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By William W. Wellborn

Abrasives are substances both natural and synthetic that are used to grind, polish, abrade, scour, clean, or otherwise remove solid material usually by rubbing action (as in a grinding wheel), but also by impact (pressure blasting). Facts about the discovery, early use, adaptation, and development of abrasives prior to the 20th century are few. From the sparse, earliest information that is available, it would appear that the science of abrasives began when, to fashion his tools, man began to select certain kinds of rocks in preference to others because of their peculiar properties. There is evidence that stones were sawed by some sort of grinding machine as early as 4000 BC in Egypt. The real expansion of abrasives use coincided with the beginning of metallurgy in the Middle East around 2000 BC. The earliest confirmed use of abrasives for grinding metal comes from a steel dagger found along with a sharpening stone which was dated at about 1500 BC. Natural and manufactured abrasive substances play an important role in fashioning and finishing numerous products with a wide variety of end-use applications. The mining and preparation of crude, natural abrasives for their markets and the manufacture of synthetic and other abrasive materials contribute significantly to the US economy (Table 1). In addition, just as sawing to specific dimensions enhanced the value of the Egyptian stone, the use of abrasives in fabrication of modem products, and the addition of abrasives to others, add value. GENERAL CONSIDERATIONS The most important physical properties of materials that qualify as abrasives are hardness, toughness (or rigidity), grain shape and size, character of fracture (or cleavage), and purity (or uniformity). To fabricate bonded abrasive products such as grinding wheels, additional considerations include stability under high heat and bonding characteristics of grain surfaces. The economic factors of cost and availability are always important. No one singular property is paramount for any use. For some applications extreme hardness and toughness are needed, as in diamonds for drill bits. For other uses, factors of greatest importance are hardness and ability to break down slowly and, in the process, develop fresh cutting edges when grains become worn. In garnet sandpaper, for example, highly cleavable, friable, or extremely tough grains are not desirable. For other uses, extreme hardness may be objectionable. Abrasives for dentifrices and for glass-cleaning soaps are cases in point. For the most efficient uses in more critical applications, different types of abrasives are rarely completely interchangeable. While crushed quartz and garnet are both used in sandpaper, the two abrasives are not at all interchangeable in their applications. In the final analysis, the choice of a high-grade abrasive depends upon the quality and quantity of work performed by the abrasive per unit of cost. Initial cost of an artificial abrasive may be much greater than that of a natural abrasive, but the artificial mineral may do so much better work and do it so much faster that the ultimate cost is less. It is for this reason that artificial abrasives have largely replaced natural abrasives.

Jan 1, 1994

By John E. Litz

The Denver Process is a chemical/metallurgical process for treating radioactive or mixed wastes by extraction and concentration of the radioactive and/or hazardous component(s). The process was developed at bench scale, evaluated at pilot scale, and will be implemented at full scale to remediate a low-level radioactive sludge. Six unit operations comprise the process: feed handling, leaching, solvent extraction, electrodeposition, precipitation, and tailing preparation. Unit operation design may be modified to tailor the process for varying feed stock composition. Radioactive or mixed waste is reacted with aggressive reagents in an aqueous phase, solubilizing targeted components. These components are then selectively extracted and concentrated in subsequent steps. Waste solids and solution exit the process as non-hazardous non¬radioactive waste. Dehydration of aqueous waste by submerged combustion eliminates aqueous discharge.

Jan 1, 1994

By Roberto Waisberg I.

El presente texto refiere brevemente la historia del proceso de producción y costos del oro en las minas de Chile hasta el año 1956. Desde la llegada de los españoles al Perú, el prestigio de producción de oro en Chile despertó la codicia de los españoles, registrándose un aumento de producción de 2000 kg. entre los años 1545-1560, que posteriormente se reducirá a la mitad de los años restantes del siglo. Luego, en el siglo XVIII, la producción asciende de 400 kilos nuevamente hasta alcanzar los 3100 kg anuales. La guerra con España y el nuevo gobierno producen una decaída nuevamente en el oro, con promedios de 300 kg. Sin embargo, el periodo de 1931-1956 representa la etapa más importante enl la historia cuprífera de Chile puesto que el volumen de producción registra 6 000 kg. anuales. La contribución de este auge se debe a las medidas gubernamentales que buscaron resarcir la crisis económica producida en 1930, a través de la transformación de sus planteles de beneficio y posterior adaptación para la flotación de minerales de oro, asi como la construcción de plantas de cianuración, desarrollando la explotación en diversos yacimientos mineros. De otro lado, con el advenimiento de la Segunda guerra mundial la producción desciende a 2938 kg, siendo el valor más bajo registrado hasta 1933. Posteriormente y pese a la reducción de producción aurífera, hay un repunte del contenido del oro de minerales y concentrados, generando un incremento de la minería del cobre en los últimos años. En 1955, la producción anual de la pequeña y mediana minería establece un 73.8% para el primero, mientras que el segundo llega a 26.4% del total. El incremento adicional del oro dependerá de la ampliación de la planta fundicional de Paipote y de la Fundición de la zona central, estimándose un total de 1000 kg anuales. Así para interesar a los productores auríferos en la reiniciación de sus explotaciones, deben establecerse políticas de estabilidad en los precios que permitan afrontar los desembolsos de la instalación de las faenas y la rehabilitación de las minas.

Jul 20, 1959

Prevent the explosion of methane gas or coal dust by replacing air with inert gas inside highwall mine excavations. Approach Use combustion gases as a source of inert gas with low oxygen content to blanket the cutting, head of highwall mining machines. This gas is placed strategically at the collar of the excavation where it enters the hole to replace the mined coal volume. The elimination of oxygen from the hole prevents explosions of gas or dust. How It Works Highwall mining methods arc used when surface mining becomes uneconomical. One method of highwall mining involves use of a horizontal auger to enter the coalbed from a surface mine bench. The coal is mined in a series of horizontal, parallel, and large-diameter holes. Historically, little methane has been associated with auger mining. However, mining technology has extended the reach of auger mining, which can increase the occurrence of methane. Methane explosions at auger mining operations have been increasing. Methane has been the primary fuel source of these ignitions, but coal dust also may have been involved.

Jan 1, 1993

By J. W. Leonard

Breakage, handling and treatment techniques that result in the placement of greater quantities of coal in fixed volume haulage units such as ships, railroad cars, trucks, mine cars, shuttle cars and scoops can do much to increase productivity by improving bulk density to reduce delivered cost, surprisingly enough, only in steel industry by-product coking has this subject been investigated and applied to practice, where as much as 20% or more productivity has been achieved. Research at UK Mining using popular coal mine loading cycle simulators shows that for typical Central Appalachian mining conditions a 10 to 20 percent gain in bulk density can produce a 15 to 30 percent increase in the loading cycle time of fixed volume shuttle-car/scoop based mining systems.

Jan 1, 1989

By L. B. Lindemuth

To those who are in a position to carry the thought to a conclusion; I would like to present a theory for the cause of blisters in galvanized sheets. Blisters that are caused from piping and shrinkage cavities will not be considered because the reason for their formation is entirely obvious. Blisters that occur periodically with no apparent system of ocation, number, size or grouping, and then disappear with no. apparent change in manufacturing method, are the ones to which this theory would apply. Examination of blisters of the latter type has shown that they contain hydrogen gas at atmospheric pressure, and that occasionally traces of oxide or sulfide are found on the inner surface of the blisters. These two characteristics have led to the rather general idea that hydrogen gas generated in pickling, and that iron oxide supposed to be inherent in the ingot and causing weak spots were, in some manner, responsible for the development of blisters in galvanizing. The theory which I wish to advance is that these blisters are caused by moisture that has penetrated the steel after the annealing operation. This moisture comes from the atmosphere, or from pickling or washing and when it is suddenly expanded into superheated steam in galvanizing, the increase in volume is so great that it cannot escape with sufficient rapidity through the small opening where it entered as moisture.The pressure so generated forces the sheet apart along those lines of lamination ;hat may be found in practically all flat sections of steel rolled comparatively cold. The increase in volume. from moisture to steam at 800° F. is 2244 times the original volume of the moisture. The pressure generated by ;he moisture heated to 800° F. and entirely confined would be enough to expand or explode any container. This would explain the source of sufficient pressure to cause a sheet to blister.

Jan 3, 1927

By J. J. Strohl, H. J. Schwellenbach

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By H. J. Schwellenbach, J. J. Strohl

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.' By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

In the western Woodlark Basin, propagation of seafloor spreading is occurring directly into the continental crust of Papua new Guinea. The spreading is not strictly in a back-arc tectonic setting, rather it appears associated with microplate rotations related to the complex convergence of the Indo-Australian and Pacific plates in the southeastern Pacific region. Volcanic rocks along the western Woodlark spreading axis range from normal mid- ocean ridge basalts to basaltic andesites with back-arc geochemical affinities. Rare peralkaline rhyolites also occur. Hydrothermal deposits occur at a basaltic andesite volcano located on the neovolcanic zone near its westernmost propagating tip. Franklin Seamount is a 250m-high cone constructed largely of pillow lavas and talus which sits astride a 450m-high ridge of similar lavas oriented subnormal to the spreading direction. At its crest, 2150m below sea level, there is a 100m-deep caldera.

Jan 1, 1993

By Edwin A. Blackey, Joseph G. Engels, James T. Cahill

An extensive geotechnical instrumentation program was undertaken during the construction of a 6.7 m (22 ft) inside diameter tunnel to monitor the ground-structure response and to check assumptions made for design. The tunnel was advanced through sedimentary and igneous rock using a fully shielded tunnel boring machine (TBM) with both temporary support and final lining provided by precast concrete liner plates erected in the tail section of the TBM. The parameters measured and evaluated include rock deformation, strain in the precast concrete liner, changes in tunnel cross section and groundwater response. In addition, an automated camera system is described which was mounted on the TBM to provide a continuous photographic record of the bedrock conditions exposed at the tunnel springline. The performance of various types of instrumentation equipment in an underground construction environment is also evaluated.

Jan 1, 1981

THERE is probably no phase of basic open-hearth steelmaking that is more of an art and less of a science than mold and pouring practice. It varies widely from plant to plant; consequently it is obviously impossible to set down in detail in this chapter the many ramifications of practice that are now used. The discussion that follows therefore attempts only to give an outline of the more common methods, with examples sufficient to indicate some of the variations in procedure, and to emphasize a few of the many underlying problems that face the open-hearth operator at the pouring platform. MOLD DESIGN Ingot molds are hollow castings of considerable wall thickness. Despite this, non-uniformity in wall thickness should be strictly avoided to prevent the setting up of unnecessary stresses. Molds are cast in various forms, sizes, and shapes depending on the type of steel, and on whether the ingot is to be withdrawn from the mold or the mold from the ingot. The selection of ingot-mold material, which is usually cupola or blast-furnace iron containing approximately 3.5 per cent carbon, 1 per cent silicon, 0.9 per cent manganese, 0.2 per cent phosphorus, and 0.07 per cent sulphur, is governed by numerous considerations that affect mold life. (1) Factors in addition to wall thickness that should be given primary consideration at the foundry are casting skin, shrinkage stresses, and casting temperature. Factors in Mold Design. In use, mold thickness is of greatest importance, for it must be heavy enough to freeze the molten metal rising in the mold and to reduce the temperature of the steel below that of the melting point of cast iron to prevent the inner surface of the mold from melting, which would result in the welding of the ingot to the mold. The skin of an ingot, which solidities immediately upon contact with the relatively cold walls of the mold, must soon become thick enough to contract on the fluid mass and pull away slightly from the sides of the mold. The separation of the ingot from the mold depends entirely on the heat taken up by the mold from the steel.

Jan 1, 1944

By E. L. Warburton

Before entering into a discussion of wet washing, it will be advisable to consider the various phases and conditions sur- rounding the problem of cleaning coal from the Corbin mines. Corbin is located in the south-east part of British Columbia in the Crowsnest Pass district, about 35 miles north of the International boundary and about three miles west of the Flat-head Range of the Rocky Mountains. It is reached by a rail-way, about thirteen miles long, from McGillivray, B.C., a point on the Crowsnest branch of the Canadian Pacific rail-way, owned and operated by the Eastern British Columbia Railway Company, a subsidiary company of Corbin Coals, Limited. The coal is assumed to occur here, as principally elsewhere in the Rockies, in the Kootenai series of the Lower Cretaceous, but owing to the disturbed condition of the surrounding country and, therefore, the magnitude of the geological study involved, this has never been definitely worked out. The coal is of the usual Crowsnest high-grade steam-coal, containing from 9 to 12 per cent. of ash, but appears to have 4 to 5 per cent. less volatile combustible matter than other coals being mined in this district, a representative analysis being about as follows:

Jan 1, 1924

By A. J. Jacobs, G. Katz

MANY of the studies of intermetallic compounds containing uranium were performed during the early 1940's under the aegis of the Manhattan Project. Subsequently, much of this work was declassified and published after 1949. In some instances this early work, completed with relatively impure materials, represents the only source of X-ray diffraction powder data available in the literature. In the case of the intermetallic compound U,Fe some discrepancies in the powder pattern exist in the literature. Since this compound has been prepared in this laboratory, it is believed that the publication of the indexed powder pattern would help clarify the conflict in the literature. In studies of the U-Fe system, Gordon and Kauf-mannl and Grogan and Clews2 independently reported the intermetallic compound U,Fe in 1950. The crystal structure was described by Baenziger, Rundle, Snow, and Wilson3 in 1950. It is isostructural with U,Mn, U,Ni, and U,Co and crystallizes in a body-centered tetragonal lattice. The lattice constants determined by Baenziger et al. are: giving an X-ray density = 17.7 g cm-3 for 4 mol per unit cell. The compound used in this study was prepared by melting a stoichiometric mixture of the constituent metals in an argon-arc furnace using a nonconsum-mable tungsten electrode. Before the argon of 99.98 pct purity was introduced into the furnace, it was passed through CaC1, to remove water vapor and afterwards through a stainless steel tube containing calcium turnings heated at 600°C for the removal of nitrogen and oxygen. The purity of the uranium used was better than 99.9 pct and that of the iron was 99.97 pct. Before and during melting a zirconium getter was used to further purify the argon atmos-

Jan 1, 1962

By Salah Badr

Longwall mine layouts with their entries, chain pillars, face support systems, advancing mining faces and compacting gobs represent a geomechanically complex mining operation. Geomechanical aspects are further complicated if mining occurs at depths exceeding 350 m, as stresses induced adjacent to the excavations exceed the unconfined strength of the coal and result in fracturing of the sidewalls of entries and chain pillars during development. Traditional methods of geomechanical analysis, whether empirical or numerical, are unable to represent such behavior satisfactorily. Achieving some meaningful results requires explicit representation of the three-dimensional boundary conditions and implementation of appropriate constitutive models for the rock mass. This paper describes a numerical modeling methodology for deep longwall mines operation using the FLAC3D finite difference de, which incorporates a strain softening constitutive law for modeling failure of coal. An important aspect of the modeling method includes accounting for stress relief provided to the pillars and entries due to the increasing load applied to the compacting gob.

Jan 1, 2003

By Bruce W. Cavender

During the last few years, the minerals industry has been confronted by increasingly stringent environmental regulations that are significantly altering the economics of the industry. Heightened media attention, threats of litigation and decreasing support by the general public have exacerbated the industry's difficulties. Regardless of their merit, environmental issues present a serious threat to the continued viability of mining companies. Firms have reacted to these regulations by escalating their environmental activities to meet the new requirements. They have simultaneously adjusted their corporate schedules to reflect the substantially longer time periods required to obtain construction and operating permits and ensure compliance with legal requirements before mine abandonment. Companies have also been forced to adapt to a considerably greater amount of uncertainty and expense associated with the development of mineral projects. This has resulted in a significant erosion of the profitability of some mining companies. Most of the changes in corporate environmental activities have been reactive. Pending regulations present formidable but uncertain technical requirements for compliance, and future regulatory developments are difficult to predict. This problem is aggravated by depressed market prices for many mineral commodities. Past expansions of environmental standards have led to the closure of mines and processing facilities due to economic hardship, as occurred in the smelting segment of the US copper industry in the 1980s (Cro¬zier, 1992).

Jan 1, 1994

By Hannweg L. A.

In a first for the diamond industry, Koffiefontein Mine in the Free State, South Africa, opted for the front cave mining method to extract a 90 metre high block of kimberlite ore. Block caving, because of its high initial capital requirements, resulting in a negative internal rate of return (IRR),was ruled out as a feasible mining method for Koffiefontein Mine. Front caving is essentially a combination of the traditional block cave and sub-level cave mining methods, but uses two or more levels of retreating drawpoints. Two extraction levels, in the case of Koffiefontein Mine, are spaced 12 m apart vertically, with the upper 478 m level (48Level) acting as the undercut/extraction level, and the lower 490 m level (49 Level) as the extraction level. Caving is achieved by undercutting a small column of ore, which is then drawn, on both the 48 and 49 Levels, until the planned extraction percentage is achieved. The next column of ore is then mined by advancing the undercut (48 Level), and the process is repeated until the entire ore block is depleted to the contact. Caving-by-gravity techniques of mining rely on closely controlled draw of the caved ore to ensure proper caving action and to minimize the dilution of ore with waste material from the open pit. This paper discusses the need for, and implementation of, a draw control system at Koffiefontein Mine — a computerized production management system, and a draw control system based on linear programming, to allow control over the draw rate and mining sequence — a major factor in all production planning to optimize production efficiency, ore recovery and cave management. Keywords: Draw control, Koffiefontein Mine, Front cave, Caving

Sep 1, 2001

By R. A. Anderson, T. E. Tietz, J. E. Dorn

Recent investigations1,2,3,4 have conclusively shown that the strain hardened state of metals depends upon the temperature and strain rate of pre-straining as well as on the total plastic strain. A typical example of the effect of temperature of prestraining on the work hardening of metals is reproduced in Fig 1. When pure aluminum is strained at liquid nitrogen temperature the upper solid stress-strain curve (OFGH) is obtained. But when pure aluminum is prestrained to e2 = 0.153 at atmospheric temperature and the test is then completed at liquid nitrogen temperature, the curve OBCD results. Thus a prestrain E2 = 0.153 at atmospheric temperature gives a flow stress C when continued at liquid nitrogen temperature, whereas a higher flow stress G results when the specimen is strained exclusively at liquid nitrogen temperature to the same strain E2. Consequently the metal strain hardens more rapidly at the lower temperature. Obviously the strain alone is not a measure of the strain hardened state in metals. Inspection of Fig 1 reveals that the flow stress at C is identical with that at F. This fact suggests that a strain E2 = 0.153 at atmospheric temperature induces the same work hardened state as a strain E1 = 0.058 at liquid nitrogen temperature. But this concept of equivalent strains is also in error. If curve CD is shifted to the left so that point C falls on F, point D becomes D'. Inspection of the curves reveals that the rate of increase of stress with strain at C is greater than the rate of increase of stress with strain at F even though the flow stresses are identical. Consequently C and F do not represent identical work hardened states. It therefore becomes important to ascer- tain how the strain hardened states at C and F might differ from each other. In a more recent report1 it was shown that straining at higher rates of strain increases the amount of strain hardening in a manner quite analogous to straining at lower temperatures. These observations suggested that the flow stress is not only a function of the strain but also a function of the strain rate-temperature history of straining, perhaps in accord with the Zener- ?H Hollomon parameter p = Ee /RT (where

Jan 1, 1950

By Stewart Gillies AD

Resource rent taxes are taxes which only become effective at such a time as a pre-determined rate of return has been achieved by an operation. Resource Rent Taxes have from time to time been considered for implementation in both Australia and New Zealand. The offshore petroleum industry in Australia is already subject to RRT as is the Roxby Downs (Olympic Dam) Mine in South Australia. Both the mining and petroleum taxation regimes in Papua New Guinea include a resource rent tax component called Additional Profits Tax.

Jan 1, 1995

By Roswell Johnson

To know the rate of decline of oil wells is very important, yet ordinarily we are prevented from getting this rate because the oil from several wells is put into one or a few tanks as soon as the wells have been connected to the power. The object of doing this is economy of tanks, pipes, labor and accounting. FIG. 1. But we are making a great sacrifice for this small economy. The advantages of the individual records are that we would know better: 1. When the well has ceased to be profitable. 2. When a well has gotten into bad condition. 3. The real record of the yield per acre and the rate of decline of the wells. The lease total is too mixed as to age of the wells for existing records to have much significance. To obtain such individual well records and more accurate logs are the great desiderata of satisfactory management of oil properties. 4. Which wells show the slowest decline-information which is valuable in locating new wells or purchasing adjoining leases.

Jan 6, 1917

By A. A. Bauer, M. S. Farkas, F. A. Rough

An investigation of the d-phase relationships between the uranium-zirconium and uranium-molybdenum systems was conducted. A ternary cut joining U-31.5 at. pct Mo to U-74 at. pct Zr is presented on the basis of differential thermal analysis, metallographic and X-ray diffraction data. HyDothetical isothermal ternary sections, inferred from the determined d-phase relationships, are presented for the uraniztm-zirconium-molybdenum system. A ternary eutectoid, at 454ºC, was found in zirconium-rich alloys. Its composition, based on metallography and electron-beam micro-analysis, is estimated as U-61.7 at. pct Zr-7.1 at. pct Mo. URANIUM binary systems have been studied extensively in the past to supply basic information needed in reactor-fuel development. Alloys of uranium with zirconium, molybdenum, and titanium are of particular interest because, in these binaries, extensive or complete solid solubility between the body-centered-cubic forms of these elements and y uranium exists. In addition, an intermediate phase of limited solid solubility forms in each of these systems by transformation of the high-temperature ?-uranium phase. Previous investigation of the phase relationships existing between the intermediate 6 phases in the uranium-zirconium-titanium and the uranium-molyb- denum-titanium ternaries have already been reported.1,2 The results of an investigation of the phase relationships existing between the 6 phases in the uranium-zirconium-molybdenum system is reported here. The phase equilibria that occur in the solid-state portions of the pertinent binary systems are shown in Fig. 1. Phase nomenclature appearing throughout the report is as shown in these diagrams. The crystal structure of the phases in this system are given in Table I. EXPERIMENTAL PROCEDURES Alloys were cast and fabricated to 1/4-in.-diam rod. Specimens were heat treated and examined by metallographic and X-ray techniques. Thermal-analysis data were obtained.

Jan 1, 1960