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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 H. R. Page, A. E. Jr Lee

From the inception of zinc retorting on a commercial scale in the United States in 1890,' the retort employed has undergone wide variations in its composition and manufacture, facilitating in part equally remarkable improvements in furnace capacities. The early day hand made clay retort was charged with carbonates or silicates or with coarse dead roasted concentrates mixed with a large proportion of charge fuel resulting in a relatively low zinc burden and fired 24 hr in direct coal fired furnaces. Its modern counterpart is fabricated in hydraulic presses from clay mixtures containing sizeable amounts of either silicon carbide or silica flour, charged with sintered flotation concentrates to more than three times the early day zinc burden and fired 24 to 48 hr in gas fired furnaces. This paper does not attempt to describe in detail the early day clay retort practice as it is well outlined in treatises by Ingalls,2 Lodin,3 Liehig,4 Hofman5 and others. A brief review of clay retort practice is presented together with a description of the major developments since 1912. Clay Retorts The Belgian type retort, both in the circular and elliptical forms, has been used almost exclusively. Typical dimensions of press made clay retorts around 1910 are shown in Table 1. Variations in these dimensions were used at different plants according to local conditions to a maximum inside diameter of 9 in. and inside length of 54 in. However, the effective heat penetration in a 24 hr firing cycle and the tendency of the retort to bend limited the retort size. Use of the elliptical vessel was an attempt to present a stronger cross-section resisting the tendency to bend and to increase the burden without increasing the depth of heat penetration. One exception to the 48-54 in. length was the 60 in. retort used as early as 1905 at Palmerton by means of supporting the last 12 in. at the butt end with a specially designed furnace back-wall. This backwall construction with the 60 in. retort had been developed and used at Bethlehem by G. G. Con-vers and A. B. DeSaulles. An attempt was made at Blende, Colo. to use even larger retorts of the Rhenish type based on European practice and requiring much higher furnace temperatures. Satisfactory plastic clays capable of withstanding these temperatures were not available, and the plant never operated successfully. PREPARATION OF BATCH Composition of the clay retort by weight was 40 to 50 pct raw clay and the balance "grog." Generally speaking the mix consisted of 7 parts plastic clay to 9 parts grog by volume. Principal source of the clay used was the Cheltenham vein—sometimes referred to as "St. Louis city clay." A typical analysis was A12O3-31.0 pct, SiO2-50.0 pct, Fe2O3-2.5 pct, MgO-0.3 pct, CaO-1.5 pct and loss on ignition 14.0 pct. At the smelter the clay was weathered whenever possible and then crushed to 0.08 in. or finer. Grog consisted of calcined adobies, cleaned retort scrap and cleaned refuse fire brick such as old furnace brick, blast furnace linings, and others. Saggers from ceramic plants and calcined flint clay were later used. The grog materials were ground to 0.12 m. or finer. Occasionally coke dust up to 10 pct of the mix was substituted for a part of the grog following European practice.² Particle size of the grog has a major influence on the retort properties—the larger the grain, the better can the retort withstand thermal shocks, resist bending at furnace temperatures and resist corrosion from slag; the smaller the grain, the lower the loss of zinc vapor through the retort walls. Grog forms the skeleton of the retort, and the clay shrinks around its grains to act as a binder. In the drying process, the grog has a stabilizing effect on the drying rate, decreasing shrinkage and giving up previously absorbed water to the surrounding clay to minimize the danger of cracking or checking.² Grog and clay were mixed through a horizontal pug mill with 10 to 20 pct water added, depending on whether the retort was to be formed by hand or mechanically, more water being required for the hand process. The batch or "mud" extruded from the pug mill was cut in convenient lengths for handling, stacked in piles or in special rooms, covered with wet burlap and allowed to "rot" or age from 1 to 8 weeks to increase plasticity. HAND MOLDING If the retort was to be molded by hand, the mud was repugged after the rotting period and given to the molders. Their molds consisted of 3 sheet iron or wood cylinders, each one third the retort length and defining the outer shape of the retort. Beginning with the bottom section, mud was placed in the form and tamped with a rammer

Jan 1, 1950

By AIME AIME

PRODUCTION of nickel in Cuba, a new source of this metal, has been started by the Nicaro Nickel Co., subsidiary of the Freeport Sulphur Co. Construction of the Nicaro plant in Oriente Province, Cuba, and of a refinery at Wilmington, Del., has been "substantially completed," according to Langbourne M. Williams, Jr., president. The initial rate of output is less than capacity operation, which will be reached when "the customary problems incident to the commencement of full- scale production are eliminated." The plant in Cuba for the treatment of the extensive nickeliferous ores of that region, and the refinery, were built by Nicaro for the account of the Defense Plant Corp. construction has been under way for eighteen months. The plants, when operating at capacity, are expected to increase America's sup-

Jan 1, 1944

By C. A. Heiland

IN the third year of war, geophysical oil exploration broke all records to keep pace with the demand for increased reserves. Geophysical prospecting for strategic and other minerals also grew in scope. The trend of geophysical oil exploration throughout the war is indicated in Fig. 1. This figure resembles the one presented in the last annual review issue of M&M (February 1944, p. 62), corrected for data on-operations in the last months of 1943 and extended into 1944. The average number of geophysical oil prospecting crews in the field in 1944 was 451, whereas the total in 1943 was 340. An upward trend in the use of all geophysical methods is clearly evident throughout the year. Seismic operations were 67 per cent, gravimeter surveys 28 per cent, and other methods 5 per cent of the total. Seismic reflection and refraction crews increased 181/2 per cent over the 1943 average, gravimeter operations increased 73 per cent, and other methods gained 92 per cent. The increase in seismic and gravimeter operations, particularly the latter, is probably due to an increased production of equipment and the organization of new service companies. Miscellaneous methods gained largely through a revival of magnetometer work, and an increase in the use of torsion balances and electrical equipment. In comparison with the prewar average of 1941, it is noteworthy that seismic operations increased 58 per cent, gravimeter operations 88 per cent, and miscellaneous other methods gained as much as 283 per cent. Unfortunately the expansion in the number of geophysical crews does not indicate a proportionate increase in efficiency or in areal coverage, as will be brought out later.

Jan 1, 1945

By Charles F. Park

Mining in the Pacific Coastal Region has passed through three stages of development. First came the gold rush days, a period when gold and silver were the objects of intensive search. Second was the development of the high-grade base-metal deposits and of the smaller and more refractory precious metal deposits. And third was the modern period that covers the discovery and development of the rarer metals, of large low-grade deposits, and of the industrial and nonmetallic minerals. Brief discussions of the salient geologic features of the region and of a few of the better known mineral deposits are given. The regional geology is favorable for continued exploration and discovery of new deposits and districts.

Jan 1, 1968

By A. H. Daane, B. J. Beaudry

The yttl-irtm-nickel system has been investigated by metallo-graphic, thermal ard X-ray methods. There are nine intermetal-lic compounds present: Y3Ni, Y3Ni,, YNi,, YNi,, Y,Ni,, YNi,, and Y,Nil, which rodergo peritectic decomposition and YNi and YNi, which melt congruently. There are eutectics at 34.8, 57.5 and 93.3 at pct Ni which melt at 805°, 950 and 1285 C, respectively. Crystallographic data are given for YNi, YNi,, YNi,, YNi5and Y,Ni,. The terminal solid solubility is low. In recent work by Fisher and Fullhart' in which yttrium crucibles to be used for containers of uranium-chromium eutectic were clad with a high nickel steel, a low melting phase formed where the yttrium was in contact with the steel. To investigate the cause of this phenomenon, Daane, et al. made a survey study of yttrium systems with chromium, manganese, iron, and nickel and found an 18 wt pct Ni (weight percent) alloy to melt at approximately 950°C. voge13 in a study of the systems of cerium, lanthanum, and praseodymium with nickel found six compounds of the same formula in each system, i.e., R3Ni, RNi, RNiz, RNi3, RNi4, and RNi5. Since there is a great similarity between yttrium and the rare earths, compound formation was likely to occur also in the yttrium-nickel system. voge13 observed that the ability of transition metals to form compounds with the rare earths diminishes in the order nickel, cobalt, and iron while no compounds are formed with manganese, chromium, and titanium. Available data on yttrium systems with these elements indicate similar behavior, except that a slightly greater tendency toward compound formation is apparent. Simple eutectics with limited solid solubility have been found in the yttrium-titanium4 and yttrium-chromium5 systems, but at least one compound is present in the yttrium-manganese system.' In the case of yttrium-iron, at least three compounds are present,5 while only two are present in the cerium-iron,' and no compounds are present in the lanthanum-iron5 system. On the basis of both the trend in the alloying behavior of yttrium and rare-earth metals with the first transition series metals and the large number of compounds in the cerium-nickel system, one might expect six or more compounds to form in the nickel-yttrium system. A consideration of Hume-Rothery's rules of alloying based on size factor, electronegati-vity, and valence predicted low terminal solid solubility and possible compound formation. The present study was undertaken to confirm these predictions of

Jan 1, 1961

By W. L. Zeigler

The mill is a departure from gravity concentration and has gained a reputation for the low initial cost of erection, extreme simplicity and the low cost of milling on the refractory character of the ore which is treated. The structure is of heavy timbers with board walls and corrugated roofing. The floors are concrete and the general layout is of the ground floor only. All the machines are driven by individual motors and the transmission usually found in such a plant is entirely absent. (Fig. 2.) ORE The ore from the Galena mine is sheared Wallace quartzite replaced with quartz, siderite, pyrite, sphalerite, gray copper (tetrahedrite) and very fine-grained argentiferous galena. Much of the silver is undoubtedly associated with the gray copper, the amount, however, apparently varies with different localities in the vein. An average assay of the ore is 6.0 per cent. lead, 1.3 per cent. zinc, 0.3 per cent. copper, 17.5 per cent. iron and 4.5 oz. silver. The concentrates produced are of an average 53 per cent. lead, 40 oz. silver, 8 per cent. zinc and 11 per cent. iron, while the tailings are 0.5 per cent. lead, 0.4 oz. silver. With the present market for zinc no attempt is made to produce a zinc concentrate.

Jan 1, 1927

By S. C. Oliphant, George H. Fancher

Two samples of drilling mud from the Hastings oil field, Texas, were tested, and the solids in each were separated into small fractions of a limited range in particle size. The mineral composition of each fraction by X-ray analysis is reported, and the results are interpreted in relation to other measured quantities, such as the physical properties of the original samples and the density of the various fractions. It was found that only 7 to 8 per cent by weight of the solids was of colloidal dimensions and that neither of the samples contained montmorillonite but did contain illite. Considerable calcite was also found. These data should be useful to the geologist interested in the origin and history of sedimentary strata as well as to the petroleum engineer.

Jan 1, 1942

By R. A. Kinzie

THE major operating costs in a cement plant are labor, power, and fuel. The opportunities and methods of savings in labor and power parallel other industries. Because our industry's use of fuel is large and its method of use is unique, the devices adapted for saving fuel are of interest. In normal practice the raw material in the form of a dry powder or a mud is fed to a rotating inclined cylinder where it is dried, calcined, and clinkered. In a Lepol kiln the raw material is rolled up into little balls with water in the nodulizer, fed to a traveling grate on which it is dried and partially calcined and tben to a short kiln where the process is finished. The more usual methods of saving heat have been the use of long kilns with chains in wet process plants and the use of waste heat boilers in dry plants (fig. 1). At the Santa Cruz plant we have used a Lepol system kiln and achieved a fuel 'economy of 4.7 gal per bbl of clinker or about 704,-000 Btu per bbl. This can be compared with our old kilns using the same fuel and raw material which consume 7.8 gal per bbl or 1,130,000 Btu per bbl. These are gross Btu and not the net figures usually used in reports of foreign companies. Our average consumption for 1948 was 708,000 Btu per bbl of Lepol clinker, based on cement sold, and oil purchased. The overall fuel consumption for 1948 was 6.61 gal per bbl. The Lepol system consists of a nodulizer, a grate, and a kiln. There is the usual accessory equipment which is similar to that used in other cement plants. The following description covers the equipment in operation at our plant. The nodulizer is a cylinder 8 ft 10 in. in diam by 18 ft long set on a slope of 1/2 in. per ft, and rotated at about 10 rpm. It is driven by a 50 hp wound rotor motor drawing about 9 kw. The inside of the shell is cleaned by a continuously operating scraper which draws 0.1 kw. The dry raw material is fed to the nodulizer where, being sprayed with water, it rolls into small spheres about 90 pct being between 1/2 and 1 in. in diam and containing about 15 pct water. The nodules fall on the traveling grate where the hot gas from the kiln is pulled down through the

Jan 1, 1951

By R. A. Kinzie

THE major operating costs in a cement plant are labor, power, and fuel. The opportunities and methods of savings in labor and power parallel other industries. Because our industry's use of fuel is large and its method of use is unique, the devices adapted for saving fuel are of interest. In normal practice the raw material in the form of a dry powder or a mud is fed to a rotating inclined cylinder where it is dried, calcined, and clinkered. In a Lepol kiln the raw material is rolled up into little balls with water in the nodulizer, fed to a traveling grate on which it is dried and partially calcined and tben to a short kiln where the process is finished. The more usual methods of saving heat have been the use of long kilns with chains in wet process plants and the use of waste heat boilers in dry plants (fig. 1). At the Santa Cruz plant we have used a Lepol system kiln and achieved a fuel 'economy of 4.7 gal per bbl of clinker or about 704,-000 Btu per bbl. This can be compared with our old kilns using the same fuel and raw material which consume 7.8 gal per bbl or 1,130,000 Btu per bbl. These are gross Btu and not the net figures usually used in reports of foreign companies. Our average consumption for 1948 was 708,000 Btu per bbl of Lepol clinker, based on cement sold, and oil purchased. The overall fuel consumption for 1948 was 6.61 gal per bbl. The Lepol system consists of a nodulizer, a grate, and a kiln. There is the usual accessory equipment which is similar to that used in other cement plants. The following description covers the equipment in operation at our plant. The nodulizer is a cylinder 8 ft 10 in. in diam by 18 ft long set on a slope of 1/2 in. per ft, and rotated at about 10 rpm. It is driven by a 50 hp wound rotor motor drawing about 9 kw. The inside of the shell is cleaned by a continuously operating scraper which draws 0.1 kw. The dry raw material is fed to the nodulizer where, being sprayed with water, it rolls into small spheres about 90 pct being between 1/2 and 1 in. in diam and containing about 15 pct water. The nodules fall on the traveling grate where the hot gas from the kiln is pulled down through the

Jan 1, 1951

By A. B. Kinzel

ALTHOUGH vanadium is an important constituent of almost every brand of high-speed steel manufactured today, little is known as to its role in this series of alloys. The now standard 18 per cent tungsten and 4 per cent chromium constituents were the result of developments and improvements before 1903. In that year, vanadium was added to the steel by Dr. J. A. Mathews. This proved to be an outstanding con-tribution, so that the standard high-speed steel of today is generally referred to as the 18-4-1 type-18 per cent W, 4 per cent Cr and 1 per cent V. Various modifications have been attempted, since then, but with the exception of the introduction of cobalt, no striking improvements have occurred. Vanadium has been increased to 2 per cent in many of the common brands, and at least one brand contains 5 per cent V in addition to the usual tungsten and chromium. The carbon has been gradually increased so that it is not uncommon to find 0.80 per cent in the standard grade, although the carbon in the 5 per cent vanadium steel mentioned above is approximately 1 per cent, and even 1.2 per cent has been made experimentally. The role of vanadium has never been well understood, and although vanadium has been increased to 2.5 and even 5 per cent in certain high speed steels, it has not been generally established that the higher vanadium contents 'result in superior cutting properties. However, in all of the investigations, of which the writers are aware, the interrelation of the vanadium and carbon contents was not considered as a prime factor. The carbon content was held at 1.2 per cent or less-generally at the usual 0.65 to 0.75 per cent.

Jan 1, 1932

Production of crude oil in Bahrein Island during 1938 reached the new high of 8,297,998 bbl., an increase of 535,734 bbl, over the 1937 production of 7,762,264 bbl. The monthly average during 1938 was 691,500 bbl. and the daily average 22,734 barrels. To the close of December 1938, a total of 52 producing wells have been drilled. During the year 13 producing wells were completed and 11 wells were drilling or partly completed as of Dec. 31, 1938. The total footage drilled as of Dec. 31, 1938, amounted to 165,339 ft., of which 60,739 ft. was drilled during 1938. The production from the main pay comes from a depth of approximately 2200 ft. and the gravity of the oil is 33" A.P.I.

Jan 1, 1939

Production of crude oil in Bahrein Island during 1938 reached the new high of 8,297,998 bbl., an increase of 535,734 bbl, over the 1937 production of 7,762,264 bbl. The monthly average during 1938 was 691,500 bbl. and the daily average 22,734 barrels. To the close of December 1938, a total of 52 producing wells have been drilled. During the year 13 producing wells were completed and 11 wells were drilling or partly completed as of Dec. 31, 1938. The total footage drilled as of Dec. 31, 1938, amounted to 165,339 ft., of which 60,739 ft. was drilled during 1938. The production from the main pay comes from a depth of approximately 2200 ft. and the gravity of the oil is 33" A.P.I.

Jan 1, 1939

By Blake M. Caldwell, Harrold D. Owen

The continued use of permanent-type, we11 completion Iras pointed LIP the need for more powerful through-tubing perforating equipment. A new expendable shaped charge perforator has been developed in which the charges are run through the tubing in a vertical position and then opened out to a horizontal position when at the desired shooting zone. The performance of this new tool is comparable to that of conventional casing type perforators. This paper describes the new tool and its applications. Performance data in targets rind under various actual well conditions are presented. INTRODUCTION The general acceptance and increasing use of pernianent-type well completion by the oil industry has indicated the need for through-tubing perforating equipment with performance comparable to that of conventional casing type perforators.'.' To achieve the performance desired, a new, expendable, shaped charge perforator has been developed with adequate power for effectively perforating both 51/2-in. and 7-in. casing below tubing. This has been accomplished without resulting severe damage to the casing and can he obtained under extremes of pressure and temperature. The design of this new tool is a radical departure from previous concepts of tubing guns since it was felt that the two major limitations imposed by previous designs severely handicapped performance. The first of these limitations was the restriction imposed upon the length of the charges by the necessary small inside diameter of the carrier. This resulted from the necessity for compromises in designing a hermetically sealed gun small enough to allow free passage through 2-in. tubing, yet with sufficient wall thickness to withstand some minimum pressure. In an effort to overcome this limitation, the charges were in some cases placed at an angle with the axis of the well with resulting loss in effective depth of penetration. The second limitation concerned the extreme distance of the charges from the casing when the gun was fired. An example of the latter situation is demonstrated by the fact that a conventional 13/4-in. tubing gun centralized in 51/2-in., 17-lb casing has a clearance of more than 1-9/16-in. on all sides. This means that a considerable portion of the force from each charge must be expended in penetrating this 1-9/16-in. of well fluid before it even reaches the casing. With the gun centered in 7-in., 23-lb casing, the amount of fluid penetration required of each jet is in excess of 21/4-in. In a situation where the 13/4-in. gun is touching one side of the 51/2-in., 17-lb casing, the jets from the charges on the opposite side have more than 3%-in. of fluid to penetrate. A similar situation in 7-in., 23-lb casing necessitates the penetration of more than 4-9/16-in. of fluid. The penetrating effect of the charges is severely impaired in passing through so much fluid. Under such conditions, there is usually insufficient force left to penetrate the casing. Previous studies have shown the decided disadvantage of having to shoot through such an excessive quantity of well fluid.'.' The problem. then, was how to position large powerful charges close to the casing prior to firing and still manage free passage through 2-in, tubing. The solution appeared to be in some method of folding sealed charge units in an open type, expendable. 13/4-in. carrier for travel through the tubing, and unfolding them upon their arrival at the shooting zone in the casing. It was desired to control the mechanism electrically from the surface through the regular shooting cable, making the release of the charges independent of the physical characteristics of the well, yet with a provision to prevent the gun from firing until the

Jan 1, 1955

By Robert W. Thompson, Darry A. Ferguson

In August, 1977 the United States Congress passed Public Law 95-87 which has been commonly referred to as the Surface Mining Control and Reclamation Act (or SMCRA). This title is somewhat misleading in that the act creates as many regulatory requirements for underground coal mines as for surface coal mines. How- ever, this paper will concentrate on the effects of SMCRA on designs and stability for various activities of surface coal mines, specifically those concerning "Disposal of Excess Spoil". In the past, the objective of most laws and regulations applicable to the mining industry has been to achieve goals of safety or environmental concerns for all types of mining. However, SMCRA is unique in that it sought to regulate-all-activities of coal mining from initial data gathering, exploration, mine design and facilities design to reclamation activities and design. More specific to the stability of the activities of surface coal mining, special criteria for design and construction performance standards have been promulgated for the construction and maintenance of roads, dams and embankments, coal processing waste, ex- cess spoil piles, backfilled reclamation and pit slides. The criteria for design for the previously mentioned facilities has been enumerated by regulations and in many ways restrict the use of site specific design parameters. The restrictions imposed by regulations significantly inhibit the engineer's ability to develop the best practical and economical design and in some cases may re-

Jan 1, 1983

By Nicholas J. Grant, John S. Benjamin

A series of composite alloys containing a high volume of NiAl, Ni3Ah or CoAl, bonded with 0 to 40 vol pct of a ductile metal phase, were prepared by powder blending and hot extrusion. The binder metals were of four types: pure nickel or cobalt, near saturated solid solutions of aluminum in nickel and cobalt, type 316 stainless steel, and niobium. Sound extrusions were obtained in almost all instances. Studied or measured were the following: interaction between the alunzinides and the binders, room-temperature modulus of rupture values, 1500° and 1800°F stress rupture properties, hardness, structure, and oxidation resistance. Stable structures can be produced for 1800°F exposure, with interesting high-temperature strength and good high-temperature ductility. Oxidation resistance was excellent. A large number of experimental investigations have been made of the role of structure on the properties of cermets and composite materials. Gurland,1 Kreimer et al.,2 and Gurland and Bardzil3 have indicated the preferred particle size in carbide base cermets to be about 1 µ, with a hard phase content of 60 to 80 vol pct. The optimum ductile binder thickness was noted to be 0.3 to 0.6 µ.1 Complete separation of the hard phase particles by the binder is important in reducing the severity of brittle fracture.' The purpose of the present study was to produce structures comparable to the conventional cermets, using a series of relatively close-packed intermetal-lic compounds rather than carbides as the refractory hard phase, and to study the effects of binder content and composition on both high- and low-temperature properties. The selected intermetallic compounds were particularly of interest because of the potential they offered in yielding room-temperature ductility. The highly symmetrical structures are known to possess high-temperature ductility and room-temperature toughness. Based on a ductile binder, the alloys were prepared by the powder-metallurgy route to avoid melting and subsequent alloying of the matrix, and were extruded at relatively low temperatures. It was expected that the composite alloy would retain useful ductility. In contrast, infiltration and high-temperature sintering led to alloying of the matrix and to decreased ductility. The systems Ni-A1 and Co-A1 were selected for this study. In the Ni-A1 system the compounds NiA1, having an ordered bcc B2 structure, and Ni3Al(?1), having an ordered fcc L12 structure, were chosen. In the system Co-A1 the intermetallic compound CoAl with an ordered bcc B2 structure was used. ALLOY PREPARATION The intermetallic compounds, see Table I, were prepared by using master alloys of Ni-A1 and CO-A1, with additions of either cobalt or nickel to achieve the desired compositions. The master alloy in crushed, homogenized form, was melted with pure nickel or cobalt in an inert atmosphere, cold copper crucible, nonconsumable tungsten arc furnace. The resultant intermetallic compounds were homogenized at 2192°C in argon, crushed, and dry ball-milled in a stainless mill to -100 and -325 mesh for the Ni-A1 compounds and to -325 mesh for the CoAl compound. Finer fractions were separated for some of the composite alloys. Several ductile binders were utilized. These included Inco B nickel, 5µ ; pure cobalt, 5 µ, from Sher-ritt Gordon Mines, Ltd.; fine (-325 mesh) niobium hydride powder; fine (15 µ) type 316 stainless-steel powder; and near-saturated Ni-A1 and Co-A1 solid-solution alloys, also in fine powder form. The niobium hydride was decomposed above about 700°C in processing of the compacts in vacuum to produce niobium powder. The Ni-7.1 pct A1 and the Co-5.3 pct A1 solid-solution alloys were prepared from pure nickel or cobalt and pure aluminum by nonconsumable tungsten arc melting under an inert atmosphere. The ingots were homogenized, lathe-turned to fine chips, and dry ball-milled in air to -325 mesh powder. These solid-solution alloys are designated NiSS and CoSS; see Table I. Subsequently the hard and ductile phases were dry ball-milled as a blend. Experiments clearly established the need to coat the hard particles with the ductile binder to optimize subsequent hot compaction by extrusion. Ordinary dry mixing usually resulted in nonhomogeneous alloys which were quite brittle. Conventional cermets are consolidated by liquid phase sinteiing or infiltration, which resulis in undesirable and uncontrolled alloying of the binder phase. For this study, a loose (unsintered) powder-extrusion process was emploved, minimizing reactions between binder and hard particle, thereby permitting much greater control of composition and structure. The constituent powders were first mixed in the desired

Jan 1, 1967

By Frank S. Hurley

To comply with the California Mine Safety Orders pertaining to transportation of mine workers over mine railroads, the management of the New Idria mine in San Benito County, Calif., designed special man-cars. The new regulations, dated October 1958, make it mandatory for mine man-cars to be designed for carrying five or more men to and from the mine workings. At the New Idria mine, largest single producer of mercury in the U.S., each shift normally employs about 70 men. However, with the possibility of a future increase in the work force, seven man-cars were constructed with a total carrying capacity of 84 men. The general design of the mine man-cars was dictated by safety regulations plus the physical characteristics of the mine-low headroom, limited side clearance in haulageways, and track irregularities due to ground swells. The cars measure 10 ft long, 44 in. wide, and 50 in. high and were built at an average cost of $450 per car. All assembly work was done at the mine site.

Jan 7, 1960

By R. D. MacDonald

The search for metalliferrous deposits in the Labrador-Ungava Trough of Canada dates from 1929 when non-ferrous minerals were the main quest of prospectors in this area. Many gossans, resulting from the oxidation of sulfides, were located and examined but none of these led to any orebodies of commercial value. However, such investigations did much to arouse interest in this area and fostered further mapping and prospecting. Since this early period in the history of prospecting in the Trough, the presence of iron ore deposits and potential iron producing sites has commanded the center of attention of exploration personnel. The iron-bearing rocks of southwestern Labrador have in recent years come into prominence as possible sources of iron ore of the beneficiation type. Exploration and development work has been carried on in concession areas granted by the Newfoundland Government to the Labrador Mining and Exploration Co. Ltd. and to the Newfoundland and Labrador Corp. In the course of evaluating its holdings in the Wabush Lake area (Fig. l), Labrador Mining and Exploration Co. has conducted geological, geophysical and diamond drilling programs.

Jan 10, 1960

By P. A. Laxen, H. R. Spedden, A. M. Gaudin

WHEN a mineral particle is fractured, bonds between the atoms are broken. The unsatisfied forces that appear at the newly formed surface1 are considered to be responsible for the adsorption of ions at the mineral surface. A knowledge of the mechanism and extent of ion sorption from solution onto a mineral surface is of interest in the development of the theory of flotation.2,3 Study of the adsorption of sodium from an aqueous solution on quartz offers a simple approach to this complicated problem. The availability of a radioisotope as a tracer element meant that accurate data could be obtained.4,5 Three main factors which appeared likely to affect the adsorption of sodium are: 1-concentration of sodium in the solution, 2-concentration of other cations in the solution, and 3-anions present in the solution. Hydrogen and hydroxyl ions are always present in an aqueous solution. By controlling the pH, the concentration of these two ions was kept constant. The variation in the amount of sodium adsorbed with variation in sodium concentration was then determined under conditions standardized in regard to hydrogen ion. The effect of concentration of hydrogen ions and of other cations was also measured. A few experiments were made to get a preliminary idea on the effect of anions. The active isotope of sodium was available as sodium nitrate. Standard sodium nitrate solutions were used throughout these experiments except when the effects of other anions were studied. It was found that sodium adsorption increased with sodium-ion concentration, but less rapidly than in proportion to it. Increasing hydrogen-ion concentration, or conversely decreasing hydroxylion, brings about a comparatively slight decrease in sodium-ion adsorption. Increasing the concentration of cations other than hydrogen or sodium decreases somewhat the adsorption of sodium ion. It would appear as if the kind of anion is a secondary factor in guiding the amount of sodium ion that is adsorbed. Materials and Methods Quartz The quartz was prepared as in previous work in the Robert H. Richards Mineral Engineering Laboratory4 except for the refinement of using de-ionized distilled water for the final washing of the sized quartz, prior to drying5 To minimize the laborious preparation of quartz, experiments were made to determine whether the sodium-covered quartz could be washed free of sodium and re-used. The experiments were successful as indicated by lack of Na' activity on the repurified material and by its characteristic sodium adsorption. Table I gives the spectrographic analyses of the quartz used. The quartz ranged from 16 to 40 microns in size, averaging about 23 microns (microscope measurement), and had a surface of 1850 sq cm per g (lot I), 2210 (lot II) and 2000 (lot III) as determined by the Bloecher method.6 Radioactive Sodium Method of Beta Counting for Adsorbed Sodium: Na22, the radioisotope of sodium, possesses convenient properties.7 It has a half-life of 3 years, thus requiring no allowance for decay during an experiment. On decay it emits a 0.575 mev ß radiation and a 1.30 mev ? radiation. The decay scheme is illustrated in the following equation: [Y Nam S. - 'Net 3 years] The ß radiation is sufficiently strong to penetrate an end-window type of Geiger-Mueller counting tube. This, in turn, makes it possible to use external counting, a great advantage in technique. Furthermore, it permits the assaying of solids arranged in infinite thickness, while assaying evaporated liquors on standardized planchets. The equipment used was standard and similar to that employed by Chang8 The original active material was 1 ml of solution containing 1 millicurie of Na22 as nitrate. This active solution was diluted to 1000 ml. Five milliliters of this diluted active solution was found to give a quartz sample a sufficiently high activity for accurate evaluation of the sodium partition in the adsorption measurements. Also, 1 ml of final solution gave a sufficiently high count for precision on the liquor analyses. The sodium concentration of the diluted active solution was 1.2 mg per liter, so that 6 mg of sodium for 60 ml of test solution and 12 g of quartz was the minimum amount used. The active solution was stored in a Saftepak bottle. Procedure for Adsorption Tests: The method consisted of agitating 12 g of quartz with 60 ml of solution of known sodium concentration for enough time to establish equilibrium between the solution and the quartz surface. The quartz was separated as completely as possible from the solution by filtering and centrifuging. The activity on the quartz and in the equilibrium solution was measured and the partition of the sodium was calculated from the resulting data. The detailed procedure for the adsorption test is set forth in a thesis by Laxen5 In brief, it included the following steps: 1-Ascertainment of linearity between concentration of Na22 and activity measured. 2-Evaluation of factor to translate activity on solid of infinite thickness in terms of activity on an evaporated active film of minute thickness, on the various shelves of the counter shield. 3-Taking precautions to avoid evaporation of water during centrifuging.

Jan 1, 1952

By Robert F. Brindisi

There are approximately thirty to forty operating mills in the United States which are currently producing uranium yellow cake. This figure includes a significant number of in situ and by-product operations however the conventional acid leach flowsheet depicted in Figure 1 is the most common process used by domes- tic yellow cake producers. Ground ore is leached with sulphuric acid and the solids are separated from the pregnant liquor and washed in a series of counter- current decantation thickeners. After clarification of the pregnant liquor, uranium is recovered from the solvent extraction/stripping circuit by the addition of ammonia with precipitation of ammonium diuranate, or simply yellow cake. The yellow cake precipitate is then thickened in a conventional thickener before re- porting to the dewatering/washing unit operation. De- watered solids subsequently report to thermal drying, and clarified liquor from the dewatering/washing step is recycled to the yellow cake thickener. The yellow cake dewatering/washing unit operation is generally accomplished using either a solid bowl centrifuge or a rotary vacuum drum filter. In most cases the underflow from the yellow cake thickener is pumped to the dewatering/washing unit operation at approximately fifteen to twenty percent solids. For the purpose of this discussion ammonium diuranate, the product of ammonia precipitation from a conventional acid-leach uranium mill flow sheet, as discussed earlier, will be considered. With only slight adjustment this discussion will be equally applicable to sodium diuranate or uranium perioxide precipitates and to either in situ or by-product operations where such precipitates are generated. Some of the domestic uranium mills which have selected solid bowl centrifuges instead of vacuum drum filters for the dewatering/washing unit operations are listed in Figure 2. Why have these uranium mills, as well as others not listed, chosen to use the solid bowl centrifuge for this particular application? It is simply a matter of performance, reliability and economics. In the remainder of this paper the performance, initial cost, installation cost, maintenance and power consumption of both the solid bowl centrifuge and the rotary vacuum drum filter will be com- pared. In order to provide a consistent basis for comparison the appropriate sized (centrifuge or filter) unit for the dewatering/washing of 400 to 500 pounds per hour of yellow cake solids will be considered. Such a flow will usually be generated by a typical 3,000 ton per day (acid-leach) uranium mill.

Jan 1, 1980