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By Felipe E. Olivares S

"El presente texto revisa el origen y distribución de los metales raros como parte de un proyecto de recuperación de subproductos de valor comercial en la fundición de La Oroya. Para ello se reeexaminan ocho minerales raros (bismuto, selenio, teluro, cadmio, talio, indio, bismuto, indio de alta pureza), cuya composición de elementos raros se registra en la concentración de algunos productos o residuos. Estos minerales son del grupo del cobre Chalcópirita, Enargita, Tetrahedrita, Bornita y Chalcosita; del plomo: la Galena, Jamesonita y Gratonita, y la Esfalerita que proporciona el zinc. Los concentrados y minerales que entran al proceso de fundición se reparten en tres circuitos principales: cobre, plomo y zinc. Los circuitos de los primeros están ligados entre sí por el sistema de Cottrelles y los procedimientos de las mismas fundiciones, mientras que el último actúa de forma independiente. Por último, es competencia del Departamento de Investigaciones, la recuperación de dichos subproductos a través del proceso de concentración y fundición, como de a búsqueda nuevos métodos en el campo minero y metalúrgico. "

Jul 20, 1956

By L. W. Beckstead

In a conventional zinc refinery plant, zinc concentrates were roasted at 950°C. The calcine was ground and leached in a series of leaching vessels using sulfuric acid. The pH in the first leaching vessel was maintained at 2.5, while the pH of the solution exiting the last vessel was about 5.4. The leached residue typically contained (in percent) 0.1 Ag, 0.2 As, 0.7 Cd, 2 Cu, 15 Fe, 10 Pb, and 15 Zn, and it was classified as hazardous material according to the EP-toxicity test. In order to make the residue non-hazardous, and to recover most of the metal values from the residue, a brine leach process was developed. The process consisted of leaching the zinc leach residue in CaCI, brine solution at pH 2 for 1 hour at 90°C. Silver extraction was about 95 percent, while lead, copper, cadmium, zinc, and iron extractions were about 80, 50, 50, 30, and less than 0.5 percent, respectively. More extreme conditions of acidity could have been used in the leaching stage to increase zinc dissolution, but iron dissolution would have significantly increased which would have complicated subsequent recovery steps. The leached and washed residue passed the EP-toxicity leach test and could be disposed as non-hazardous. Silver, copper, and lead in the leach solution were recovered by zinc cementation, while zinc and cadmium were recovered by sulfide precipitation. The resulting solution can be recycled in the process after magnesium precipitation.

Jan 1, 1993

By C. O. Lee, R. H. Feierabend, Z. W. Bartlett

DESPITE the fact that American brimstone production has increased 2 times since the prewar period 1935-1939, the demand for sulphur exceeds the supply. To alleviate this situation efforts are being made to increase the production of sulphur and bring the shortage to an end. The most important developments are being undertaken on the Gulf Coast. One new brimstone mine has been put into operation and three others are being developed. These projects are expected to account for a large part of the additional sulphur production now anticipated for the free world.'.' Except for a small amount of byproduct sulphur from sour natural gas and refinery gases, American brimstone is produced by the same process devised by Frasch 60 years ago, consisting essentially of the injection of superheated water at about 325°F into sulphur-bearing formations through wells which are drilled from the surface by standard oilfield methods. Sulphur deposits vary in depth from a few hundred feet to 2500 ft from the surface; in a typical deposit the sulphur occurs as crystalline aggregates in porous limestone and gypsum in concentrations varying from 5 to 40 pct, the thickness of the formation ranging from a few feet to several hundred.3 The deposits are usually surrounded by formations impermeable to the flow of hot water, the top and sides being enclosed by impermeable sediments and the bottom by dense impermeable anhydrite. As the hot water leaves the well bore, it percolates through the formation and transmits its heat to the rock and sulphur. When the sulphur reaches 240°F it melts, separates from the rock, and flows through the hot channels downward to the bottom of the well. Here it is freed from incoming water and from formation water, which saturates the porous rock, by gravity separation, liquid sulphur being 1.8 times as heavy as water. It then is pumped to the surface in a relatively pure state, 99.5 pct or better. Chief contaminants are organic impurities, such as petroleum residues, which are soluble in sulphur and tend to discolor it.' Sixteen deposits of commercial magnitude have been discovered in the Gulf Coast area. Five of these deposits have been depleted and abandoned with production totalling more than 26,886,000 tons. Eight properties are operating and have produced 65,875,-000 tons to Jan. 1, 1952, for a total Frasch-type production of 92,761,000 tons. Three new deposits are under development. Aforementioned deposits are listed in Table I. Each of these deposits has been found in the cap-rock overlying shallow intrusive salt plugs or domes and buried in several hundred to as much as 2500 ft of sediments. All salt domes do .not contain sulphur; only 16 sulphur deposits have been found in a total of 231 salt domes which are known to exist in the Gulf Coastal states of Texas, Louisiana, Mississippi, and Alabama. Ten of the salt domes lie in the Gulf of Mexico offshore from the Texas and Louisiana coasts. Additional offshore domes may be found, but their value as a sulphur potential is questionable owing to the great hazard and high cost that can be visualized at these exposed locations where depths of water reach 50 ft. Marsh operations of Freeport Sulphur Co. are in the southeast section of Louisiana. In general, the country is low, flat, relatively uninhabited marshland broken by many shallow lakes and bayous which extend for irregular distances inland. Geologically, the coastal area is of sedimentary origin, having been built up gradually under the sea by deposition of material carried down by large streams to their deltas. Offshore bars were formed enclosing

Jan 1, 1953

By C. O. Lee, Z. W. Bartlett, R. H. Feierabend

DESPITE the fact that American brimstone production has increased 2 times since the prewar period 1935-1939, the demand for sulphur exceeds the supply. To alleviate this situation efforts are being made to increase the production of sulphur and bring the shortage to an end. The most important developments are being undertaken on the Gulf Coast. One new brimstone mine has been put into operation and three others are being developed. These projects are expected to account for a large part of the additional sulphur production now anticipated for the free world.'.' Except for a small amount of byproduct sulphur from sour natural gas and refinery gases, American brimstone is produced by the same process devised by Frasch 60 years ago, consisting essentially of the injection of superheated water at about 325°F into sulphur-bearing formations through wells which are drilled from the surface by standard oilfield methods. Sulphur deposits vary in depth from a few hundred feet to 2500 ft from the surface; in a typical deposit the sulphur occurs as crystalline aggregates in porous limestone and gypsum in concentrations varying from 5 to 40 pct, the thickness of the formation ranging from a few feet to several hundred.3 The deposits are usually surrounded by formations impermeable to the flow of hot water, the top and sides being enclosed by impermeable sediments and the bottom by dense impermeable anhydrite. As the hot water leaves the well bore, it percolates through the formation and transmits its heat to the rock and sulphur. When the sulphur reaches 240°F it melts, separates from the rock, and flows through the hot channels downward to the bottom of the well. Here it is freed from incoming water and from formation water, which saturates the porous rock, by gravity separation, liquid sulphur being 1.8 times as heavy as water. It then is pumped to the surface in a relatively pure state, 99.5 pct or better. Chief contaminants are organic impurities, such as petroleum residues, which are soluble in sulphur and tend to discolor it.' Sixteen deposits of commercial magnitude have been discovered in the Gulf Coast area. Five of these deposits have been depleted and abandoned with production totalling more than 26,886,000 tons. Eight properties are operating and have produced 65,875,-000 tons to Jan. 1, 1952, for a total Frasch-type production of 92,761,000 tons. Three new deposits are under development. Aforementioned deposits are listed in Table I. Each of these deposits has been found in the cap-rock overlying shallow intrusive salt plugs or domes and buried in several hundred to as much as 2500 ft of sediments. All salt domes do .not contain sulphur; only 16 sulphur deposits have been found in a total of 231 salt domes which are known to exist in the Gulf Coastal states of Texas, Louisiana, Mississippi, and Alabama. Ten of the salt domes lie in the Gulf of Mexico offshore from the Texas and Louisiana coasts. Additional offshore domes may be found, but their value as a sulphur potential is questionable owing to the great hazard and high cost that can be visualized at these exposed locations where depths of water reach 50 ft. Marsh operations of Freeport Sulphur Co. are in the southeast section of Louisiana. In general, the country is low, flat, relatively uninhabited marshland broken by many shallow lakes and bayous which extend for irregular distances inland. Geologically, the coastal area is of sedimentary origin, having been built up gradually under the sea by deposition of material carried down by large streams to their deltas. Offshore bars were formed enclosing

Jan 1, 1953

By R. D. Saltsman

Under contract with the Federal Bureau of Mines, Bituminous Coal Research, Inc., performed a statistical study to find less expensive and time-consuming methods of monitoring the noncombustible content of coal mine dust. Calculations based on thermodynamic considerations verified the reasonableness of Federal rock-dusting requirements. The results of extensive sampling in southwestern Pennsylvania mines indicated that, at the 95-percent confidence level, the present system of sampling at 200-foot intervals may be replaced by sampling at 500-foot intervals, and that a 1-inch scoop floor sampler may replace the standard 6-inch scoop for floor sampling. Further field testing is recommended.

Jan 1, 1975

By C. E. Jordan

The fundamental characteristics of ultrafine chalcopyrite flotation were investigated by the Bureau of Mines to establish techniques to improve recovery of ultrafine chalcopyrite. Intense agitation during conditioning was found to improve flotation recovery of the ultrafine chalcopyrite. A split-circuit flotation design, treating the plus 38-µm-size ore by the conventional method and treating the minus 38-µm-size ore with intense agitation conditioning followed by flotation, recovered 95.6 pct of the copper in a 28.3-pct-copper concentrate, This was a 5-pct increase in copper recovery over the conventional chalcopyrite flotation method.

Jan 1, 1986

By Milos Krofta, Donald B. Aulenbach, James P. Van Dyke

INTRODUCTION An innovative Ø.1-MGD (378.5-m3/day) biological and phyicochemical wastewater treatment system at Hancock, Massachusetts, USA, has been treating the domestic sewage generated from the four-season Jiminy Peak timeshare resort since 1985. The influent raw sewage is collected and equalized in an equalization tank, then biologically treated by two rotating biological contactors (RBC). A portion of the RBC effluent is discharged to a secondary sedimentation tank to simulate conventional waste treatment by a conventional RBC system. For parallel treatment the remaining portion of the RBC effluent is discharged to an innovative Sandfloat (SASF-8) clarifier to simulate a two-stage biological and physicochemical treatment by RBC-Sandfloat. Both secondary sedimentation effluent and Saandfloat effluent are later combined together and discharged to a tertiary sand filter and two UV disinfection'units for final polishing. The operation and performance data of Jiminy Peak plant in the period of May 1985 to March 1988 was fully documented in the Lenox Institute technical report No. LIR/Ø3-88/292 (391 pages, March 1988). This report presents only background information and complete plant data in the period of April 1988 to January 1989. Dr. Lawrence K. Wang and Mr. James P. Van Dyke have been the Consulting Engineer and Director of Special Projects, respectively, for the Jiminy Peak Project since 1985. (Key Words) Innovative and Alternative Technology, Rotating Biological Contactor, Sandfloat, Dissolved Air Flotation, Sedimentation, Tertiary Filtration, Ultraviolet, Disinfection. BACKGROUND INFORMATION Jimmy Peak, Inc., Hancock, Massachusetts, has designed and constructed a secondary wastewater treatment system consisting of preliminary/primary treatment facilities, two rotating biological contactors, a dissolved air flotation clarifier, a secondary sedimentation clarifier, a sand filter and two ultraviolet (UV) disinfection units. Total average wastewater flow is approximately Ø.1 ) (378.5 cubic meters per day). In accordance with the Consulting Engineer's preliminary design, the total wastewater stream is initially collected and equalized by an equalization tank, and then biologically treated by rotating biological contactors (RBC). Part of the RBC effluent is diverted to a Krofta Sandfloat (SASF-8) for physicochemical clarification in order to simulate an innovative RBC-Sandfloat two-stage waste treatment system. The remaining part of the RBC effluent is diverted to a sedimentation tank for secondary clarification in order to siumulate a conventional secondary RBC-sedimentation waste treatment system for the purpose of parallel comparison. Both Sandfloat effluent and sedimentation effluent are later combined together and discharged to the subsequent sand filter and UV disinfection unit for final polishing. Figure 1 shows the flow diagram of Jiminy Peak Plant. The Massachusetts Clean Waters Act (M.G.L.c.2l,s.26-53) was amended by Chapter 246 of the Acts of 1973 to authorize the Massachusetts Department of Environmental Quality Engineering (DEQE), Division of water Pollution Control, Boston, Massachusetts, to regulate discharges into all waters of the Commonwealth of Massachusetts, including groundwaters. The DEQE regulates discharges through the issuance of discharge permits which impose limitations on the amount of pollutants which may be discharged in the effluent, together with monitoring and reporting requirements and other conditions to insure adequate treatment of all liquid wastes prior to discharge. The DEQE reviewed the application submitted on behalf of Jiminy Peak, Inc. for a permit to discharge sanitary wastes from the Jiminy Peak secondary wastewater treatment system located off Brodie Mountain Road, Hancock, Massachusetts, and subsequently developed the conditions contained in the Groundwater Discharge Permit No. Ø-188. More specifically, Jiminy Peak, Inc. (i.e. the permittee) has been authorized to discharge into the ground from the Jimmy Peak secondary wastewater treatment facilities for which the Permit No. Ø-188 was issued, a treated effluent

Jan 1, 1989

By H. R. Green

The design of slurry launders has usually been based on strictly empirical concepts. An examination of the most common procedures reveals that they do not account for many of the variables that are recognized as Significant for slurry transport. These may include flow rate, volume concentration of solids, solids specific gravity, solids size distribution particle shape, launder geometry, and roughness of the wetted surface. It was decided to develop a design procedure, which would accomplish two things: First: Take into consideration most of the known significant variables and systematize the procedure to assure consistent results. Second: Provide a rational basis for examining and utilizing operating data to refine and improve the system. To accomplish this, it was necessary to develop a basic design concept. This concept has been developed through a complex development history, and yet still appears workable and technically sound. That concept may be outlined as follows: First: The solids transport velocity is the fundamental basis for slurry launder design. Second: The stream configuration, hence the launder size, is integrated with the solids transport velocity so that the actual stream velocity exceeds the solids transport velocity. Third: The launder slope is that which will achieve the required actual stream velocity. It was originally planned to utilize a series of charts and nomographs for the design procedure. A nomograph, based on data given in Taggart's Handbook of Mineral Dressing (1), was developed to determine the solids transport velocity. The velocity thus determined was used with Manning charts to determine the launder slope .and configuration as though the fluid were water. It was planned to utilize a "slope adjustment factor" to increase the slope to compensate for the apparent viscosity of the slurry. All attempts to develop a suitable correction factor failed. This failure was principally due to the complex relationship of viscosity to the required slope. It was then decided to revise the entire approach. Specifically, the present procedure is based on the Darcy-Weisbach flow equation, which is rationally preferable to the Manning equation. The Camp (2) equation is used for the solids transport velocity. The necessity of reference to the Moody curves for the Darcy friction factor is avoided by using the Colebrook and White (3) equation. The resulting mathematics require iterative solutions for several of the unknowns.

Jan 1, 1978

By Alan H. Goode

Of the 15 million tons of glass discarded in municipal wastes annually, about 2 million tons pass through incinerators and can be recovered by ore processing techniques. Although the primary use for this recovered waste glass would be as cullet in the manufacture of new glass containers, this approach is not always practical because of long shipping distances, high freight costs, and high purity requirements. Alternatives to the use of waste glass as cullet would be either as landfill or in secondary products. Although glass is considered a good landfill material, it is still wasteful of potentially valuable raw materials. Secondary products that can be made from waste glass1 include a variety of industrial and construction items. As part of the Bureau of Mines' program in secondary materials recovery, the Tuscaloosa Metallurgy Research Laboratory has conducted research concerned with the development of ceramic building products from waste glass. Materials employed in this research were obtained from municipal incinerator residues fractionated in the Bureau's continuous residue processing plant, College Park Metallurgy Research center.' The studies reported herein3,4 present research results using waste glass in the formulation and production of clay bricks, not only to produce improved products, but to reduce production costs, conserve natural raw materials, and most important, to reduce manufacturing energy requirements.

Jan 1, 1974

By O. E. Brown

The measurement principle of the Coulter Counter for particle size analyses is described. The application of the Coulter Counter to cement particle content and size analysis and to the size distribution of clinker from standard grindability tests are described. Methods of calibration, sample preparation and data reduction are presented. The results of the analyses obtained are compared with characteristics associated with performance showing both concordance and non-concordance of test data., Future uses of the Coulter Counter for particle size analysis in research and plant control adaptations are indicated.

Jan 1, 1960

By J. B. Winter

The present Inspiration Consolidated Copper Company Smelter/Acid Plant complex located near Miami, Arizona, was designed and built specifically to conform to the ambient air quality standards imposed by the Federal Clean Air Act of 1971. The present plant replaced a facility which had been in operation since 1915. The electric furnace, siphon converters and acid plant complex were commissioned in May, 1974 (1). At that time, the Plant was considered to be one of the most extensively monitored and instrumented facilities of its kind in the world. Several control areas were located in close proximity to the controlled process, in addition to so-called central control rooms in the Smelter and Acid Plant. The control rooms were staffed with officials below the foreman level. Instrumentation consisted of conventional analog devices, except for a computer used in monitoring ambient air environmental conditions and limited use of programmable logic controllers. The harsh service conditions prevalent in the plant rendered a significant portion of the instrumentation unreliable and, in many cases, inoperable. Failure of the facility to obtain the total sulfur collection for which it was designed was, in part, due to serious inadequacies in control system logic and sulfur recovery averaged 79% in the early years. After careful study of the smelting process, engineering hardware and high labor cost inherent in the organizational structure, management made a judgement decision to convert to centralized control of the process with high level supervision 24 hours per day. A distributed digital control system with supervisory arid data acquisition/information reporting computers was selected to satisfy the requirement of central control. Operations are now controlled and coordinated from a primary control room in the Smelter and a satellite control area in the Acid Plant.

Jan 1, 1983

Prior to 1972, Peko Mines Limited at Tennant Creek used a central mill at Peko, which consisted of a gravity circuit, flotation and cyanidation of tailings, to treat their gold ore from surrounding mines. The Warrego concentrator, 60km to the west, was built to handle the Warrego and Gecko copper orebodies, which were being developed concurrently with the concentrator construction The Warrego concentrator consisted of a flotation circuit, using autogenous grinding at 100 - 120 tph, compared to the more elaborate Peko circuit employing rod mill - ball mill grinding capable of treating only 18 - 26 tph. During development of the Warrego copper mine a gold rich ore pod was located, because of which the mining plans were adjusted to allow simultaneous, separate extraction of the copper and gold ores. However, the plans for the concentrator were not affected as the gold ore could be carted 60km to the Peko central mill for processing.

Jan 1, 1978

By W. T. Millar

Graphite in the Suffern deposits near Hillburn, N. Y., consists of flakes ranging in size from fines to one-half inch in diameter. The Bureau of Mines investigated the Suffern deposits in the spring of 1943. Gneiss outcrops were examined, and five trenches excavated and sampled. The graphite is not confined to any definite zone but occurs in small, local pockets. Eleven samples were collected and combined for laboratory tests. Tests by the Bureau of Mines indicate that the graphite ore can be beneficiated by stage crushing, "skin flotation" of coarse flakes, and wet flotation of minus 10-mesh flakes. Over 90, percent of the graphite was recovered.

Jan 1, 1949

By Rajesh Sonkar, Sanjeev Kumar

The deployment of Dragline machines in opencast coal projects involves huge capital expenditure as well as operating costs. The mine operators always endeavor to ensure smooth excavation of material by the Dragline machines while meeting the major objectives of safety, minimal disturbance to environment and optimal operating costs. Optimization of fragmentation through well designed blasts is fundamental in meeting part of the objectives is good fragmentation of rock or mineral proposed to be lifted by the Dragline machine. If the overburden lying over mineral is not properly fragmented, it may result in formation of ‘Toe’ the removal of which is the most arduous task in opencast coal mines. The mine operators always endeavor to achieve best possible results in blasting to ensure smooth operation of the high cost Dragline Machine. However, despite continuous efforts by the blasting engineers involved in bench blasting for the machine, the formation of ‘Toe’ is an ever nagging problem faced in large opencast coal projects eventually reducing the overall operational efficiency of mining.

Jan 1, 2011

Continued discussion of the paper of R. W. Raymond, presented at the Cleveland meeting, October, 1912, and printed in Bulletin. No. 70, October, 1912, pp. 1111 to 1122. See also discussion printed in Bulletin No. 74, February, 1913, pp. 319 to 329. R. B. BRINSMADE, Puebla, Puebla., Mexico (communication to the Secretary *) :-Though I heartily approve of several minor points emphasized by Dr. Raymond (such as the general desirability of local administration, rather than Federal, and the evils inseparable from the Federal Apex Law), I feel obliged to dissent from his four final propositions. As Dr. G. 0. Smith has already discussed these propositions from the standpoint of the experience of other countries, I shall base my criticism upon certain general economic principles, which will be grouped under the following five topics: 1. The dual theory of property. 2. The essential factors in fixing prices. 3. The relation of land ownership to the distribution of wealth. 4. The relation of land ownership to speculation and development. 5. The relation of land ownership to the political power. I shall give here only a brief outline of these topics, as they are fully treated in the standard treatises on political economy.1

Jan 5, 1913

By J. F. Lewis

Believing that one of the essential points in mining, as in all other business, is to know the expense incurred in each particular department, I have carefully kept an account with each department for the past three years; and, as it may be of interest to some of our members, I herewith present it to the Institute, reduced to per centages of the total yearly expense, showing the cost for materials, repairs, expenses, tools, team accounts, powder, oil, and fuel, also for labor under its several heads: 1875—Materials. 1875—Labor, Repairs,...040 Expenses,....016 Team,. .. .. .017 Machinery,. .. .055 Tools,..006 Repairs,. .. .027 Improvements,.....030 Washer,. .. .076 Oil ..008 Pit,.. 178 Powder,..,043 Dirt,.. 130 Fuel,. .325 Superintendence,. ..050 .485 .516 100.1 1876—-Materials. 1876-Labor. Repairs,. . • .. .043 Expenses,......012 Team,.011 Machinery,. .. .090 Tools,..005 Repairs,... 027 Improvements,.....007 Washer,. .. .085 I,. ...,015 Pit,.. 139 Powder,. .. .. .030 Dirt,.. 160 Fuel,...321 Superintendence,. ..078 .444 .559 100.3 1877—Materials. 1877—Labor. Repairs,......102 Expenses,......019 Team,..031 Machinery,. .. .050 Tools,..030 Rrpairs,. .. .046 Improvements,....,031 Washer,. .. .057 I,...009 Pit,.....179 Powder,...035 Dirt,.. 162 Fuel,...200 Superintendence,. ..054 .457 .548 100.5 Average for the three years 1875, '76, and '77. Materials. Labor. Repairs,...062 Expenses,...016 Team,..016 Machinery,. .. .065

Jan 1, 1879

By V. L. Brant

The estimated total energy demand for the U. S. is 60 quadrillion BTU for 1970 and is expected to more than double by the year 2000. Today the energy crunch is here. It is evidenced by brownouts to conserve electrical energy, and the shortage of gas and oil. Many such headlines, including suggestions for energy rationing, are in the news today. Paradoxically we have an abun¬dant source of energy in the form of coal. Energy consumers, however, are geared to use fuel in other forms. It is the intent of this presentation to discuss processes under development for conversion of the energy contained in these vast coal reserves to more usable forms. Coal conversion to synthetic fuels should be of particular interest to those engaged in mining. Coal consumption in a single conversion plant is expected to range from 10,000 to 40,000 tons per day. Plants will be located at the mine mouth and will operate continuously 24 hours per day, seven days per week. Further, many such plants will be needed., For example, looking at natural gas only, the supply/demand gap is estimated to reach 13 trillion cubic feet per year by 1985. Assuming that gas from coal conversion were to fill only 101, of this gap, 15 plants producing 250 MMSCFD would be required. These plants would consume about 90 million T/Y of coal or 135 million T/Y of lignite. It will indeed be a challenge to mine, suitably prepare, and transport coal at these projected volumes without interruption. Many ideas have been conceived for better utilization of coal and these are in various stages of development. Those supported by public money are covered by the Annual Report of the Office of Coal Research (OCR), Department of Interior. The two processes to be discussed, the Consolidation Coal Company's (Consol) C02 Acceptor Process (1)(2)(3) and The Pittsburg and Midway Coal Mining Company's (PAMCO) Solvent Refined Coal Process (4)(5) are both sponsored by OCR.

Jan 1, 1970

By Jack D. Ramsdell

Electrorefined, magnesium-reduced, and sodium-reduced titanium samples were evaluated for recrystallization behavior, effect of cold rolling on tensile properties, and bend ductility. Curves were determined for the softening of 2.5 to 80 percent reduced cold-rolled metal that was annealed over a temperature range of 300° to 800° C. Upon annealing, coarsening of the grains was found to occur for reductions in thickness of less than 20 percent in titanium made by all three processes. The presence of metallic impurities greatly inhibits coarsening for reductions in thickness below 20 percent, but for reductions in excess of 20 percent grain size is essentially constant at each annealing temperature. The three grades of titanium did not harden excessively from cold rolling; both tensile and bend ductility were good for reduction in thickness up to 80 percent. Upon annealing, softening occurs between 400° and 600° C., and, based on microstructure and property studies, full annealing is accomplished after 30 minutes at 700° C. for the magnesium-reduced and electrorefined titanium and at 800° C. for the sodium-reduced titanium.

Jan 1, 1960

By S. M. Runke

An ever-increasing demand for beryllium and beryllium alloys has directed attention to the milling of low-grade domestic deposits as a possible means of supplying part of the Nation's ore requirements. The United States generally produces less than 10 percent of its annual supply. For example, new supply in 1953 was estimated at about 9,250 tons, of which the United States imported 8,500 tons. Domestic production was only 750 tons.2/ Since record high prices for beryl in 1952 and 1953 did not greatly increase domestic product ion, the United States must develop mechanical beneficiation processes to use low-grade deposits and reduce dependence on foreign sources

Jan 1, 1954

By T. H. Alden

BaCKOFEN, Avery, and Turner have shown"2 that the large tensile elongation in superplastic metals is correlated with a high strain-rate sensitivity of the flow stress. At present, the reported superplastic materials are complex eutectic or eutectoid systems with a large solubility range in at least one of the terminal phases. This note describes the behavior of a solid-solution alloy (at 22°C)5,= Sn-1 wt pct Bi, which has been processed in order to produce a fine grain size.2,3,6 The chill-cast, 1/2-in.-diam ingot was homogenized for 1 week at 130°C, extruded at O°C to 0.083 in. diam, quenched within 15 sec, and stored at -196°C. Samples were warmed to room temperature 16 hr prior to mechanical testing or metallography. During tension testing at room temperature, it was found that the flow stress reached a steady-state value which depended on the imposed elongation rate and was independent of strain to an error of about ±5 pct. In Fig. 1 is plotted log a vs log e where a is the measured steady-state flow stress and i the imposed strain rate. These quantities were calculated using "instantaneous" area and length to an accuracy in the sample diameter of ±5 x 10-4 in. The log a-log 1 curve is sigmoidal. The strain-rate sensitivity, which may be given by d log old log 6,' has a maximum value of about 0.48 between i = 5 x 10-4 and 5 x 10-3 per min. The value is well above frequently measured values for metals at creep temperatures (0.1 to 0.3) and comparable to the rate sensitivity for Pb-Sn solder: which has been shown to be superplastic.2,3 Note, however, that below or above the strain rate of 10-3 per min the rate sensitivity of Sn-1 pct Bi decreases and at high rates is more nearly like that of conventional metals. The grain size of this material is relatively uniform, Fig. 2, and may be specified by an average linear intercept i.' Grain growth occurs at room temperature but after 16 hr, L = 5. Tensile elongation was measured on finer-grain-size specimens, which had been at room temperature 15 min prior to testing. After 15 min, the samples are fully recrystallized. Three strain rates, 5 x 10-3, 10-2, and 2.5 x 10-2 per min, were employed for which the measured elongations varied from 425 to 500 pct. These results would seem to justify the term superplastic, applied to this alloy, although the ductility is inferior to that of the Sn-Bi eutectic.3 The Sn-1 pct Bi alloy, which is a single-phase material with equiaxed, fine grains, is an excellent one on which to make an estimate of the Nabarro-Herring9 creep rate. According to the model of Avery and Backofen2 a substantial part of the deformation must be accomplished by N-H creep in the strain rate-temperature regime in which d log a/d log 1 = 0.5. The maximum N-H creep rate is given by9 L2kT where v is the atomic volume and D the self-diffusion coefficient. For this test v = 3.O x 10-23 cu cm, D = 1.4 e-23,300/RT = 2.5 x 10-l7 sq cm per sec,7 L = 5x 10+4 cm, T = 295°K, and a = 2000 lb per sq in. = 1.38 x 10' dyne per sq cm. The calculated creep rate is E= 4 x 10-11 per sec which may be compared with the measured creep rate at this stress, Fig. 1, of 6.7 x 10-5 per sec. The N-H creep rate is seen to be negligible. Earlier consideration2 of this question was in error because incorrect diffusion coefficients were used. The assistance of R. R. Russell in metallography is acknowledged gratefully.

Jan 1, 1967