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By H. C. Fuller, E. L. Anderson, F. S. Wartman, M. A. Cook

"This paper gives an account of the production of ductile-grade titanium powder in 100-pound batches by reduction of purified titanic chloride with magnesium, followed by grinding, leaching, and magnetic separation of the reaction product. Apparatus, technique, and methods of controlling product quality are described, and some data on operating costs are given.Although man has been aware of the existence of the element titanium for over 150 years, only recently has the metal itself been available in a form pure enough, particularly with reference to absorbed gases, to allow an estimate of its physical properties. In comparison with mild steel, the tensile strength of titanium is about the same, the resistance to atmospheric corrosion greater, and the specific gravity much less.These properties suggest the usefulness of titanium as a structural material for applications requiring resistance to atmospheric attack and a good strength-weight ratio. Although known reserves of titanium, high-grade enough to be useful at present are not sufficiently extensive to predicate that titanium might supplant iron, aluminum, and magnesium for structural purposes, the reserves are large enough to make its use as a structural material quite feasible in certain fields in which its unique properties are required.As a result of these considerations, the Bureau of Mines station at Tucson initiated an, investigation into the technology of titanium in 1938. First, a critical survey and small-scale trial of virtually all suggested methods7/ for producing the metal was made. The method devised by Kroll8/ seemed best adapted as a source of metallic titanium for, although the purity of the product was not as high as that of products from some other processes, it seemed that modifications in technique necessary for larger-scale production could be made more easily.Essentially, in applying the Kroll process, titanium tetrachloride was reduced to metallic titanium by adding the chloride slowly to molten magnesium contained in a molybdenum-lined iron reaction chamber filled with a protective atmosphere of argon. After reduction to the state of completion desired, the reduction mass and chamber were allowed to cool. Then the chamber was opened, and the reaction mass, consisting of titanium, magnesium chloride, and some unused magnesium, was removed by suitable means. Most of the magnesium chloride and metallic magnesium were removed by leaching in dilute hydrochloric acid. The leached sponge was then ground wet in a ball mill, releached in strong acid, and dried Removal of free iron by magnetic separation completed the production process.Development of the Kroll process by the Bureau was undertaken first at the station at Salt Lake City, in 1942 and the amount formed per run was increased in several steps from a few grams to pounds.9/ In addition to increasing the size of batch over that used by Kroll, the technique was simplified somewhat by the use of unlined reaction vessels."

Aug 1, 1949

By W. A. SLOAN, G. L. Oldright, J. D. Sullivan, E. K. PRYOR, A. F. HALLET, R. S. Dean, S. L. BROWN, F. S. Wartman, A. J. Thompson

Lack of sufficient publication funds has resulted in the accumulation of considerable unpublished experimental work on copper metallurgy by the Bureau staff . The present report gives extended abstracts of this work , as well as a discussion of the best direction for future efforts . The experimental work reported in this series of papers was done in cooperation with the University of Arizona , Tucson , Ariz . The investigations covered in the third , fifth , and seventh papers were done under the direction of G. L. Oldright and the remainder under the direction of V. H. Gottschalk .

May 1, 1934

By C. W. Davis

During the second year of its existence, the Ore-Testing Section has actively prosecuted work in the standardization of testing methods, in conducting routine tests and analyses for other Government agencies, and in the determination of recommended processes for the treatment of representative ores from mining districts. Close cooperation has been maintained between the members and units of this Section and those of other Sections, J. A. Woolf and J. B. Clemmer have continued to act in an advisory capacity and R. E. Head has contributed valuable assistance in the microscopic examination of ores and metallurgical products that did not respond readily to standard treatment methods.

Jan 1, 1938

By C. W. Davis

During the second year of its existence the Ore-Testing Section has actively prosecuted work in the standardization of testing methods, in conducting routine tests and analyses for other Government agencies, and in the determination of recommended processes for the treatment of representative ores from mining districts. Close cooperation has been maintained between the members and units of this Section and those of other Sections. J. A. Woolf and J, B. Clemmer have continued to act in an advisory capacity and. R. E. Head has contributed valuable assistance in the microscopic examination of ores and metallurgical products that did not respond readily to standard treatment methods.

Jan 1, 1938

By R. L. Grayson

In 2010, MSHA began holding hearings on a new coal mine respirable dust rule, which proposes a 1.0 mg/m3 standard, a separate 0.1 mg/m3 quartz standard, and the use of single-shift samples for compliance. Small mines largely comprise the ?hot spots? where dust diseases of the lungs increased dramatically over the past decade. This paper presents findings on the feasibility of small mines? compliance with the proposed new 1.0 mg/m3 standard. It gives probabilities of noncompliance by MSHA district, seam height, and selected ?hot spot? counties. Because silicosis and respirable quartz dust appear to be the primary issues related to a recent doubling of the prevalence of dust diseases of the lungs, multiple regression models for prediction of the average total quartz content of compliance samples are given. These findings have significant implications on the need for better dust control methods.

Jan 1, 2012

By E. G. King

Thermodynamic calculations involving the oxides of tungsten and molybdenum have been hampered by lack of accurate entropies and high temperature heat contents. This paper presents work by the Federal Bureau of Mines on new low-temperature heat-capacity data leading to the evaluation of the 298.15° K. entropies of tungsten dioxide and tungsten trioxide, as well as high-temperature heat contents (above 298.15° K.) of molybdenum dioxide, molybdenum tri-oxide, tungsten dioxide, and tungsten trioxide. Included are heats of fusion of the two trioxides. The heats of formation at 298.15° K. of these four oxides are tabulated as measured by Mah4 5/ of this laboratory. Included also are the 298.15° K. entropy of Mo02 measured by Kin 6/ of this laboratory, and that for Mo03 according to Kelley and King.7/ These basic data provide a consistent compilation from which to perform accurate thermodynamic calculations involving these compounds. No previous low-temperature- heat-capacity measurements have been reported for tungsten dioxide. Seltz, Dunkerley, and DeWitt8/ measured the low-temperature heat capacity of tungsten trioxide. The only previously-reported high-temperature heat-content measurements for any of the oxides were conducted by Cosgrove and Snyder,9/ who measured molybdenum trioxide to 1,300° K.

Jan 1, 1960

By Timothy L. Roberts

For many years practicing geotechnical and structural engineers in Texas and Louisiana were reluctant to include augered cast-in-place (ACIP) piles as a viable deep foundation alternate. Three primary concerns were the lack of a final driving resistance to confirm adequate capacity, the inability to inspect ACIP piles prior to installation, and the fact that pile integrity is highly dependent on the skill of the ACIP pile contractor's field personnel. The Deep Foundations Institute (DFI) Committee on Augered Cast-In-Place Piles has responded to these quality control concerns by establishing standards of practice for ACIP pile installation and ACIP pile inspection. Regional areas of the country have different subsurface soil and rock conditions and many times this will influence local installation and inspection requirements. The best way structural and geotechnical engineers can address quality control concerns is to be familiar with the DFI standards of practice, to know the special requirements for a particular project site, and to incorporate them into a well-written ACIP pile specification. The purposes of this paper are to: 1) review the main fundamentals of a successful quality control program, 2) describe how these fundamentals can be applied to the installation of ACIP piles, and 3) add personal experiences from ACIP pile projects in Texas and Louisiana. Nondestructive test methods are frequently used as a quality control tool to evaluate questionable piles. An additional brief discussion of two nondestructive test methods is presented to complete the paper.

Jan 1, 1998

By Matthew N. Plis

This U.S. Bureau of Mines report presents the documentation for COALVAL, a coal property evaluation software package developed on Lotus 1-2-3,4 version 2.2, spreadsheets. The software is compatible with version 3.1 as well, and may provisionally be run on the earlier 2.01 version. COALVAL is a menu-driven program that produces a prefeasibility-level cost analysis of mine-planned coal resources. The package contains cost models for each of five coal mining methods commonly employed in Appalachia: auger, contour strip, mountain top removal, continuous miner, and longwall. Other models, such as a dragline cost model, will be incorporated as the Bureau's Coal Recoverability Program matures. COALVAL allows mine operators, evaluators, consultants, and Government entities to input resource data and the various production, operating, and cost variables that pertain to their property. The program can evaluate up to 25 seams, each to be mined with up to five different mining methods, within a given area. Summary spreadsheets listing the cost per clean ton to mine the resources, f.o.b. the tipple, are produced for each property, seam, and mining method/seam combination.

Jan 1, 1993

By Steven W. Hunt, Safdar A. Gill

Design and construction of temporary shafts for underground facilities often require evaluation of bottom stability due to rupture (blow-ins) caused by water pressure acting on a bottom plug. This paper provides an overview of shaft bottom stability evaluation methods for the case of a relatively watertight soil plug which overlies an artesian aquifer. Methods of considering plug perimeter shear, plug buckling and pressure relief are discussed. Minimum factors of safety are suggested. Future use of the shaft subgrade is also considered. Bottom plug failure case histories at several shafts are evaluated. This paper should be useful to project designers, ground support system design engineers and contractors involved with construction of shafts for underground facilities

Jan 1, 1997

By J. F. Breedis

The stabilization of austenite in Fe-Ni alloys against martensitic transformation after reversion has been attributed' to the lattice imperfections remaining from previous transformation. More recently, however, such stabilization has been associatedzy3 with compositional changes which occur in martensite during slow heating. Dislocation arrangements in deformed austenitic stainless steels strongly influence transformation to martensite on quenching.4 Therefore, it is of interest to observe the transformation behavior of stainless steel after reversion, especially since previous studies have been confined to Fe-Ni alloys. The transformation-product morphology also differs in the two iron-alloy systems;6 plate martensite forms in the latter while stainless steel transforms to lath-type martensite. Samples of Fe-16 wt pct Cr-12 wt pct Ni (maximum C + N = 0.005 wt pct, 0 = 0.02 wt pct), prepared by vacuum melting and casting from high-purity components, were annealed at 900°C for 24 hr, after cold rolling to produce a 40-p grain diameter. The M, temperature of this alloy is -80°C and samples are fully austenitic after annealing. The amount of martensite present after quenching, annealing, and requenching was determined magnetically, to within 0.01 wt pct, using a vibrating specimen magnetometer. Nearly 14 pct of martensite forms during the initial quench in 1 min at -196°C. Two-min and 2-hr anneals were done in air and in evacuated quartz capsules, respectively. All specimens were electro-polished before making the magnetic measurements.

Jan 1, 1967

By Steven G. Vick

Effects of the August 1995 failure of the Omai tailings dam in Guyana continue to be felt throughout the mining industry. This article is taken from the final report of the independent body of experts commissioned by the government of Guyana to establish the cause of the failure. It explains what happened, why and how to avoid it in the future. Despite environmental, health and safety effects that were negligible by any objective measure, the failure has received wide international attention. Coming on the heels of the shift in exploration targets and development opportunities to developing countries, it has intensified interest in tailings dam safety and environmental management issues. An informed understanding of the nature of the failure and its effects is key to responding to these concerns.

Jan 1, 1996

By Hyung-woo Kim, Kiyoung Sung, Sup Hong, Jung-Yeul Jung

Many concepts for the commercial production of deep-seabed manganese nodules have been studied since the 1970s. However, the bearing capacity of the deep seabed is not strong because it is composed of fine particles with high moisture content. So it is impossible to carry manganese nodules in the collection system. Therefore, the validity of continuous mining by riser pipe from the seabed to the vessel is highly appreciated. A continuous mining system is composed of mining vessel, lifting pipe, buffer system, transport tube (flexible pipe), and self-propelled mining robot. A conceptual diagram of this system is shown in Fig. 1(a). Korea Research Institute of Ships and Ocean Engineering (KRISO) developed a deep-seabed self-propelled mining robot called "MineRo" to collect manganese nodules in 2012 (as shown in Fig. 1(b)). In 2013, real sea experiments at around 1380 m depth were carried out to evaluate the collecting and driving performance of MineRo. However, we found that the roller in MineRo did not work well at times. To elucidate the problem, we analyzed the rheological characteristics between the bushing and drive shaft under deep-water conditions.

Sep 1, 2014

By D. W. Davis, P. Landais, L. J. Robb

Abstract -Nodules of carbonaceous matter have been observed in several Archaean peraluminous granites adjoining the gold- and uranium-bearing Witwatersrand Basin. These nodules, which resemble solid bitumen, usually replace uraninite or uranothorite and are paragenetically late. They show many similarities to kerogen seams and nodules associated with the conglomerate horizons in the Witwatersrand Basin, including composition, maturation, intimate association with uraninite, and a late paragenesis. U-Pb isotope ratios for granite-hosted and Witwatersrand kerogens define a single, albeit somewhat diffuse, Pb-loss discordia suggesting an upper intercept age of circa 2300 Ma. Carbon isotope analysis indicates lighter ?13C values for the granite-hosted nodules than for the sediment-hosted kerogen seams. Raman spectroscopy indicates that granite-hosted kerogen is more highly ordered than Witwatersrand seam kerogen. A model is presented which links the granitoid­ and sediment-hosted organic matter to an event of oil generation in the Witwatersrand Basin circa 2300 Ma ago. Light hydrocarbons migrated through the sediments and into underlying/adjacent granitic rocks, undergoing polymerization and accumulation in response to radiation accompanying the presence of high-uranium phases.

Jan 1, 1994

By P. -E. Boileau, P. Marcotte, Y. Aldien, S. Rakheja

A number of biodynamic models of the hand-arm system have evolved on the basis of measured driving-point mechanical impedance (DPMI) responses to facilitate analyses of the coupled hand-tool system [1]. The parameter identifications in such models are based upon minimization of an error function of the model and the target impedance data, which may not yield a unique solution. Consequently, a number of model structures and parameter sets could be realized that would equally satisfy the target curve. Moreover, the vast majority of the reported models exhibit acute deficiencies due to excessive static deflections of model masses, presence of a low frequency mode and very light masses in the order of 1.2- 4.8 grams. The models also do not characterize the dependency of the biodynamic responses on many factors, namely the hand forces, hand-arm posture and vibration intensity. This study aims at development of a hand-arm biodynamic model with considerations of the hand forces, and both the DPMI and power absorption measures, to enhance the uniqueness of the model. Methods Two different model structures are chosen for identifying the model parameters on the basis of measured DPMI and absorbed power characteristics of the hand-arm system under zh-axis vibration over a range of hand-grip and push forces. Owing to the strong influence of the hand-handle coupling forces, the models were initially derived for fixed hand forces, namely 30 N grip and 50 N push forces, as suggested in the ISO 10068 standard [2]. The equations of motion for the model are formulated and solved to compute both the DPMI and absorbed power responses. A constrained minimization function comprising weighted errors of both the DPMI and absorbed power is formulated and solved to identify the parameters. Alternate functions corresponding to different combinations of hand forces are then applied to identify hand-force dependent model parameters. Variations in the model parameters are investigated as functions of the grip, push and coupling forces through linear regression analysis. Regression-based models are formulated for deriving the hand-handle forces dependent model parameters. The validity of the model is also examined under selected combinations of hand forces. [ ]

Jan 6, 2006

By S. I. Angadi, C. Eswaraiah, Sunil Kumar Tripathy, T. Santosh, Nilima Dash

Banded hematite quartzite (BHQ) and banded hematite jasper (BHJ) ores represent a promising potential iron ore resource in the near future. The hematite and quartz in BHQ and the hematite and jasper in BHJ are closely related and require fine grinding for the liberation of hematite phases. The present study investigates the mineralogical features and comminution properties of BHQ and BHJ ores. In the BHQ ore, it is observed that thick bands of hematite have fine inclusions of silicate minerals, whereas silica bands have few grains of iron-bearing minerals. The hematite and quartz minerals in the iron and silicate bands of the BHJ ore appear to be more intricately associated and complex in nature, presenting difficulties for liberation. Bond work index studies were carried out for the BHQ and BHJ ores with different product fineness in order to evaluate the energy requirements in fine grinding. A higher Bond work index value was found for the BHJ ore (15.4 kWh/ton) than for the BHQ ore (12.4 kWh/ton). A decrease in the product particle size was associated with an increase in the work index value. Single-particle breakage experiments were designed by varying feed particle size and input energy. The specific energy consumption for the BHJ ore was higher than that for the BHQ ore. The comminution parameters showed that the BHJ ore is harder to grind than the BHQ ore, which may be due to the mineral composition and association of hematite phases with the silicates.

This modern plant consists of a battery of 72 Otto Wilputte Compound Under-jet Ovens with extensive plant for the recovery by-products. The ovens produce 9000 tons of furnace coke per week together with some 17,000,000 cub. ft. of coke oven gas, which is used for heating purposes throughout the works and for heating the ovens.Coal is supplied from the, company's colliel'ies and is stored under a Coal Handling Bridge of 264 ft. span and 365 ft. overall, capable of handling 250 tons per hour. The coal storage space of 1600 ft. can. be extended to a maximum of 2800 ft.Preparation of the coal for coking involves fine grinding which is carried out as follows:-Coal is dumped from rail waggons on to a conveyer belt which delivers it to a 12 ft. x 22 ft. Bradford Breaker, with a capacity of 400 tons per hour. The coal, then approximately 4 in. in size, passes, to the mixing bins; and after careful blending is reduced in the,hammer mills to a fineness suitable for the ovens.The fire coal is elevated to the main storage bin over the battery for subsequent charging by electric larry car. The average cokingtime is 15 hrs. 36 mins. At the end of ,this period the doors, are removed and the coke pushed from the oven by a mm. The coke, is, then quenched with fresh, water and sent to the Blast Furnace in 50-ton Copper Waggons.The gas is exhausted from the ovens by Brown Boveri turbine, exhausters of 23,000. cub, ft, per min. capacity. Exhaust steam from these is used for various. purposes in the By-Product plant. After removing the last traces of tar by Electrostatic Precipitators', the gas is bubbled through sulphuric acid, the ammonia forming ammonium sulphate, which is subsequently washed, dried and bagged for sa,le as a fertilizer. The gas then goes to the Benzol Scrubber Towers, where benzol and other derivatives are recovered and later redistilled and purified. The cleaned gas is then piped to the main works for heating purposes, being used with tar to fire the steel furnaces, and also the soaking pits and reheating furnaces.BLAST FURNACES.The company is at present operating two Blast Furnaces of 800, and 1000 tons nominal daily capacity respectively. The latter was...

Jan 1, 1938

By H. Lyn Bourne

This marks the 15th consecutive year for this column. During 1995, readers were asked to supply royalty data for construction raw materials especially for states where none is available. The response was discouraging and royalty data are still missing for several states for either sand and gravel or crushed stone. Several readers have said that this article has merit but more new information is needed to justify its continuation. Several articles and pieces of news relative to mineral royalties appeared in the trade journals. R. Lee Aston (E&MJ, November, 1995, p. 16N) described South Australian royalties as amended by the government in1994. The royalty rate (a percentage) was lowered but the assessed value for minerals was raised and the current rate amounts to 20 cents//t (18 cents/st).

Jan 1, 1996

By Brad Thrall

"Viceroy Resource Corporation of Vancouver, B.C. has recently developed the Brewery Creek Mine, an open pit heap leach gold mine 55km East of Dawson City in Yukon, Canada. Mineable reserves are 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797000oz) of gold. Ore processing employs the heap leach technology using run of mine ore. 1996 highlights included construction of 158,000 m2 of leach pad, placing 1.9 million tonnes of ore at a grade of 1.46 gpt on to the leach pad, construction and successful commissioning of the ADR plant, and pouring of the first gold bar on November 15th.Permitting this mine has drawn significant attention because the heap leach technology, while proven in other jurisdictions, has not been used in Yukon where winter site temperatures routinely reach -45 C for extended periods.IntroductionFollowing a comprehensive governmental and public review of the proposed project, a Class A Yukon Water license was issued in August, 1995. Construction started in 1995 and leaching commenced in October 1996. The Brewery Creek Mine is in Yukon, the home of the Klondike Gold Rush nearly 100 years ago. This is the only part of Canada not glaciated in the Wisconsin advance so there is oxide ore resulting from about 80,000 years of weathering -unique in Canada. Discontinuous permafrost is present on site and has required special consideration in the design. Winter temperatures dip below minus 50 C, demanding an approach to mining and leaching that is different than in more temperate climates.Project DescriptionThe mine comprises 8 shallow, largely oxidized gold deposits along a 7km corridor. An 8 year mining program will extract 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797,000 oz) of gold. The strip ratio will be 1.35: 1. Overall gold recovery is projected at 78%. There will be 2.3 million tonnes of ore placed on the pad annually, with an average of 2.57 tonnes (83, 000 oz) of gold recovered per year. Because of the cold winters, the mine will operate approximately 210 days per year, while the leaching process will operate on a year round basis. Run of mine ore is placed on the heap 150 days per year. The recovery process utilizes conventional sodium cyanide leaching , carbon adsorption, caustic desorption, and on-site smelting. Peak employment during mining is 145 and winter staff is reduced to 45."

Jan 1, 1997

By Francisco Alberto Marreiros da Rocha Filho

No processo de laminação a quente ocorre ocasionalmente a perda de ductilidade a quente que pode causar o fenômeno denominado como “abertura de cabeça” da barra durante a sequência de passes. Essa perda de ductilidade a quente pode ser severa e causar interrupções no processo produtivo. Esse fenômeno pode estar associado a diversos fatores entre eles: composição química do aço; defeitos presentes no tarugo; condições do processo de laminação; e temperatura do aço durante o processo, principalmente nos passes de desbaste e intermediários de laminação. No processo em estudo, a perda de ductilidade a quente ocorre associada à laminação de vergalhões a partir de tarugos oriundos diretamente do lingotamento contínuo. O presente trabalho tem como objetivo avaliar a influência da temperatura e da microestrutura formada ao longo do processo na perda de ductilidade a quente das barras através de um exame comparativo entre os vergalhões laminados com e sem o uso da etapa de reaquecimento.

Oct 30, 2017

By P. Lozada

The objective currently demanded from controlled blasting techniques is to minimize damage to the remaining rock mass while protecting the integrity of mine walls and ensuring the quality of the fragmented material for easier extraction. Finding this balance represents one of the three main challenges faced by large open-pit mines. They must achieve mine stability by complying with slope designs while also maintaining good results in terms of fragmentation and equipment productivity. The main constraints are related to unfavorable geological conditions, water presence, limited drilling diameters and available explosive products, as well as managing drilling times and blast scheduling, among others. For slope control, in addition to presplitting, blasting designs generally include buffered rows, drilled with smaller diameters, and loaded with fewer kilograms of explosives than production holes. The objective is to reduce the powder factor (kg/t) to minimize the energetic influence on the walls and attenuate damage. However, this often requires using very short charge lengths compared to the bench height (one-third or less) and reducing the pattern geometry, which increases the number of holes per ton and drilling costs. This traditional technique of reducing powder factor in the buffered rows has the main limitation of not being able to distribute the explosive energy throughout the entire column, leading to fragmentation issues mainly in the steaming area. Measurements conducted by Orica in their operations indicate at least a 30% increase in P80 particle size in the buffer zone associated with this condition. To improve energy distribution, techniques such as decking, or vertically decoupled charges can be employed. However, these techniques cannot be applied under wet conditions, and strict control over implementation quality is required to avoid deviations, which increases loading operational times and resource required. In this sense, the use of less energetic and lower-density explosives is a better alternative. On one hand, they allow for complete charge columns in the buffered rows, ensuring homogeneous fragmentation, and on the other hand, they reduce the energetic influence on the slopes, significantly decreasing vibration levels. With less energy per kilogram of explosive (MJ/kg) and less charge per linear meter (kg/m), it becomes possible to handle higher specific charges and longer explosive columns, providing flexibility in drilling and loading design options without compromising slope stability.

Jan 1, 2024