No. x tunnel was 80 ft. long, running north into the hillside. For the first 35 ft. this adit exposed a typical, iron-stained gossan with streaks and stains of malachite, and small seams of water-soluble nickel sulphate. Beyond this, it encountered a relatively fresh peridotite, well mineralized with blebs of pentlandite. Systematic sampling yielded an average assay of 1.5 per cent nickel and 0.3 per cent copper. Down the hillside to the south and 35 ft. lower in elevation, No. 2 tunnel was driven under No. I and in the same direction, traversing 50 ft. of gneiss before it reached the peridotite. At this point the latter formation showed only scattered blebs of sulphides, not of commercial grade. To the west, on the same contour line as No. I tunnel, and 800 ft. around the hillside from it, No. 3 tunnel was driven 50 ft. into the hill to explore a strong gossan outcrop that assayed 0.5 per cent nickel (Fig. I). From the geological exposures on the hillside and in the tunnels, it appears that the main mass of the hill is formed of an ancient and altered gneiss complex. Into this, a sill-like injection of amphibolite (altered peridotite) was intruded, which, at least in some places, contains commercial quantities of nickel sulphide. Its structure is obscure, but it appears to dip into the hill at an angle of about 15°. In the vicinity of No. 3 tunnel a large dike of quartz diorite cuts the peridotite or amphibolite; about 140 ft. north of the portal of the tunnel, a dike of andesite 00 ft. wide cuts boldly up the hillside to break the continuity of the peridotite in that direction. The geological evidence available was not adequate to decide whether the peridotite forms a continuous sill, or occurs as isolated bodies in the gneiss. Not all of this peridotite carries commercial amounts of pentlandite, but if the mineral-

Jan 1, 1946