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By Harold A. Krueger

FACILITIES and mining operations of the National Lead Co., St. Louis Smelting and Refining Division, near Fredericktown, Mo., are situated in a famous mining area. Copper, lead, nickel, and cobalt have been mined here for more than 100 years, work having been started on a high sulphide copper outcrop in 1847. Lamotte sandstone is characterized by differential compaction on a rigorously eroded pre-Cambrian surface. The Bonneterre formation was therefore a good host for minerals not generally found in mineable quantities in these midwestern areas. Unusually complex minerals, however, make beneficiation difficult, and because of irregular ore thicknesses and elevations many engineers and operators have not attempted to mine the property. Others have tried who failed. This paper deals with economic, efficient, and competitive methods of mining these highly irregular orebodies, as compared to the open-stope, room-and-pillar methods normally used for horizontal-bedded lead deposits. For the purpose of this study it should be understood that the ore is found in two distinctly different types of occurrences, one to be designated as basin ore and the other as contact ore. Mining of basin ore is complicated by many faults, fractures, cross faults, and breaks. Contact ore is complex because it is found on flanks or slopes of pre-Cambrian knobs or highs. The dip of the mining floor for the latter type varies between 18" and 45". Occurrences of both types of ore are complicated by water courses or solution channels which carry unconsolidated shale, lime, sand, and dolomite. This material is also found between the bedding planes of the members of the Bonneterre formation. The water found where there are fractures, faults, and channels makes it very fluid and tacky, see Fig. 1, particularly after it has been blasted and handled by loading and hauling machines. Much of the ore can be wadded and thrown without dispersing. During early operations by the Buckeye Copper Co. in 1861 and the North American Lead Co. from 1900 to 1910, conventional narrow-gage railroad and side dump mine cars were used with hand shoveling. The complications of mining the contact ore, the only type attempted at this time, can be appreciated when it is realized that operators were obliged to use mules for haulage. Haulageways constructed on these slopes were of necessity similar to wagon trails or goat trails up the side of a mountain. In other words, it was merely a matter of going from side to side of the strike length of the slope, gaining a little in elevation on each shuttle trip. Production totaled only one to two tons per manshift. A few years later, about 1913, the property was purchased by combined Canadian interests known as the Missouri Cobalt Co., and the use of trolley locomotives was initiated. Between 1900 and 1928 a land agent using churn and diamond drilling methods prospected scattered sections of the area. In 1928 the first property was purchased by the present company, then operating as the St. Louis Smelting and Refining Co. Check drilling and prospecting was carried out by the company at various times between 1928 and 1939 to correlate the erratic mineralization. Much information about both types of orebodies was accumulated, but it was still questionable as to whether money should be invested to work these occurrences. In anticipation of high lead and copper prices, about the time World War II started, it was decided to develop and bring into production some of this ore. In 1942 No. 1 shaft was put down on the largest basin-type orebody and in 1943 No. 2 shaft was put down on contact-type ore. Operations were expanded when No. 3 shaft was completed in 1943, and progressed further in 1948, when National Lead Co. dewatered and opened No. 5 and 6 mines, old workings of the North American Lead Co. and the Missouri Cobalt Co. Because of the differential compaction of Lamotte sandstone over the pre-Cambrian porphyry, in some instances mineable thicknesses of basin-type ore occurred 20 to 30 ft above the sand. This is the exception rather than the rule, since most of the mineralization starts at the sand and is variable in thickness. The ore was attacked, therefore, by development drifts and crosscuts at the lowest possible elevation, where the ore immediately overlying the Lamotte sandstone could be drained and made accessible for mining. It was planned to connect to the drifts and crosscuts with raises to mine ore deposited 20 to 30 ft higher. The higher orebodies were thus mined as slusher levels. Slusher hoists were used to drag the ore into the raises, which were made into hoppers. The ore was then loaded into 32x32-in. ore cans, hauled to the shaft by battery locomotives, and hoisted by the conventional Tri-State method. The rate of efficiency was 5 to 6 tons per manshift underground. The contact-type ore was attacked in a similar way, except that the orebodies were not nearly so wide, so that they were more flexible for slusher loading into cans. This advantage was offset, however, by haulage complexities, since the railroad was constructed on steep slopes. Through experience and ingenuity, many improvements were made in mining both types of ores. The two levels, so-called, in the basin-type ore-bodies were connected as previously planned, more efficient locomotives replaced the older ones, and a

Jan 1, 1954

By N. J. Grant, B. C. Giessen, Paul Predecki

ALLOYS of gold, silver, and aluminum with elements of the groups BII, BIII, BIV, and BV were prepared by a rapid quenching technique (splat) and were examined by X-ray diffraction. Five new intermediate phases were found and will be described briefly herein. For the gold and silver systems, the concentration ranges having an electron/atom ratio e/a of 1.4 to 1.5 ("3/2 Hume-Rothery phases") were studied primarily. Master alloys were prepared from high-purity metals (99.9+ pct or better) by melting either in evacuated fused silica capsules or by nonconsum-able-electrode arc melting in an argon atmosphere. Small pieces, 20 to 50 mg, of each alloy were blast-atomized to form a splat, by a technique similar to that described by Duwez and Willens.1 The technique used for this study is described in detail in Ref. 2; it utilizes a resistance-heated graphite crucible with a small hole at the bottom, directed toward a metal substrate or quenching plate. The prepared alloy rests over the fine hole, through which it is expelled by an explosion shock wave in the form of fine droplets (1 to 50 µ) of molten metal onto a copper or silver substrate, which is maintained at about -190°C. The resulting very high cooling rates (see Ref. 2 for quantitative measurements) can prevent the process of nuclea-tion and growth in many instances, resulting in the formation of metastable phases. The splat particles were transferred to a GE-XRD5 diffractometer and maintained at -190°C, where they were examined with CuKa radiation. The samples were then allowed to warm to room temperature or were heated to higher temperatures until the equilibrium structures formed. Of fifteen alloy systems considered, nonequi-librium structures were encountered in six; these are described below and summarized in Table I. In the system Au-Sb a metastable £ phase (A3 type, hcp, a = 2.898 + 0.002A; c = 4.731 * 0.004A; c/a = 1.633) was found in the concentration range Au + 13 to 15 at. pct Sb. This phase is isomorphous with the stable phases in the systems Au-Cd, Au-In, and Au-Sn, all at an average e/a ratio of 1.4 to 1.5. The concentration range of one-phase metastable was deduced from the small amounts of supersaturated gold solid-solution phase present in the splat product. It was found that ? could also be retained by splatting onto a substrate held at room temperature: however, decomposed into the equilibrium phases Au + AuSb2 after heating to 200°C for 1/2 hr, or on holding the powdered splatted alloy at 20°C for several months. Calorimetric measurements will be made in an attempt to decide the question whether ? is metastable at all temperatures or whether it is a stable phase at low temperatures. There is evidence that another phase, possibly also close-packed but with a different stacking sequence, can be obtained by rapid quenching of alloys with a different antimony content. Klement, Willens, and Duwez3 reported the existence of an amorphous phase on quenching Au-Si alloys (25 at. pct Si) to - 196°C. They found that on heating to room temperature another phase of unknown crystal structure was formed. This was confirmed (see Table I); however, the new crystalline phase, designated as ?, could also be formed simply by rapid quenching to room temperature, and even was found to exist already in the as-cast Au + 20 at. pct Si alloy. It was found that ? decomposed into Au + Si on the specimen surface at room temperature. This behavior, and the question whether or not there is an equilibrium-temperature region for ?, have not yet been resolved. It is probable that ? (Au + 20 to 21 at. pct Si) is cubic of the -brass type (D81-3) with a = 9.60, + 0.01A and N = 52 atoms per cell [compare 6 (CU-Sn)4]. Except for two very weak lines, the powder pattern of about thirty lines could be indexed on this basis; however, a determination of the atom positions has not yet been attempted. For Au-Ge the C phase was observed at about 21 at. pct Ge as reported by Luo et at.5 Lattice parameters a = 2.876A, c = 4.73,A, c/a = 1.64 were found. In the Au-Pb system, formation of a ? phase was not observed, but in the lead-rich region at 75 at. pct Pb, broad peaks belonging to an amorphous phase were found. The maximum diffracted intensity occurred at 28 = 32.4 deg which is about 1 deg larger than the position of the (111) line of lead (Cuka). For Ag-Pb, an amorphous phase analogous to the one found in the Au-Pb system was observed; this metastable phase exists probably at about 75 at. pct Pb. Since no lead-rich alloys were tested, all alloys consisted of silver + amorphous phase at -190°C. In A1-Ge alloys, line-rich and complex powder patterns were obtained at about 30 at. pct Ge; they bear similarities to those of aluminum and germanium, but are of lower symmetry; the existence of more than one intermediate phase is possible. The authors are grateful to the Kennecott Copper Corp. for Fellowship support, and ARPA (Contract

Jan 1, 1965

By C. W. Arnold, H. L. Stone, D. L. Luffel

The purpose of this research is to determine (I) the efficiency of small banks of enriched gar driven by methane in displacing oil from a porous medium and (2) the effects of variation in bank size and composition of that efficiency. Most of the experiments were conducted in a sand-packed tube 20-ft long and 1/2-in. in diameter. The hydrocarbon system generally used was methane, butane and decane at 2,500 psia and 160°F. The results of these experiments indicate that, in the regions contacted by the gas, a small bank of an oil-miscible gas driven by methane can displace all of the oil in a piston-like manner. If the enriched gas is of such composition as to remain immiscible with the oil, displacement of oil is less efficient than for the miscible case, and the gas bank travels through the sand with a velocity less than that of the driving gas. These data along with theories discussed imply that smaller banks and less total gas are required when the enriched gas and oil are miscible. INTRODUCTION Widespread application of enriched-gas drive to the recovery of oil rests upon a key factor — the use of limited quantities, or "banks", of enriched gas. At the present time, the value of liquefied petroleum gas or other enriching agents discourages their use in a continuous injection technique, or even in a large bank, except in a few isolated reservoirs. If small banks of enriched gas driven by methane were as effective in displacing oil as is continuous injection, the enriched-gas drive process might be applied to a larger number of reservoirs. Previous research on the mechanics of the enriched-gas drive process reported by Stone and Crurnpl and by Kehn, Pyndus and Gaskell has utilized continuous injection of enriched gas. This work has shown that two types of displacements occur. With gases containing sufficient intermediates. the oil is displaced misciblv and complete recovery is obtained from the regions swept. When gases are used which contain insufficient intermediate hydrocarbon for miscible displacement, oil is displaced immiscibly. In the latter type, selective solution of the intermediate hydrocarbons causes a swelling and reduction in viscosity of the oil and leads to an increased recovery over that obtained by dry-gas (methane) drive. The size of the enriched-gas bank necessary for efficient displacement of oil is determined by those factors which cause deterioration of the bank. A differentiation may be made between those factors which operate on a microscopic scale and those which act on a macroscopic scale. On the smaller scale, the enriched gas mixes in the direction of flow by diffusion and convection with the fluids immediately preceding and following it. On the larger scale, the gas may by-pass the oil by flowing through permeable streaks, by overriding the oil because of density difference, or by fingering because of unfavorable viscosity ratios. In such cases, the enriching material tends to mix with the oil both laterally and in the direction of flow. The increase in effective area available for diffusion and dispersion of the enriching components leads to a faster degradation of the bank and a need for a larger bank than is necessary for those cases in which no by-passing occurs. The effects of such macroscopic factors in the deterioration of enriched-gas banks have been reported in a separate paper by Blackwell, Terry and Rayne. The present study was confined to the factors which operate on the smaller scale, in particular to the behavior of banks of enriched gas in sands uniformly swept by the gas. Experiments were designed to answer the following questions. 1. Can small banks of enriched gas driven by methane be used to secure oil recoveries comparable to those obtained by continuous injection of enriched gas? 2. What is the optimum bank size (the minimum bank size necessary to obtain a recovery comparable to that obtained by continuous injection of the same enriched gas)? 3. How many total pore volumes of gas must be injected to obtain the maximum recovery when the optimum bank size is used? 4. What is the effect of varying the number of enriching components in a gas bank? This report describes the experimental investigations and discusses the results in terms of their significance to reservoir behavior.

By Y Gu

The Mineral Liberation Analyser (MLA) employs the technologies of scanning electron microscopy (SEM) and energy dispersive X-ray analysis (EDX). The use of backscattered electron images to define phase boundaries produces very high resolution image analysis and subsequently detailed and accurate particle images and data output. The mineralogy, elemental deportment, grain size, liberation characteristics and mineral associations of a large number of samples can be described. The MLA uses a comprehensive set of analysis methods to meet a wide range of applications. The data from ore characteristion projects completed during feasibility studies are important aids for effective plant design. The MLA can be used to establish variation in the mineralogy of an orebody, identify and quantify the deportment of deleterious elements, and provide valuable information to support comminution and flotation test work. In established plants the MLA data is an important component of plant survey work and monthly composite samples. Changes in feed characteristics and/or operating conditions can be monitored and solutions to any deterioration in recovery and grade performance found. The MLA has extensive applications in the minerals industry. Rare and precious metal samples can be accurately measured using high-resolution grey-scale images. Several new measurement techniques have been developed to perform rapid searches of a large number of samples to provide efficient and accurate analyses of fine-grain, low concentrate material. Quantifying the mineral variability in mineral sands operations is essential to maximising grades and recovery. The MLA produces particle-based data on density, mineralogy, size and shape and has developed measurement techniques to provide the elemental quantification required for the characterisation of these deposits. Nickel sulfide ores can be analysed to provide modal mineralogy, elemental deportment and liberation and locking data. The low-grade and variable nickel mineralogy of these deposits makes the use of quantitative mineralogy an essential component of maximising returns. Coarse particle measurements preserve the in situ textures of ores and support flotation and breakage test work for feasibility studies. The deportment of penalty and bonus elements can be quantified and the association and grain-size data can be used to identify potential separation issues.

Jan 1, 2006

By W. W. Anderson, D. G. Girton

The photoluminescence of Pr, Nd, Ho, Er, Tm, and Yb in CdS, and Ho, Er, Tm, and Yb in ZnSe has been observed from crystals Prepared by diffusion using rare earth metals and an excess chalcogen pressure. For a given temperature, time, and chalcogen pressure the spectral characteristics were very reproducible from run to run, and the emission intensity for Nd, Er, and Yb in CdS was as high or higher than the best vapor phase doped crystals we have grown. For a few rare earths it was found that certain conditions of diffusion tend to yield optimum rare earth emission intensity with respect to the background lattice emission. Photoluminescence measwements of Yb in CdS as a function of depth gave a profile which was neither a Gaussian nor complementary error function. Part of the profile appears to arise from a fast component of the diffusion and the other part from a slow diffusing component. At 960°C and 33 atm S pressure, a com -plimentary error function approximation of the slow diffusing component gave a diffusion coefficient of D = 1.3 x 10-9 sq cm per sec. MOST of the studies of emission from rare earth ions in II-VI compounds have been reported on crystals doped during growth,1,2 although Kingsley and Aven prepared ZnSe:Er by diffusion for paramagnetic resonance and fluorescence studies. Pappalardo and Dietz prepared CdS:Yb by diffusion, but they made optical absorption measurements., We know of no study on the properties of rare earth diffusion in the II-VI compounds. To date we have diffused Pr, Nd, Ho, Er, Tm, and Yb into CdS, and Ho, Er, Tm, and Yb into ZnSe and observed the rare earth emission spectra. For a given temperature and chalcogen pressure, the emission characteristics are very reproducible from run to run and for Yb, Nd, and Er in CdS, as good as the best crystals we had prepared by doping during vapor phase growth.2 The emission of Pr, Ho, and Tm has been observed in CdS prepared by diffusion for the first time. Previous attempts2 to prepare these later three materials by vapor phase growth were unsuccessful. The problem of obtaining reproducible characteristics in II-VI semiconductor compound work is well known.5 Not only is it difficult to reproduce results from one laboratory to another but it is sometimes difficult to reproduce results from one growth run to another under ostensibly identical conditions within one laboratory. This situation has been particularly bothersome in research on the luminescence of rare earth activated ZnS1 and Cds2. Crystals from one vapor phase growth run would show very strong rare earth line emission while crystals from a nearly identical run would show no rare earth emission. It was also observed on occasion that the intensity of the rare earth emission was not constant over the entire volume of a single crystal. MATERIAL PREPARATION AND INSTRUMENTATION Vapor phase grown boules of CdS were supplied by Dow Corning. This material was characterized by a free electron concentration of n - 3.5 x 1015 cm-3 and Hall mobility of 350 sq cm per v sec at room temperature. There were microscopic voids and decorated precipitates in some samples. The precipitates annealed out at diffusion temperatures but the voids remained. Single crystal rectangular samples of mm dimensions were sawed from the boules. The ZnSe was polycrystalline, UHP grade from Eagle-Picher. Poly-crystalline samples were sawed from the ingots. The samples were lapped, polished on one side, etched in a solution of 0.5 M K2Cr2O7 in 16 N H2SO4, and thoroughly washed in distilled water. A sample, excess sulfur (or selenium), and 5 mg of rare earth metal (turnings) were sealed in a 3.6 cm3 quartz ampoule at about 2 X 10-5 torr. The high chalcogen pressure used (1 to 30 atm) prevented thermal etching of the crystals and affected the diffusivity and solubility of the rare earth ions in the crystal lattice. For meaningful or reproducible results, it is thus necessary to specify the vapor pressure at which the diffusion was carried out. It is assumed that a negligible amount of the chalcogen was used in the formation of rare earth sulfides or selenides Our sulfur vapor pressure calculations are based on data assuming S2, S6, and S, molecules only in which case the equilibrium constants are given by6 where the pressures are expressed in torr. Selenium vapor consists of a mixture of Se2, Se4, Se6, and Se6 molecules. The selenium vapor pressure was calculated using equilibrium constants given by The status of the rare earth source during diffusion is unknown, i.e., the partial pressures of the rare earth metal and of the rare earth chalcogenides has not been determined. All emission spectra were recorded at 77°K on a Perkin-Elmer model 98-G spectrometer using a 640 line per mm grating. No correction was made for the spectrometer and detector spectral sensitivity. Excitation was by means of an XBO 1600 w xenon arc

Jan 1, 1970

By M. M. Wong, D. E. Couch, G. M. Martinez

The Bureau of Mines electrowon hafnium metal with an average oxygen content of' 150 ppm at 700°C from an electrolyte containing 27 wt pct LiCl, 62 wt pct RbCl, and 11 wt pct HfC14. The average anode and cathode current efficiencies were 90 pct at anode and initial cathode current densities of 86 amp per sq ft. Haf-nium metal with an average oxygen content of 440 ppm was electrowon at 800oC from an electrolyte containing 90 wt pct KC1 and 10 wt pct HfCl4. The average anode and cathode current efficiencies were similar to those obtained in the LiCL-RbCl-HfCl, electrolyte. The chlorine gas given off at the graphite anode was vented through either a silica or a graphite tube to prevent cell corrosion. THE current method for the commercial production of high-purity hafnium is the thermal decomposition of Hfl4.1 The iodide method is not adaptable to continuous process techniques. Nettle, Hiegel, and Baker2 studied the electrorefining of hafnium from hafnium sponge containing 800 ppm oxygen. They failed to obtain hafnium with 600 ppm oxygen in their initial deposits and obtained AEC specification for oxygen only after 75 pct of the soluble hafnium had been removed from the electrolyte. Calculations using their data indicated this was approximately 4 lb of hafnium. The electrolyte was then used to produce approximately 3 lb of hafnium with a low oxygen content. However, no data are shown concerning the amount of anode material initially used or what percent of it was dissolved, therefore, results are not suitable for evaluation of a continuous operation. In general, it was not possible to consistently obtain low oxygen content metal with the electrolytes described by Nettle, Hiegel, and Baker. Wong, Hiegel, and Martinez3 investigated the electrorefining process for hafnium and showed that even by strict control of electrolyte composition only relatively low oxygen reduction could be obtained. The oxygen contained in the hafnium anode material tended to transfer to the cathode deposit and only a limited purification was possible. Both the "iodide" and the "electrorefining" processes depend upon hafnium sponge as a starting material. The sponge is normally produced by magnesium reduction of HfC14 ' and does not meet AEC specifications for hafnium metal. Since only 30 pct of the anode feed could be utilized3 in the electrorefining cells, the Bureau of Mines developed an electrowinning process. HfC14 was used as the feed material for the electro-winning process described in this report. Many of the electrolytes used in the electrorefining studies3 ap- peared to be suitable carrier-electrolytes for HfC14. However, in the initial studies on electrowinning, it was desirable to use electrolytes that had low solidus temperatures and could be operated over a wide temperature range to investigate parameters of the process. Therefore, electrolytes containing LiC1, NaC1, KC1, RbC1, CsC1, and HfC14, in various combinations were explored. EQUIPMENT Chlorinator. Hafnium carbide was chlorinated to produce HfC14 in the batch-type chlorination shown in Fig. 1. Chlorination temperatures were measured with a thermocouple placed in the center of the HfC charge. A flow meter was used to monitor the helium and chlorine. The exhaust side of the silica chlorina-tor tube was equipped with a flask for collecting organic material released during the initial heating of the HfC. The temperature of an internal heater, which extended from the HfC14 condensing flask to the hot end of the chlorinator, was adjusted to prevent the HfC14 from condensing before entering the collection flask. Helium and excess chlorine were exhausted through the lid of the collection flask to an aqueous NaOH solution. Sublimer. Initial studies were conducted using a sublimer, Fig. 2, made by placing a 13/8-in. OD nickel thimble 11 in. long, inside a 11/2-in. ID nickel bell 12 in. long, and locking it in place. This unit was loaded with HfC14 and partially immersed in the molten electrolyte for sublimation directly into the electrolyte. In another sublimer shown in Fig. 3, the HfC14 was contained in a "resin reaction flask". Quartz wool, previously heated to 600aC, secured between two nickel wire screens, was placed just above the HfC14 powder. The lid contained a vacuum outlet, a gage, an argon inlet, and an air-cooled pipe for condensing the HfC14. This sublimer was evacuated and heated. The sublimation temperature was not critical and the sublimer operated satisfactorily at all temperatures between 250" and 350°C. Electrolytic Cell. The electrolyte chamber, Fig. 4, was made of mild steel 8-in. schedule 20 pipe, 30 in. long. The exterior was metallized with a Ni-Cr alloy. The electrolytes were contained in a 16 gage nickel or iron liner with a nickel heat shield on top. The cell was heated by a resistance furnace. A 21/2-in. ID by 25 in. long air lock was connected to one port of a two-port cell cover assembly through a slide valve. The cover assembly of the air lock was electrically insulated from the cell and was equipped with a rubber sleeve that provided for the passage of the cathode lead. This allowed the cathode deposits to be removed and a new nickel cathode to be introduced without allowing air to enter the cell. A tube-rod assembly was bolted to the other port on the cell cover assembly and was sealed by a packing seal. The tube-rod assembly consists of a graphite

Jan 1, 1970

By Jun Lu, Min Chen, Wei-jian Su, Jian-zhong Luo, Yu-tao Lei

"In the case of treating phosphatizing wastewater containing zinc, zinc had an impact on the solution chemistry, especially on pH. Fe2(SO4)3 and CaCl2 were used as precipitators to remove phosphate from zincic phosphatizing wastewater and the optimum pH was investigated. Zinc ion had a significant effect on the optimum pH. The optimum pH ranges were 12.10(inlet concentration of 15.67mg/L, outlet concentration of 9.53mg/L), 3.04-5.50(inlet concentration of 11.97mg/L, outlet concentration of 0.44-0.26mg/L), and more than 10.72(inlet concentration of 12.24mg/L, outlet concentration of less than 0.47mg/L) under the conditions without precipitator, with Fe2(SO4)3 and with CaCl2 respectively.Introduction Phosphate coating is commonly used for protective coating of machine elements. Including oil removal by caustic washing, rust removal by acid washing, phosphatization and inactivation, the phosphatization process produces wastewater containing inorganic salt of high concentration, especially the phosphate, beyond the environmental standards and polluting environment seriously. What’s more, the wastewater from zincic phosphatizing technology also contains another pollutant, zinc. At present, Chemical Precipitation (CP) and Biological Nutrient Removal (BNR) are two of most common technologies of phosphate removal. BNR, such as A/O, A2/O and UCT technology, is suited for wastewater containing organ phosphorus of low concentration and CP for wastewater with phosphate of high concentration [1]. Researches on CP treating wastewater containing phosphate of high concentration, mainly focus on adding quantity of precipitators[2-5] and coagulants[2-4] and the choice of technologies[3-9]."

Jan 1, 2008

By Frank G. Horino, V. E. Hooker

INTRODUCTION The determination and use of mechanical properties of rock in engineering and rock mechanics are rapidly developing. Many of these properties are determined on intact rock specimens; thus, their application and repre¬sentation of rock mass properties may be limited. How¬ever, relative information often provides useful guidance in the solution to mine design and stability problems. Summarized in this chapter are some of the stan¬dardized techniques and procedures currently used to obtain these mechanical properties. Typical applications of the use of these properties are also presented. Stan¬dardized techniques include those advanced by the American Society of Testing and Materials (ASTM), International Society of Rock Mechanics (ISRM), US Bureau of Mines (USBM), Canadian Dept. of Mines and Technical Surveys, South African Institute of Min¬ing and Metallurgy, and other individual investigators. Information on the mechanical properties of rock and the behavior of the rock under a given system of stresses represents a necessary part of the information for rational engineering design for any given mining op¬eration. The mining method, the type and extent of sup¬port, the extraction ratio, the overall dimensions of the mine, and the orientation of the rooms and pillars are all decisions that are influenced by the mechanical prop¬erties of the ore, roof, and floor material under various stress systems and the magnitude and direction of the in situ stresses (Hooker, Bickel, and Aggson, 1972). Initial mechanical property information regarding a structure or mine property is generally obtained by two basic techniques: (1) static and dynamic property tests are conducted on intact and fractured rock specimens of exploratory drill core, and (2) dynamic properties are obtained by borehole logging techniques. When mining access becomes available, and as the mining horizon is expanded, additional information can be ob¬tained to verify preliminary mine design values. This chapter presents some of the standardized tech¬niques and equipment currently used in obtaining me¬chanical property data in the laboratory. The properties considered are: (1) uniaxial compressive strength of intact rock core specimens, (2) uniaxial compressive strength of rock cores containing planes of weakness, (3) triaxial compressive strength of intact rock core specimens, (4) triaxial compressive strength of cores with a plane of weakness, (5) Young's modulus, (6) Poisson's ratio, (7) density or apparent specific gravity, (8) modulus of rupture, (9) indirect tensile strength, and (10) creep characteristics. Where possible, an at¬tempt will be made to evaluate each property measure¬ment in relation to the problems of rock mechanics and application of results. TEST SPECIMENS The selection and care of drill core for laboratory testing require some consideration. It is recognized that laboratory-determined properties are not necessarily rep¬resentative of an in situ rock mass property. However, relative information between beds or zones of interest is still valuable information in selecting mining horizons and preliminary design criteria. To provide statistical data the number of drill core samples selected to repre¬sent each of the areas of interest should be from a mini¬mum of three to a maximum of ten test specimens. A judgment must also be made on site as to whether the recovered drill core should be wrapped and sealed in plastic to preserve moisture. On the one hand investiga¬tions of air-dried and saturated specimens have shown that moisture significantly affects the elastic properties and strengths of many rock materials (Obert, Windes, and Duvall, 1946; Colback and Wiid, 1965); on the other hand it is apparent that most core drilling is done with water which may saturate the specimen to a greater extent than in the in-situ condition. Whether or not the decision is made to retain the moisture, the core should be delivered to the laboratory as soon as possible after recovery for subsequent specimen preparation and testing. Specifications Shape: The shape of the specimens influences lab¬oratory testing in two ways: (1) time and cost of sam¬ple preparation and (2) strength of the material. Cy¬lindrical specimens of drill core are by far the least time-consuming to prepare for static or dynamic labora¬tory testing. In addition, the cylindrical shape lends it¬self to a more uniform stress distribution throughout the sample than other shapes, such as rectangles and hexa¬gons. The compressive strengths of various shapes have been studied (Grosvenor, 1963, and Price, 1960), and results indicate that the cylindrical specimens usually provide the highest strength for a given height-diameter ratio. However, reduction in strength from a cylindrical shape to a rectangular in situ pillar is not regarded as significant in relation to other considerations such as planes of weakness in a pillar or safety factors in the design process. Length-Diameter Ratio: The length-to-diameter ra¬tio, LID, has a significant effect on the compressive strength. Various recommendations have been made to use standard LID ratios ranging from 2 to 2.5 to 3 (ASTM, 1975c; ISRM, 1972). However, past work by others such as Obert, Windes, and Duvall (1946) has shown that excellent results can be obtained using LID ratios from 2 > (LID) > >/s. In selecting an LID ratio for testing, one should keep in mind the amount of material available for testing. In many instances, this may be limited. Thus, a shorter specimen such as 1: 1 LID may be necessary to provide enough test data for statistical analysis of results. Sec¬ond, it may be desirable to obtain elastic constants dur¬ing the test. This generally requires instrumentation such as linear variable differential transformers (LVDTs) or strain gages near the center of the specimen. In this case, an LID of 2.5 or 3 is desirable so that the instru

Jan 1, 1982

By R. V. Sara

Determination of the Hf-C phase diagram was conducted primarily by metallographic and X-ray diffraction studies on appropriate alloys. The only intermediate phase observed in this binary system was HfC. This phase was found to be homogeneous between 34.0 and 48.0 at. pct C at 2200°c and between 36.0 and 49.3 at. pct C at 3150°C. The lattice-parameter variation was also determined for HfCI-, compositions prepared at 2200° and 3150°C. The most most refractory composition, with a melting point of 3830°c, was established at 47.5 at. pct C from melting-point data. Solidus temperatures of 2240' and 3150°C occur on the high-kafnium and high-carbon sides of the monocarbide, respectively. The invariant point between HfC and carbon is located at 66.0 at. pct C, whereas the 2240°C solidus corresponds to the peritectic temperature at which hafnium is formed from HfC and hafnium-rick liquid. Hafnium has a melting temperature of 2208°C and is capable of taking carbon into solution to the extent of 10.5 at. pct at this temperature. ALTHOUGH the Hf-C phase diagram has not been previously evaluated experimentally in its entirety, the belief has been that the general configuration would resemble the chemically similar Group rV carbide systems, Ti-C and Zr-c.' These binaries are characterized by a single carbide phase with a simple NaC1-type structure which is maintained over wide compositional ranges. This family of carbides has high thermal stability that increases substantially as the atomic number of the metal component increases. This trend characterizes HfC as one of the highest melting materials. According to Agte and Alterthum,2 the melting point for this monocarbide is 3890°C, a value that has been quoted quite extensively for the past several decades. Recently, in repeating the work of Agte and Alter-thum, Adams and Beall3 determined the melting point of HfC to be 3895°C. A significant departure from the commonly accepted version of the Hf-C system was reported by Avarbe and his coworkers,4 who proposed that there is an extreme stabilization of a hafnium in a narrow field to 2820°C, above which it melts peritectically to form HfC and liquid. Their study was also concerned with the melting temperature of various HfCI-, compositions, but the peak melting point was taken from the work of Agte and Alterthum. Avarbe and his associates were not concerned with the high-carbon region of the system. However, three widely varying temperatures have been reported for the HfC-C solidus by other investigators. Cotter and Kohn5 observed approximately 2800°C; Portnoy et al.' reported 3260°C; and, more recently, Krikorian7 indicated 2915°C. Equally as uncertain is the solidus between hafnium and HfC. As noted above, Avarbe et al. report a peritectic temperature of 2820°C, Krikorian7 measured 2150°C, and Benesovsky and Rudy' estimated 2000°C on their diagram. I) EXPERIMENTAL PROCEDURE The starting materials for this study consisted of reactor-grade hafnium hydride obtained from Fair-mount Chemical Co., Newark, N.J., hafnium carbide supplied by Wah Chang Corp., Albany, Ore., and Union Carbide spectroscopic-grade graphite, SP-1. The graphite analysis indicated impurities at levels of only 0.5 ppm or less. According to the suppliers, the hydride and carbide contained the typical impurities listed in Table I. The samples required for these studies were prepared by dry blending either graphite and hafnium hydride or hafnium hydride and hafnium carbide powders for approximately 5 min in a "Spex Mixer Mill". The latter combination was used only for preparing several samples in the HfC1-, melting-point studies. Small pellets, varying between 3/16

Jan 1, 1965

By M. I. Abbas, M. A. Shahin, M. S. Ahmed

"The influence of annealing of galvanized steel (galvannealing) on its corrosion kinetics in a saturated solution of Ca (OH)2 has been studied by means of the polarization resistance method.Galvannealing heat treatment was carried out at 450°C and 550°C for varying times. Annealing at 450°C for 24 hr or at 550°C for 5 hr showed the highest corrosion potential (Ecorr) ~= - 400 m V/ SCE as well as, the lowest corrosion current density (icorr ~= 3 mA.cm-2, while Ecorr and icorr of the unheated specimens were ~= - 1400m V/SCE and == 20 mA.cm-2, respectively.Polarization curves were also obtained for the as-received galvanized specimens together with the galvannealed ones.INTRODUCTIONHeat treatment of galvanized steel (i.e . galvannealing) is carried out at a temperature range of 450°C to 550°C for a time sufficient to convert the free zinc coating to an alloyed one consisting of layers of iron-zinc intermetallic compounds. Table I is a review of the properties of the Fe-Zn intermetallicphases while Figure I is a reproduction - of the Fe-Zn equilibrium phase diagram taken from Hansen.Several authors have studied the formation of intermetallic phases through the interdiffusion of Zn/Fe couples. Their results confirmed the stability of intermetallic phases that could be shown up on the outer surface as a result of galvannealing of galvanized steel.Bernick and Sievertv"" investigated the processing of galvan nealed steel. They found that the galvannealed coating with the best forming qualities consisted of a mixture of Zeta phase and Eta phase."

Jan 1, 1989

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit? The rocks are well-consolidated Gila conglomerate, quartz , monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ransome.3 In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special, fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3 1/4x4 1/4-in. tool joints; 8-in. bits, 2 3/4x 3 3/4-in. tool joints; 6-in. bits, 2 1/4 x3 l/4 -in. tool joints; 4-in. bits, 1 5/8 x2 5/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property, To allow for 2 3/8-in. drill pipe with tool joints, 3 1/2-in. core barrels and bits were used. With the standard 3 1/2-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 5/8 to 3 7/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven-machines. The open churn drill hole was cased with 2 1/2-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in unconsolidated material like the formations drilled by

Jan 1, 1952

By K-H. Frohne

This report presents a study of the Cooper sand of the Fork Run pool of Warren and McKean Counties, Pa., and a prediction of the secondary oil recovery by gasflooding the reservoir. A computer program employing a modification of the Craig, Geffen, and Morse secondary recovery calculation was used to make the gas flood performance forecast and is presented in the appendix. The Cooper sand reservoir in the Fork Run pool lies at a depth of about 1,900 feet and covers an area 1/2 mile wide by 2-1/4 miles long. Original oil in place is estimated to have been 8,266,000 STB. The primary recovery mechanism is solution gas drive, and ultimate primary recovery may reach 20 percent over a well life of 90 years. Two computer predictions were made for gas flooding the Cooper reservoir. The first was for a 25-acre pilot five-spot utilizing existing wells, but project economics were poor since only 25 percent of injection gas is utilized for pilot oil recovery. Recovery was 71,000 STB, or 22 percent of oil in place at flood initiation, but required 25,700 scf gas injected per STB of oil recovered. In the second prediction, pilot area performance was projected for the entire Fork Run pool, and economics became good. Recovery was 1,753,000 STB from 25 five-spots after the injection of only 6,400 scf per STB of oil recovered.

Jan 1, 1969

Australia is the world's leading producer and exporter of mineral sands. Despite increased production from some other countries in recent years, Australia maintained its No.1 position in 1988 among world producers of mineral sands concentrates. Production of the titanium minerals -- ruffle, ilmenite and leucoxene, as well as synthetic ruffle -- by the seven companies currently comprising the Australian industry increased in 1988 in response to continuing strong demand for feedstock by the pigment industry. Production of co-product zircon remained at the 1987 level, reflecting declining zircon grades. Exports of mineral sands concentrates increased in 1988; export earnings exceeded $300million. Prices for mineral sands concentrates are now at a historical high and, assuming that the world economy remains strong, real prices should continue to trend upwards in the short term. Exploration in the Eneabba-Cooljarloo region and the Scott River area, both in WA, the Murray Basin, Vic, and the central Queensland coast area, has identified substantial resources of heavy mineral sands. These discoveries have reversed the trend of the early 1980s of falling resources of mineral sands. In 1988, BMR estimated that economic demonstrated resources (EDR) of ilmenite, rutile, zircon and monazite combined was about 72.8 million tonnes. However, a substantial portion of the resources on the east coast are frozen as a result of environmental considerations. The proportion of mineral sands concentrates processed in Australia increased with the commissioning of two new synthetic rutile plants; Australia is now the world's leading producer of synthetic rutile (total capacity 272000 tonnes), most of it for export. Australia's two TiO2 pigment producers have both increased capacity, lifting national capacity by some 70p ercent to over 120000 tonnes/yr. As part of the expansion one producer converted from the sulphate to the chloride process. Other new pigment plants are being proposed for the future.

Jan 1, 1989

By L. E. Kun, R. O. Oelofsen, E. J. J. van Vuuren

High-rate floc-blanket clarification of gold pregnant rotary filtrate was introduced at the Vaal Reefs South gold plant of the Anglo American Corporation during February 1978. A. daily 12 000 to 15 000 t of filtrate, with an average suspended-solids concentration of 350 mg/I, is continuously clarified to a concentration of less than 2S mg/I of suspended solids in a modified plant pachuca (10 m in diameter) at throughput rates of 7,7 to 8,5 ma/m2.h by the addition of 5 g/t of ferric chloride inorganic coagulant and 0,2 g/t of Superfloc NI00 flocculant. The pachuca overflow is then clarified further in seven Stellar candle pressure filters of 54,S m2 surface area to under 5 mg/I of suspended solids. The combined clarification cost when the hopper clarifier is used in conjunction with the Stellar filters is R0,0032 per ton of liquid treated, which is approximately 60 per cent lower than the cost of Stellar filtration on its own. The savings are mainly due to the reduced use of filter pre-coat, the consumption of which has dropped by 70 per cent since the introduction of hopper clarification.

Jan 1, 1979

By Gordon E. Agar

"The effectiveness of two this compounds for copper sulphide depression in molybdenite flotation has been examined with batch bench scale tests.Both thioglycollate and trithiocarbonate proved to be good depressants producing results similar to those f or cyanide and hydrosulphide. Ferrocyanide was ineffective on the samplematerial tested.The depressing effect of thioglycollate was readily reversed by an addition of xanthate.IntroductionThe separation of molybdenite from copper sulphide minerals has been practised for a long time(1,2). The predominant scheme is to depress the copper sulphides while capitalizing on the ready floatability of molybdenite to effect the separation. There are many recipes for depressing the copper minerals(3), among them the addition of sodium cyanide, particularly if the copper is present primarily as chalcopyrite; some form of sulphide whether as sodium sulphide, sodium hydrosulphide or phosphorous pentasulphide when there is a mixture of copper sulphides such as bornite, chalcocite and chalcopyrite; potassium ferro cyanide may be used when chalcocite is the principal copper sulphide; peroxide or hypochlorite; as well as heating in the presence of lime. There are different mechanisms by which these reagent s depress the copper sulphides. Sodium cyanide and the sulphide-producing compounds apparently reduce the potential of the copper sulphides to the point where they ar e rendered unfloatable by xanthates . Another study(4) showed that after cyanide addition, chalcopyrite was rendered un floatabl e until the REDOX potential (measured with a gold electrode with a saturated calomel reference) had become more positive than -250 mY. If the chalcopyrite pulp was conditioned after the cyanide addition the REDOX potential increased slowly, whereas if air was introduced during the conditioning the potential r ose much more rapidly. The difference in time to reach the desired level of - 250 mV was as much as 20 minutes' 25 minutes being required when air was absent but less than 5 minutes when air was admitted . These observation s demonstrate some of the problems encountered when using cyanide to depress sulphides such as chalcopyrite . The system adjusts to counteract the depressing effect of the additive. To overcome this tendency it is necessary to either add an overdose at the beginning or to more or less continuously add the regulating reagent. In this study cyanide solution was added under automatic control to maintain a specified REDOX potential."

Jan 1, 1984

By C. H. Li, R. J. Stokes

This paper compares the fracture behavior of tungsten rods in three conditions: recrystallized. recovered, and wrought. Notched specimens suhjected to a 50 in.-lb impact load showed ductile-brittle transitions at 700, 4.90°, and 440°C, respectinely. The recrystallized material had an equiaxed pain structure and jracbred by simple cleavage from a grain boundary source at all temperatures up to 700°C. The wrought and recovered material had an elongated fibrous structure and at low temperatures fractured by cleavage originating from the notch. As the transition temperature was approached cleavage was preceeded by more and more intergvanular splitting which deflected the crack front into planes parallel to the tensile axis. The enhanced toughness of wrought and recovered tungsten was attributed both to its inability to initiate cleavage because no pain boundaries were suitably oriented perpendicular to the tensile stress and to its inability to maintain cleavage because of intergranular splitting ahead of the crack. It has been appreciated for a long time in a qualitative manner that the room-temperature brittleness of fully recrystallized tungsten may be alleviated by working the material at relatively low temperatures.' More recently this difference in mechanical behavior between wrought and recrystallized tungsten has been examined quantitatively by measurement of the tensile properties as a function of temperature. In these experiments brittleness has been expressed in terms of ductility or reduction in cross-sectional area upon tensile fracture or in terms of the bend radius before fracture under bending.' This work has shown the existence of a fairly sharp transition from brittle to ductile behavior with an increase in temperature. The ductile-brittle transition temperature for recrystallized material is approximately 200°C higher than for wrought material. An increase in strain rate, small additions of impurity,' or an increase in grain size4 shift the respective transition temperatures to higher values, but the difference between them remains approximately the same at 200°C. A number of explanations for this embrittlement by recrystallization have been given. It has been blamed either on the concentration of impurity at the grain boundaries, the increase in grain size, or the change in texture which occurs upon recrystallization. The present paper examines the effect of different heat treatments on the notch-impact behavior of commercial powder-metallurgy tungsten rods. The change in the ductile-brittle transition temperature for this method of loading and the fracture mode has been related to the different mi-crostructures produced by heat treatment. EXPERIMENTAL PROCEDURE Commercial swaged powder-metallurgy tungsten rods 1-3/8 in. in length and 1/8 in. in diameter were machined to introduce a sharp V notch 0.030 in. deep. To change the microstructure from that of the as-received wrought material some of the specimens were subjected to an anneal in nitrogen either at 1300° or 1400°C for 8 hr or at 1600° or 2000°C for 1/2 hr. The notched rods were then placed in a miniature Charpy-type impact machine and struck at their midpoint (opposite the notch) with a hammer designed to deliver 50 in.-lbs of energy. The strain rate at the base of the notch was estimated to be approximately 100 sec-1 at the instant of impact. The specimens were heated in situ to the desired impact temperature. The microstructures produced by the various anneals were studied by both X-ray diffraction and metallographic techniques. Fig. 1 reproduces the microstructures observed metallographically following a 10-sec electroetch in a 10 pct KOH solution. Fig. l(a) shows the elongated fibrous grain structure of the as-received material. Following the anneal at 1300" or 1400°C the grain structure was still elongated as shown in Fig. l(b) but the etch pits delineated dense polygonized dislocation arrays within many of the grains. Occasionally a relatively dislocation-free recrystallized grain was found growing into the matrix. The anneals at 1600° and 2000°C resulted in complete recrystallization and some grain growth. The grains produced at 1600°C were still slightly elongated as shown in Fig. l(c) whereas the anneal at 2000°C produced equiaxed grains. The changes in grain size produced the expected changes in the X-ray back-reflection patterns; there was no indication either in the as-received material or the annealed material of any preferred orientation. RESULTS a) Impact Behavior. Fig. 2 reproduces the ductile-brittle transition curves measured in the manner described in the previous section. It can be seen that under these testing conditions the as-received

Jan 1, 1964

By B. H. Kear, B. J. Piearcey

The orientation and temperature dependence of the tensile and creep propevties oj Mav-M200 crystals halle been determined. Crystals oriented for single slip exhibit ,maximum ductility, minimum work hardening, and least creep resistance, whereas crystals in multiple slip orientations show nzinimu~n ductility, strongest work hardening, and greatest creep resistance. At 1400" and 1600°F there is a marked improzlenzent in creep strength for orientations approaching [001] and [111], indicating that interactions between dislocations gliding in intersecting octahedral slip systems play an important role in creep resistance. At 1800°F the creep strength is much less dependent on orientation, which is rationalized in terms of slip in cube planes as well as octahedral A frequent mode of failure in high-strength cast nickel-base superalloys is by intergranular fracture, particularly along those grain boundaries oriented transverse to the major stress axis. VerSnyder and ~uard' demonstrated that this problem could be overcome by eliminating transverse grain boundaries through unidirectional solidification. Piearcey and versnyder2 made use of this principle in the development of unidirectionally solidified gas turbine blades and vanes, where the growth direction of the columnar grain structure coincides with the axis of principal stress under operating conditions. The present investigation was undertaken to determine if further improvement in properties may be obtained by eliminating grain boundaries altogether, so as to take advantage of the well-known dependence of mechanical properties of single crystals upon the orientation. 1) EXPERIMENTAL PROCEDURE Single crystals of Mar-M200* were grown from the melt under vacuum by a modified Bridgman method. The melt was poured into a preheated alumina mold, and crystal growth was promoted from one end by appropriate gradient cooling. Tensile and creep specimens (2 in. diam by $ in. gage length) were prepared by a series of operations involving electrical discharge machining, precision grinding, and electropolishing. The orientations of the specimens were determined by the Laue X-ray back-reflection method. Tensile tests were carried out in aWiedemann machine with furnace attachment, using a strain rate planes. In all orientations imzpvouement in the strength charactevistics of the material can be induced by heat treatment. Creep ad stress rupture data for (001) oriented crystals are compared with similar data obtailted previously joy random polycvystalline material, and also columnar grained material having a prejevved (001) orientation. The single-crystal )material exhibits both longer rupture life and lower minimum creep rate at all temperatuves, and the rupture elongation is comparable with that in the columnar grained material. From these results it is concluded that single crystals should and useful application in gas-turbine blades and tanes. The optimum orientation for a blade is considered to be with its axis of principal stvess parallel to (001) ou (111). of 0.0001 sec-'. Load and extension were recorded directly on an X-Y recorder. The strain measuring device consisted of extension arms attached to the grips at one end and leading out of the furnace to an LVDT (linear variable differential transformer) at the other. Creep tests were performed using a standard constant load creep frame. 2) DISCUSSION OF RESULTS 2.1) Structure of Alloy. The structure and segregation in as-cast and heat-treated Mar-M200 has been described in detail else where.~ The main features are as follows: the as-cast material is heavily cored, due to pronounced dendritic segregation during solidification; the dendrites are rich in tungsten and cobalt whereas the interdendritic regions are rich in chromium, titanium, nickel, and carbon. The structure consists of -60 vol pct coherent precipitate of y', basically Ni3(A1,Ti), in a matrix of y (nickel-base solid solution), interspersed in the interdendritic regions with minor MC carbides and -y' eutectic. The y' particles, and y-y' eutectic, contain more titanium, aluminum, and nickel, and less tungsten, cobalt, and chromium than the y matrix. Heat treatment partially removes segregation, eliminates the eutectic, refines the y' dispersion in y, and gives additional partially coherent M23Cs carbide. An as-cast crystal, therefore, is composed primarily of two oriented phases (y + y'), whereas the normally heat-treated crystal consists of three oriented phases (y + y' + MZ3Cs). Typical electron transmission micrographs of the y + y' structure are shown in Fig. 1. Crystals grown in the (100) orientation develop a simple unidirectional dendritic structure, Fig. 2, since (100) happens to be the preferred direction of growth for dendrites in this material. Crystals grown in the {110) and (111) orientations, however, tend to promote equal growth, generally in separate colonies, in the two and three geometrically favored (100) growth directions, respectively. In other orientations of growth, a single (100) dendrite direction generally pre-

Jan 1, 1968

By D. Kim

Kennecott operates one of the largest and most modern copper refineries in the world at Magna, Utah, USA. The inputs to the refinery are anodes produced at Kennecott Smelter, which has extremely high contents of lead and bismuth. In order to minimize undesired effects in the electrolytic refining process, unwanted impurities, especially bismuth, must be controlled. This paper describes the process development of practicing state-of-the-art quality control and presents the pilot test results of using the MRT Technology for Bi production.

Jan 1, 2007

By F. R. Hensel

PuRe silver and silver-lead alloys have been studied as to their suitability for bearing~.' A review of the properties of thallium and the silver-thallium constitutional diagram was made by the author to analyze the possibilities of silver-thallium compositions for antifriction materia1s. t The silver-rich end of the diagram as reported in the literatureg-" was found to be sketchy and it was necessary to carry out corisiderable experimental work to arrive at definite conclusions. Some of the results of this work are reported in this paper. Test Materials In the beginning of the work, only fused alloys were investigated. Later, research Table i.—Composition of Fused Silver-thallium Alloys Tested for Antifriction Properties Percentage Alloy No. of Thallium 1533...................... 0.48 1534.................... 1.06 a31 HT................... 2.04 231....................... 2.10 232 HT................... 3.83 Balance 232....................... 4.07 Silver 233 HT................... 5.86 a33....................... 6.38 a34....................... 9.84 a34 HT................... 10.07 was carried out on electroplating methods and diffusion processes. In preparing the fused alloyst care was taken to exhaust the toxic fumes caused by the thaillium content. The compositions of the series of fused alloys are listed in Table I. The silver-thallium alloys were melted in clay-graphite crucibles and cast into preheated steel molds of 3/4-in, diameter. In the cast condition, they showed a cored structure, as indicated by the micrographs of Figs. I to 3. The etching reagent used was a mixture of 2 grams K2Cr207, 8 C.C. H2S04, and IOO C.C. H20. The cored structure is unsatisfactory for the type of corrosion resistance required for bearing applications, and homogenizing experiments were carried out at various temperatures, the results of which are shown in Figs. 4 through 8. The alloys with a lower thallium concentration, which were heated for 2 hr. at 525°C show an almost completely homogenized solid solution type of structure. With higher thallium concentration, the homogenizing temperature was dropped to 475OC. to eliminate the formation of a liquid phase. It is evident that heating for 2 hr. at this temperature did not iesiilt iii complete elimination of the cored structure. As would be expected, cold-working of the cast structure is an expedient in establishing equilibrium conditions. The fine homogeneous grain structure shown in Fig. 7 corresponds to the cold-worked area under the Rockwell ball penetration when the hardness was taken before heat-treatment. A detailed structural study was made on a cast silver-thallium alloy containing 3.67 per cent thallium. Three micrographs (Figs. 9, 10 and 11) show the transition

Jan 1, 1946

By F. R. Hensel

PuRe silver and silver-lead alloys have been studied as to their suitability for bearing~.' A review of the properties of thallium and the silver-thallium constitutional diagram was made by the author to analyze the possibilities of silver-thallium compositions for antifriction materia1s. t The silver-rich end of the diagram as reported in the literatureg-" was found to be sketchy and it was necessary to carry out corisiderable experimental work to arrive at definite conclusions. Some of the results of this work are reported in this paper. Test Materials In the beginning of the work, only fused alloys were investigated. Later, research Table i.—Composition of Fused Silver-thallium Alloys Tested for Antifriction Properties Percentage Alloy No. of Thallium 1533...................... 0.48 1534.................... 1.06 a31 HT................... 2.04 231....................... 2.10 232 HT................... 3.83 Balance 232....................... 4.07 Silver 233 HT................... 5.86 a33....................... 6.38 a34....................... 9.84 a34 HT................... 10.07 was carried out on electroplating methods and diffusion processes. In preparing the fused alloyst care was taken to exhaust the toxic fumes caused by the thaillium content. The compositions of the series of fused alloys are listed in Table I. The silver-thallium alloys were melted in clay-graphite crucibles and cast into preheated steel molds of 3/4-in, diameter. In the cast condition, they showed a cored structure, as indicated by the micrographs of Figs. I to 3. The etching reagent used was a mixture of 2 grams K2Cr207, 8 C.C. H2S04, and IOO C.C. H20. The cored structure is unsatisfactory for the type of corrosion resistance required for bearing applications, and homogenizing experiments were carried out at various temperatures, the results of which are shown in Figs. 4 through 8. The alloys with a lower thallium concentration, which were heated for 2 hr. at 525°C show an almost completely homogenized solid solution type of structure. With higher thallium concentration, the homogenizing temperature was dropped to 475OC. to eliminate the formation of a liquid phase. It is evident that heating for 2 hr. at this temperature did not iesiilt iii complete elimination of the cored structure. As would be expected, cold-working of the cast structure is an expedient in establishing equilibrium conditions. The fine homogeneous grain structure shown in Fig. 7 corresponds to the cold-worked area under the Rockwell ball penetration when the hardness was taken before heat-treatment. A detailed structural study was made on a cast silver-thallium alloy containing 3.67 per cent thallium. Three micrographs (Figs. 9, 10 and 11) show the transition

Jan 1, 1946