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By N. Mungan

Laboratory water floods were performed in oil-wet and waterwet alundum and Torpedo cores, displacing a refined oil with n-hexylamine or Triton X-100 solution. Also, some floods were performed in which a sucrose solution was displaced with n-butyl alcohol. The Purpose ot the tests was to see if the oil recotlery could be increased and to examine what role the inter-facial tension reduction and wettability change play in the recovery mechanism. it was found that the chromatographic transport of amine was influenced by core wettability. In oil-wet cores, the rate of advance of the amine band could be predicted from equilibrium chromato-graphy, while in water-wet cores the amine band moved faster than predicted, indicating a non-equilibrium process. By reducing the interfacial tension to 1.1 dyne/ cm, oil recovery was increased. More oil was recovered from Dri-filmed cores than from water-wet cores. Reversing the wettability of the porous media from oil-wet condition also resulted in some additional oil recovery. Neutral wettability floods did not increase oil recovery. INTRODUCTION Interfacial forces in petroleum reservoirs are responsible for retention of large quantities of residual oil. Increasing exploration costs have created an incentive to attempt recovery of this residual oil by new and novel processes. One novel process involves changing the interfacial forces by introducing a chemical into the reservoir during water flooding. Recently, several investigators have carried out laboratory displacement tests using amines. lt3.4 The amines adsorbed onto initially water-wet core surfaces, changing them to oil wet. Additional oil was recovered when the cores were flooded in a manner reversing the. wettability from oil-wet conditions. In all these studies, the wettability reversal was considered responsible for increased oil recoveries although it was noted that amines also reduced oil-water interfacial tension. In the present study, displacement experiments were performed in water-wet and Dri-filmed cores. The purpose was to separate the effects of lowering the interfacial tension from changing the wettability. LABORATORY STUDY All experimental data given in this paper have been obtained at a temperature of 70F + 1 and prevailing atmospheric pressure. FLUIDS AND MATERIALS The oil used in these experiments* was washed with H2S04 and passed through two 100-200 mesh silica-gel beds to remove unsaturates and surface active impurities. Normal hexylamine** of practical purity was used, and the alundum cores*** were composed of nearly pure A1203. Fused quartz (SiOi) and Lucalox (A12 )t plates were used in the contact angle measurements. DISPLACEMENT TESTS Displacement tests were performed in alundum and Torpedo cores. Before displacement runs, cores were acidized with 0.1 normal HCI, flushed with distilled water and fired at 1,200F to make them water wet. This procedure also reduced water sensitivity of the Torpedo cores. Oil-wet cores were prepared from water-wet cores by flowing several pore volumes of 1 per cent by weight Dri-film SC-87 solution in heptane through the cores, flushing the cores with nitrogen, evacuating, and finally heating the cores at 300F for 4 or 5 hours. Between floods, cores were cleaned by conventional procedures and treated again as above.

Jan 1, 1965

By K. C. Strebig, C. W. Schultz, A. A. Selim

For some time researchers have attempted to in- crease drilling efficiency (and therefore reduce costs) by adding certain chemical agents to the circulating medium. The effect of some organic additives on the diamond drilling of quartzite was recently investigated in the laboratory by personnel of the U. S. Bureau of Mines. In these tests, the penetration rate and bit life were found to increase with the use of additives. In addition, an economic analysis of a hypothetical 1000-ft hole showed that additives can reduce drilling costs by 30%. Wetting Agents Improve Drilling Efficiency Much of the interest in the use of wetting agents to improve drilling efficiency is the result of an observation made by Rehbinder, Shreiner, and Zhigach.1 These investigators postulated that drilling creates a region of increased crack formation directly below the bit and that a circulating medium that wets the rock will also penetrate the micro- cracks of the predestruction zone. Increasing the molecular affiinity (wetting ability) of these circulating media for the surface of the rock helps the deformation and destruction by softening the rock in the zone of increased crack formation. In this manner, the hardness or strength of the rock is decreased as the medium wets the rock in the predestruction zone. This is known as the "Rehbinder effect." Rehbinder's work has prompted many investigators to examine the relationship between the use of additives and the mechanical destruction of rock. Engelmann, Terichow and Selim reported that the maximum penetration of the rock by a diamond point in a solution environment was at the isoelectric point of the mineral and solution interface.2 Improvement in rock breakage with the use of additives was observed by Frangiskos and Smith.3 Joris and McLaren suggest that additives having a lower surface tension than water will also be more efficient in cooling the diamonds.4 Additives help remove fine cuttings from the face of the bit and improve efficiency. The latter two authors also report an improvement in impregnated bit life and penetration rates with the use of soluble oil additives. All of these investigations were initiated in the hope of discovering what economic benefits will accrue with the use of additives.

Jan 10, 1969

By S. C. Geol, C. H. Page

INTRODUCTION Mining at Mindola Mine, Rokana Division of Nchanga Consolidated Copper Mines Limited is the deepest on the Zambian Copperbelt. Mining extends to 1150 metre depth so far and the present plans extend to 1500 metres, although reserves exist below this level. Mindola Mine is the major producer of copper/cobalt ore at Rokana Division and produces approximately 200,000 tonnes of copper/cobalt ore per month. A standard open stoping method is the only method used for mining with minor variations in positioning drilling drives and other development, figure 1. The standard main haulage interval is 75m but in flat dipping areas the full stope height is split into two or three panels. In 1970 severe mining problems were experienced when mining deeper than 850111 below surface. The main problems included: - Heavy support requirements in drilling drives near stope face. - Premature closure of blast holes necessitating pre-charging of the holes with explosives immediately after drilling and well before the day of blasting. - Severe bumping and spitting due to rock movements on drive surfaces and remotely within the rock mass under high stress. - Low mining efficiency due to dilution from the waste hangingwall. - These stress related factors also caused some premature detonations of precharged but unprimed blast holes which became a serious safety hazard. A number of alternatives have been considered to overcome these problems. A search for leas sensitive explosives to avoid the premature detonation problem has been partially successful, however, the need for a method of reducing stress levels at the mining face is still considered essential as the mine goes deeper. This paper describes experiments with Regional Support Pillars as a method of reducing stresses at the mining face.

Jan 1, 1981

By W. Weissberger, Frankiewicz, L. Pommier

Introduction In the U.S., about one-quarter of the fuel oil and natural gas consumption is associated with power production in utility and industrial boilers and process heat needs in industrial furnaces. Coal has been an attractive candidate for replacing these premium fuels because of its low cost, but there are penalties associated with the solid fuel form. In many cases pulverized coal in unacceptable as a premium fuel replacement because of the extensive cost of retrofitting an existing boiler designed to burn oil or gas. In the cases of synthetic fuels from coal, research and development still have a long way to go and costs are very high. Another option, which appears very attractive, is the use of solid coal in a liquid fuel form - coal slurry fuels. Occidental Research Corp. has been developing coal slurry fuels in conjunction with Island Creek Coal (ICC), a wholly-owned subsidiary. Both coal-oil mixtures and coalwater mixtures are under development. ICC is a large eastern coal producer, engaged in the production and marketing of bituminous coal, both utility steam and high quality metallurgical coals. There are a number of incentives for potential users of coal slurry fuels and in particular for coal-water mixtures (CWMs). First, CWM represents an assured supply of fuel at a price predictable into future years. Second, CWM is available in the near term; there are no substantial advances in technology needed to provide coal slurry fuels commercially. Third, there is minimal new equipment required to accommodate CWM in the end-user's facility. Fourth, CWM is nearly as convenient to handle, store, and combust as is fuel oil. Several variants of CWM technology could be developed for different end-users in the future. One concept is to formulate slurry at the mine mouth in association with an integrated beneficiation process. This slurry fuel may be delivered to the end-user by any number of known conveyances such as barge, tank truck, and rail. Slurry fuel would then be stored on-site and used on demand in utility boilers, industrial boilers, and potentially for process heat needs or residential and commercial heating. An alternative approach is to formulate a low viscosity pre-slurry at the mine mouth and to pipeline it for a considerable distance, finishing up slurry formulation near the end-user's plant. Finally, at the other extreme of manufacturing alternatives, washed coal would be shipped to a CWM manufacturing plant just outside the end-user's gate. Depending on fuel specifications and locations of the mine and end-user facility, any of these alternatives may make economic sense. They are all achievable in the near term using existing technology or variants thereof. The Coal-Water Mixture CWMs contain a nominal 70 wt. % coal ground somewhat finer than the standard pulverized ("utility grind") coal grind suspended in water; a complex chemical additive system gives the desired CWM properties, making the suspension pumpable and preventing sedimentation and hardening over time. Figure 1 shows the difference between a sample of pulverized coal containing 30 wt. % moisture and a CWM of identical coal/water ratio. The coal sample behaves like sticky coal, while the CWM flows readily. The combustion energy of a CWM is 96-97% of that associated with the coal present, due to the penalty for vaporizing water in the CWM. Potentially any coal can be incorporated in the CWM, depending on the combustion performance required and the allowable cost. CWMs are usually formulated using high quality steam coals containing around 6% ash, 34% volatile matter, 0.8% sulfur, 1500°C (2730°F) initial deformation temperatures, and energy content of 25 GJ/t (21.5 million Btu per st). Additional beneficiation to the 3% ash level can be accomplished in an integrated process. There are a number of minimum requirements which a satisfactory CWM must meet: pumpability, stability, combustibility, and affordability. In addition, a CWM should be: resistant to extended shear, generally applicable to a wide variety of coals, forgiving/flexible, and compatible with the least expensive processing. It was found that a complex chemical additive package and control of particle size distribution are necessary to achieve these attributes simultaneously, while maximizing coal content in the slurry fuel. Formulation of Coal-Water Mixtures A major consideration in the manufacture, transportation, and utilization of a slurry fuel is its pumpability, or effective viscosity. Most CWM formulations are nonNewtonian, i.e., viscosity depends on the rate and/or duration of shear applied. Viscosities reported in this paper were obtained using a Brookfield viscometer fitted with a T-spindel and rotated at 30 rev/min, thus they are apparent viscosities measured at a shear rate of approximately 10 sec-1. The instrument does reproducibly generate a shear field if spindle size and rotation rate are held fixed. By observing the apparent viscosities of several slurries at fixed conditions it is possible to obtain a relative measure of their viscosities for comparison purposes. A true shear stress-shear rate relationship at the shear rates at which the CWM will be subjected in industry may be obtained using the Haake type and a capillary viscometer. These viscometers are used for specific applications. However, for comparison purposes, apparent viscosities are reported.

Jan 1, 1985

By Takahide Wakamatsu, Yoshiaki Numata, Charn Hoon Park

INTRODUCTION A technique is needed to apply the flotation method successfully to apply the flotation method successfully to low grade and complex ore bearing hard-to-float oxide minerals. Thus it is necessary to find a new selective, powerful collector or to develop a new flotation process. Some attempts to use two or more kinds of collectors combined in flotation have been made but the principle of the reaction between the collectors used and the surface of mineral as well as the effect of combined collectors on flotation has not yet been fully understood. As part of the study on the improvement of selective collection in flotation, we have been investigating the flotation of minerals using two kinds of organic reagents, one of which is a conventional collector. We found xanthates were able to adsorb even onto the surface of oxide mineral, with the help of another organic reagent, e.g. a chelate compound (1). Furthermore, a notable effect of combined reagents such as the xanthate-amino acid combination (2) was found on the reagent adsorption phenomena and the collection phenomena, depending on the combination of reagents. The present study was carried out to clarify the effects of adding two kinds of collectors to the collection of mineral in flotation in the following two cases: (1) joint use of xanthate and di-alkyl dithiophosphate for galena collection and (2) joint use of a cationic collector with xanthate for scheelite collection.

Jan 1, 1980

By A. E. Jenkins, L. A. Baker, N. A. Warner

The electromagnetic levitation technique has been successfully applied to rate studies of the de-carburization of liquid Fe-C alloys from 5.5 to zero pct C at 1660°C using gas mixtures containing 1 to 100 pct CO2 with CO or helium as diluents. The rate was shown to be independent of carbon concen -tration in the melt and only slightly affected by total pressure. The results were shown to be consistent with control either wholly or predominantly by gaseous diffusion. As well as avoiding side reactions with the crucible, levitation allowed precise calculation of reaction rates since the geometry of the system was accurately known. It also enabled interpretation using the established mass-transfer correlations for a gas flowing past a single sphere, The existence of gaseous-diffusion control and the extremely high rates of reaction reported for gas flowing past small spheres are of great practical significance in relation to current investigations into the use of spray-refining techniques in continzcous steelmaking. PREVIOUS work on the kinetics of carbon removal from liquid iron has generally involved the use of oxygen as the gaseous oxidant. Investigations have been made on commercial steelmaking furnaces and in addition laboratory-scale experiments have been attempted with a slag phase present. The general lack of agreement and the obvious difficulty in interpreting the results is at least partly due to the complexity of the systems used. This work has been reviewed by carter1 and more recently by Bods-worth2 and ward.3 The gas-liquid metal reactions have also been studied with slag-free melts by the conventional technique of blowing oxygen onto the melt contained in a crucible. The reaction between the oxidant and carbon could be followed by sampling either the melt or the gas phase. Side reactions with the crucible material have precluded direct interpretation of the results. Fujii, 5-7 using O-A mixtures, attempted to correct for the crucible reaction and reported that the reaction between carban and oxygen occurred at the melt surface and was controlled by gaseous diffusion for carbon contents greater than a critical value and by carbon-diffusion control in the metal for carbon contents less than the critical. The critical value was given as approximately 0.15 pct C. However, whereas this was the case for oxygen contents in the gas stream up to 10 pct, between 10 and 20 pct the rate was found to be approximately independent of oxygen content, and Fujii attempted to explain this observation by postulating the presence of an oxide "film".7 Turkdogan et a1.' have studied iron fuming during decarburizing and suggested gaseous-diffusion control of the reaction above a critical carbon concentration, given as approximately 2 pct C. Filippov9-12 reports in a similar manner to Fujii but makes no mention of the crucible reaction with the carbon in the melt, even though this would be expected with the temperature, crucible material (MgO), and carbon contents used. Both CO2 and O2 were used as oxidant gases. In the present work the decarburization of liquid Fe-C alloys was studied using an adaptation of the electromagnetic levitation technique, and thus the complexities introduced by crucible side reactions were eliminated. In this technique a single sphere of the Fe-C alloy is freely suspended and heated by electromagnetic induction in a flowing stream containing the oxidant gas. The full range of carbon levels up to saturation was studied, the temperature of the melts in this instance being held constant at 1660°C. The oxidant was CO2 which was used either pure or in association with a diluent gas, the total gas pressure being maintained either at 1 atm or at a reduced pressure. Thus the reaction under study was:

Jan 1, 1964

By Paul Hunt

UP to June, 1923, the Park-Utah mine had shipped about 94,000 tons of a direet-smelting ore of a gross value of $4,200,000, or about $45. a ton. These values were in gold and silver only, although the ore carried small amounts of lead, copper and zinc. Freight, treatment and deductions had averaged $15.40 a ton. At the expiration of the Pittman Silver Act there was in sight about 30,000 tons of a similar ore with a gross value of about $15 a ton in gold and silver, and a large body of partly developed lead-zinc ore of a gross value of about $20 a ton, 40 per cent. of which was represented by the lead-zinc content and the balance by the silver and gold. Mine costs had been high during the life of the Pittman Act, but as a large part of the profits had gone into equipment of the mine and the orebodies were large and had good walls, we knew that mine costs could be greatly reduced. We were therefore faced by the necessity of getting lower freight and smelter rates or of concentrating the ore. We did obtain better freight rates on lower grade ore, and a favorable smelting rate on the small tonnage of siliccous silver-gold ore, but this smelter rate could not be applied to the lead-zinc ore with any profit to ourselves. The Judge mill of the Park City Mining & Smelting Co., with which we were affiliated, was running to capacity; if we built our own mill it ineant a 5-mile freight haul to a satisfactory millsite and a year's delay. Fortunately, the International Sillelting Co. had just inherited a mill at Tooele which, with some changes, was about double the catpscity required for its own ores. Flotation tests were run on Park-Utah ore and a contract entered into with a low base treatment charge per ton, the tariff being largely fixed by the metallurgical deductions. This contract has been admirably calculated so that the miner can follow a lean orenbody to its limit. Such a procedure has no doubt made it difficult for the smelting company at times, but it has resulted in finding ore extensions that otherwise might have been missed, so In the long run it has profited too. A flat rate per ton would have been like mining up to the claim endlines in a great many of the stopes. In order to offer such a contract it can be inferred that the smelting company must have been able to make high reeoveries and high-grade products on a very lean ore.

Jan 1, 1928

By T. L. Chu, G. A. Gruber

Silica films hare been rleposited 011 silicon substmtes at 400° to 600°C by a chemical-transport technique using hydrogen fluoride as the transport agent ill a closed system. This transport takes place from a source materia1 1071: temperature to substrates at higher temperatures, as indicated by the thermochemistry of the transport reaction. The experimental variables of- the transport process, such as the substrate temperature, the pressure pi the transport agent, and so forth, have been studied. The rate -determining step of the transport process appears to he the ),ale of chemical reaction in the source region. The transported films are similar to thermally grown silica films in physical proper-ties with the exception of 'some what higher dissolrrtion rates. SILICA films deposited on suitable substrates serve many purposes in electronic devices. They are used for the fabrication of tunneling devices, the surface passivation of devices, and the shielding of devices from nuclear radiation: and as selective masks against the diffusion of specific impurities into semiconductors. Doped silica films can also be used as sources for the diffusion of impurities into semiconductors. Several oxidation and deposition techniques for the preparation of silica films have been developed to meet the requirements of these applications. The therma1 oxidation of silicon by oxygen or steam at temperatures above 900 C is commonly used in silicon technology. The deposition techniques are perhaps more advantageous since they usually require lower temperatures and are not limited to silicon substrates. Silica films have been deposited on silicon and other substrates by reactive sputtering and chemical reactions. The sputtering of silicon in an oxygen atmosphere is capable of depositing good-quality silica films on silicon' and gallium arenide. Many chemical reactions are known to yield silica at room temperature or higher. These reactions may involve intermediate steps. However, the final step yielding silica should take place predominately on the substrate surface in order to produce adherent films. When silica is formed in the gas phase by volume reactions, no adherent deposit can be obtained. Generally, the experimental conditions of a reaction can be varied so that the surface reaction predominates over the volume reaction. The chemical reactions which have been used successfully for the deposition of silica films are briefly as follows. The pyrolysis of alkoxysilanes in an inert atmosphere or under reduced pressure has been employed to deposit silica films on germanium3 and silicon4 at 650" to 750°C in a flow system. The deposition of silica films from alkoxysilanes has also been achieved at nearly room temperature by a low-pressure plasma. Device quality silica films have been deposited on germanium and gallium arsenide by the deposition of an amorphous thin silicon film followed by oxidation at 600" to 700" . Silica films for high-temperature capacitors have been produced by the hydrolysis of silicon tetrabromide at 950°C in argon and hydrogen atmospheres.7 We have developed a chemical-transport technique for the deposition of silica films on semiconductor substrates at relatively low temperatures. The thermochemistry of the transport reaction, the experimental variables of the transport process, and the properties of the transported silica films are described in this paper. THERMOCHEMICAL CONSDERATIONS The transport of solid substances by chemical reactions in the presence of a temperature gradient has been used for the preparation of films and crystals of many electronic materials. In this technique, a gaseous reagent is chosen so that it reacts reversibly with the solid substance under consideration to form volatile products. Since the equilibrium constants of most reactions are temperature-dependent, the transport of these products to regions of suitable temperature in the reaction system would cause the reverse reaction to take place. depositing the original solid. When the equilibrium is shifted toward the formation of the solid as the temperature is decreased, the solid is transported from a high-temperature zone to a lower-temperature region, and vice versa. This chemical-transport technique can be carried out in a closed or gas-flow system. In a closed system, chemical equilibrium is presumably established in the different temperature regions of the system, and the transport agent regenerated in the deposition region repeats the transport process in a cyclic manner. The local chemical equilibrium may not be approached in a flow system: however, this system offers a greater degree of flexibility. Silica reacts reversibly with hydrogen fluoride and this reaction was chosen for the transport process. The over-all reaction between silica and hydrogen fluoride may be written as: SiO2(s) + 4HF(g-) = SiF4Ur) + 2H2O(^)

Jan 1, 1965

THE following statement is contained in an article entitled "Mortgaging the Future." by Walter C. Teagle, president of the Standard Oil Co. of New Jer-sey, appearing in the current issue of the Lamp, the official organ of the Standard Oil Co.: "During the last eight months the domestic daily production. of crude oil has risen 300,000 bbl.. which means that no less a volume than 2,500,000 bbl. now are coming to market each day. . . . Every twenty-four hours 250,000 bbl. more than the consumer required are being added to our surplus oil storage. The temporary advantage of cheap petroleum products and the availability of fuel .oil, for a widespread use as a substitute for coal are not being weighed against the time when scarcity may follow on the heels of plenty and the public of tomor-row may have to pay for today's excesses. The over-production of crude and its consequences constitute a problem of national import which far outweighs the temporary advantages either to a part of the industry or to sections of the public.

Jan 8, 1927

By T. D. Tiemann

The extraction of iron oxide from several types of siliceous iron ore by digestion in sulfuric acid solutions was investigated in the temperature range from 80 to 258°F at concentrations from 10-60% H2SO4. Degree and rate of extraction were found to be critical with respect to both temperature and acid concentration, requiring temperatures in excess of 250° F and acid concentrations of 50% to effect 80-90% extraction of iron oxide in 60 min. Rate of extraction is directly related to surface area and apparently controlled by the rate of chemical reaction. The activation energy was found to be approximately 25,000 cal per g mol. Extraction characteristics of the several ores were found to be similar. The Department of Minerals and Metals Engineering at the University of Wisconsin conducts a continuing research program to develop processes for the utilization of low-grade iron ore deposits such as the nonmagnetic taconites of the Wisconsin Gogebic range. Methods under study include high-intensity magnetic and high-tension electrical concentration, flotation, direct reduction and chemical extraction. Concentration of a number of low-grade siliceous iron ores by the extraction of silica in sodium hydroxide solutions has been reported by the writer and others in the literature.1,2,3,4 Residual concentrates containing as much as 65% iron have been obtained from taconites containing less than 30%. Currently an alternative approach is under study, namely the dissolution of iron oxides from siliceous gangue by extraction in acid, with subsequent recovery of the iron from solution. Such a process for extracting iron oxide from the low-grade oclitic deposits of Alberta with hydrochloric acid was investigated and described by Gravenor, Govett and Rigg.5 The current investigation, which is entirely devoted to extraction with sulfuric acid, is the subject of this paper. In part, the objective of this research was to determine the possibility of leaching iron ores in situ with sulfuric acid. The investigation of a recovery method is left to later research. The ores investigated contained hematite (Fe2O3) as the principal iron-bearing mineral and minor amounts of goethite (Fe2O3 .H2O), magnetite (Fe3O4), and iron-bearing silicates. The gangue was essentially quartz occurring as single crystals and microcrystalline aggregates. Detailed description of the ores may be found in the literature.1,2,4,6 A comparison of the specific ores used in this investigation is shown in the partial analyses of Table I. EXPERIMENTAL METHODS All extractions were made at atmospheric pressure using beakers with a mechanical stirrer for those tests run below the boiling point, and Erlenmeyer flasks for those at the boiling point. A thermostatically controlled hot plate was used for heating, and this maintained the mean temperature within the solution to within ±3°F during extraction. The sample of ore for extraction was weighed to + 0.02 g. For tests below the boiling point, the sample was added to the beaker containing the sulfuric acid solution at the desired temperature. For the boiling-point tests, the sample was added to the Erlenmeyer flask followed by distilled water and the

Jan 1, 1968

By W. J. Carlson, I. Iwasaki

A previous article reported a combination process for recovering both metallic iron and metallic manganese from the manganiferous iron ores of the Cuyuna Range in Minnesota.' In this process, the iron was recovered by direct reduction and magnetic separation and the manganese was recovered from the nonmagnetic tailing by leaching and electrolysis. The current efficiency during the electrolysis of manganese is quite low, in the range of 60 to 70%, due to the simultaneous plating of hydrogen.2,3 The parameters that affect the current efficiency are (1) the concentrations of the manganese ion and the ammonium sulfate buffer, (2) the effect of autoxidant, (3) temperature, (4) pH, and (5) current density. To study the plating characteristics of manganese metal.

Jan 1, 1969

Jan 1, 1891

By J. Gordon Parr, A. J. Goldak

OgdEN et al.1 and Bumps et al.2 suggested that the solubility of aluminum in a titanium extended to 30 pct.* Sagcl,3 Clark and Terry,4 Anderko et al.,5 Ence and Margolin6 and Saulnier and croutzelles,7 subsequently proposed the existence of an intermetallic compound in the region of 25 pct Al. Although Anderko et al.5 suggested that the structure of this phase, Ti3Al, was of the DO,, (Ni3Sn) type, their conclusion was based only upon the results of a powder photograph that included reflections from ? = 0 to ?= 27° (CuK),some of which were reported to be indistinct. Ence and Margolin,6 referring to a 25 pct alloy (but without indicating whether the figure referred to weight or atomic pct) stated: "The structure of Ti2Al is isomorphous with Ti3Sn DO19, as pointed out by Pietrokowsky". But the point was not proven, because Ence and Margolin did not report values of calculated intensities of Ti2Al with the DO,, structure. In fact, the observed intensities they reported are in poor agreement with the values we calculate for DO19 (see Table I). Saulnier and croutzelle7 concluded, on the basis of electron micrography and microdiffraction studies, that the structure of the 25 pct A1 alloy was ordered hexagonal. Our purpose has been to unequivocally determine the Ti3Al structure. Alloys containing 25 pct Al were prepared from iodide titanium and spectrographic standard aluminum by levitation melting. The levitation chamber was evacuated to 10-6 mm Hg before filling with 99.999 + pct A. After casting, the alloy was chilled to liquid air temperature and crushed to -200 mesh in a percussion mortar. The powder, wrapped in molybdenum foil, was annealed (by induction heating) at 1000 "C for 2 hr in the argon atmosphere of the levitation chamber. No sublimate was seen on the foil or elsewhere, and therefore the assumption is

Jan 1, 1962

By M. Adnan Goksel

The paper describes the most common cold bond processes including the Carbonate-Bond, Grangcold, Cobo, MTU and others, and outlines these along with bonding mechanisms, physical and chemical characteristics of the products, and applications to various minerals and waste materia1s.

Jan 1, 1977

By D.E. Norquist, J. E. Moeller

The results of wear on marked balls, 4, 31/2, 3, and 2 in. diam are given. All balls were forged steel of practically the same chemical analysis and hardness. The results indicate that balls in a given mill for a given length of time will have equal diameter losses, regardless of size. IN order to determine the relative wear rates of several sizes of grinding balls, groups of 4, 31/2, 3, and 2 in. balls were marked individually and were charged, all at the same time, into each of two production mills grinding copper ore. The remainder, and vast majority, of the ball charge consisted of 2 in. diam, and smaller, white, cast iron balls. The original weight of each test ball was determined and recorded. Some of each group of test balls were recovered from the mills periodically, individually reidentified, and reweighed. The test balls were recovered during regular maintenance shutdowns, at approximately thirty-day intervals, so as to avoid disrupting operations. As soon as the weights had been recorded, each marked ball was recharged into the mill from which it had been taken. The test units were mechanically the same and theoretically received equal tonnages of feed. All of the test balls were of forged, alloy steel composition with only slight variations in chemistry that would not be expected to affect the physical properties. All were heat treated in the same manner to produce a high hardness, as equal as possible for all sizes and as uniform as possible from surface to center. Generally speaking, the hardness of the portion worn from the balls during the test ranged from 62 to 65 Rockwell C. The microstructure was, therefore, predominately martensitic. The method of marking the test balls consisted of forming a small, round hole, 3/16 or 1/4 in. diam, into the center of each test ball with a Thomas Metal Master disintegrator, which produces a hole in fully hardened steel without affecting the hardness of the surrounding metal. After the hole was formed, each test ball was individually weighed to the nearest gram and the weight was recorded. The balls were grouped as to size and each group assigned a code letter. The individual balls within the various groups were then assigned a number. A small copper disc was stenciled on one side with the group letter and on the other side with the individual ball number and dropped into the hole in the corresponding test ball. The hole was then plugged with a low-melting-point metal alloy, which could be melted out in boiling water. Half of the test balls in each group were charged into one mill and the other half into another mill, so that all test balls in each mill would be subject to, as nearly as possible, identical conditions for an equal length of time. Previous testing had indicated that the wear rates on balls of different sizes could not be accurately compared if weight loss, or percentage weight loss, were used as a basis for comparison. Fig. 1 is a graphic record of the average weight loss for each size group in mill No. 1, and it becomes readily apparent that the larger balls lose substantially more weight in a given length of time. Fig. 2 is a graphic record of the average percentage weight loss for each size group in mill No. 1, illustrating that the smaller balls definitely lose a greater percentage of their original weight in a given length of time. It was, therefore, decided to convert the actual weights of the test balls to diameter and follow the diameter losses periodically to see if the wear rates

Jan 1, 1951

By D. E. Norquist, J. E. Moeller

The results of wear on marked balls, 4, 31/2, 3, and 2 in. diam are given. All balls were forged steel of practically the same chemical analysis and hardness. The results indicate that balls in a given mill for a given length of time will have equal diameter losses, regardless of size. IN order to determine the relative wear rates of several sizes of grinding balls, groups of 4, 31/2, 3, and 2 in. balls were marked individually and were charged, all at the same time, into each of two production mills grinding copper ore. The remainder, and vast majority, of the ball charge consisted of 2 in. diam, and smaller, white, cast iron balls. The original weight of each test ball was determined and recorded. Some of each group of test balls were recovered from the mills periodically, individually reidentified, and reweighed. The test balls were recovered during regular maintenance shutdowns, at approximately thirty-day intervals, so as to avoid disrupting operations. As soon as the weights had been recorded, each marked ball was recharged into the mill from which it had been taken. The test units were mechanically the same and theoretically received equal tonnages of feed. All of the test balls were of forged, alloy steel composition with only slight variations in chemistry that would not be expected to affect the physical properties. All were heat treated in the same manner to produce a high hardness, as equal as possible for all sizes and as uniform as possible from surface to center. Generally speaking, the hardness of the portion worn from the balls during the test ranged from 62 to 65 Rockwell C. The microstructure was, therefore, predominately martensitic. The method of marking the test balls consisted of forming a small, round hole, 3/16 or 1/4 in. diam, into the center of each test ball with a Thomas Metal Master disintegrator, which produces a hole in fully hardened steel without affecting the hardness of the surrounding metal. After the hole was formed, each test ball was individually weighed to the nearest gram and the weight was recorded. The balls were grouped as to size and each group assigned a code letter. The individual balls within the various groups were then assigned a number. A small copper disc was stenciled on one side with the group letter and on the other side with the individual ball number and dropped into the hole in the corresponding test ball. The hole was then plugged with a low-melting-point metal alloy, which could be melted out in boiling water. Half of the test balls in each group were charged into one mill and the other half into another mill, so that all test balls in each mill would be subject to, as nearly as possible, identical conditions for an equal length of time. Previous testing had indicated that the wear rates on balls of different sizes could not be accurately compared if weight loss, or percentage weight loss, were used as a basis for comparison. Fig. 1 is a graphic record of the average weight loss for each size group in mill No. 1, and it becomes readily apparent that the larger balls lose substantially more weight in a given length of time. Fig. 2 is a graphic record of the average percentage weight loss for each size group in mill No. 1, illustrating that the smaller balls definitely lose a greater percentage of their original weight in a given length of time. It was, therefore, decided to convert the actual weights of the test balls to diameter and follow the diameter losses periodically to see if the wear rates

Jan 1, 1951

By F. D. Rosi

The slip and twinning behavior in extended titanium crystals were studied in some detail. The formation and appearance of coarse kink bands are discussed. Their crystallographic geometry was determined by X-ray analysis. A phenomenological interpretation of the complexities in kink band development is also presented. The critical resolved shear stress, coefficient of shear hardening, and plane of fracture were determined for several crystals extended at room temperature. THE slip and twinning elements observed in the room-temperature deformation of titanium were enumerated in a previous paper1 in which considerations were advanced regarding the nature and selection of these elements and their effect on the known mechanical properties of this metal. The present study concerns the crystallographic, microscopic, and mechanical aspects of flow in relation to the slip and twinning elements, and includes a prediction of slip systems, nature of slip and twin markings, inhomogeneities of plastic flow, and stress-strain characteristics. Unless otherwise noted, arc-melted titanium sponge (99.77 pct) was used in these experiments. The method of production of crystals, their dimensions, and the surface preparation for micrographic examination have been reported.&apos; For obtaining stress-strain characteristics, only those crystals which traversed the entire width of the specimen and were at least 8 mm in length were used. Tensile deformation of the crystals was performed with conventional grips for sheet specimens and a constant-stress loading beam, designed after the method of Andrade and Chalmers.&apos; The specimens were loaded by allowing sand to flow from a reservoir into a bucket suspended from the longer end of a balanced 6:1 lever arm at a rate controlled to load the specimen approximately 2 kg per min. Strain measurements were made using the Baldwin SR-4, bonded, resistance-wire strain gage, Type A-8, which permitted a reading accuracy of 2 microinches per inch. The formulas used in the evaluation of shear stress and shear strain, as in deriving the coefficient of shear hardening, are given by Schmid and Boasv n terms of the original orientation and change in length of the crystal. For calculating the critical resolved shear stress, the standard equation was used. The crystallographic nature of the unpredictable slip observed in a number of specimens was determined by the single-surface X-ray method of analysis as described in ref. 1. Experimental Results Prediction of Slip System: It was reported&apos; that room temperature slip in titanium takes place predominantly on {10i0} and in a <1120> direction, giving three potential slip systems. Hence, it should be possible to predict the operative primary system in the manner used by Taylor and Elam&apos; for alu- minum (i.e., the one with the greatest component of shear stress in the direction of slip). The stereo-graphic construction in Fig. 1 shows that this is true. In all cases, slip was found to occur initially on the (0110) plane and, in a number of cases, in the [21f0] direction. (The direction of slip was not determined for all orientations.) It follows that duplex slip can be expected when two slip systems are geometrically equally favorable for slip, as demonstrated in Fig. 2. In both crystals slip took place simultaneously on the (1700) and (olio) prismatic planes, as indicated by Fig. 2a and b. In the case of crystal B the movement of the specimen axis with increasing extension was toward the (10i0) direction, which represents the stable end orientation resulting from alternate slip on the [2ii0] and [1120] directions. This is analogous to the duplex slip process in face-centered cubic metals." Unpredictable Slip: In a number of crystals whose orientations were well within the operation of a single slip system, secondary slip occurred on planes not predicted by the criterion of maximum shear stress. Examples are shown in Fig. 3. In addition to

Jan 1, 1955

Jan 1, 1916

Jan 9, 1951

"The Tooele plant is situated at the mouth of Pine Canyon, five miles northeast of Tooele City. The site forms one terminus of the Tooele Valley railway, which runs southwest through Tooele City to Warner, a station on the Los Angeles and Salt Lake Railroad, thirty-five miles southwest of Salt Lake City.An aerial tramway, four miles in length, crosses the Oquirrh Mountains and connects the plant with the Bingham district.The plant consists of a lead-zinc flotation concentrator, lead smelter and copper smelter.SAMPLING MILL The mill is a standard Taylor and Brunton sampling mill, built in two sections, each with a capacity of 100 tons per hour. The building is of steel and concrete, 40x84 feet. A well-equipped pulping, sampling and drying room is attached.All ore is sampled here and then distributed by conveyors to cencentrator, lead smelter and copper smelter bins.CONCENTRATOR This flotation mill is arranged in two units of 500 tons capacity each, one equipped with Callow pneumatic flotation cells, the other with Mineral Separation sub-aeration machines.The feed is maintained as uniform as possible by bedding, after coarse crushing at the sampling mill. Standard practice of ball mill grinding and classification is followed."

Jan 1, 1925