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By Iga Trisnawati, Fitria Puspita, Himawan Tri Bayu Murti Petrus, Gyan Prameswara, Idi Amin, Andi Nurfaadhilah Ulfah

Laterite ores, rich in essential metals like nickel, iron and cobalt, are a crucial resource for modern technology. However, the extraction of nickel from laterite ores is a complex process due to their intricate mineral compositions and low nickel concentrations. This paper presents the results of a study to understand the leaching behavior of preroasted laterite ore and the impact of roasting on leaching kinetics. The ore was subjected to roasting at 280 and 610 °C for a duration of three hours, followed by leaching in sulfuric acid (H2SO4) under atmospheric conditions. The results show that preroasting the ore at 610 °C significantly improved the recovery of nickel during leaching, with recovery rate of up to 93 percent, and the leaching process of the preroasted ore was primarily controlled by diffusion through the ash layer. The activation energy values were found to be 20.97 kJ/mol for the ore roasted at 280 °C, and 18.97 kJ/mol for the ore roasted at 610°C. These findings suggest that preroasting can enhance the leaching performance of laterite ore, potentially paving the way for more efficient and sustainable nickel extraction processes.

Aug 1, 2024

By T. F. Oliveira, E. B. Zanelato, J. Alexandre, L. J. T. Petrucci, A. R. G. Azevedo

The paper and cellulose industry is responsible for generating large amounts of solid waste, which are produced during various stages of the production process. Currently this material goes to places like landfills and other generating costs for the company. Furthermore, cement, mortar and concrete represent a significant economic and environmental cost. This study aims to determine the pozzolanic activity index in the residue studied to verify its viability in the incorporation of the same in cement. It will be used Luxan method for this determination. The data found show that the material has the possibility of being incorporated in the mortar due to its pozzolanic activity index.

Mar 1, 2017

By P. Braga, A. Chaves, S. França, A. Luz

"In Brazil, there are no official reserves of molybdenum, and the unofficial reserves are modest. Therefore, the molybdenum ore (molybdenite) produced as a co-product or by-product of emerald mining in the region of Campo Formoso, BA, has arisen as the only economic option. The flotation process is commonly used in Brazil for the recovery of molybdenite from emerald mining tailings, but it produces low-content molybdenite concentrates, thereby resulting in a product with low economic value. The molybdenite concentrate produced in Campo Formoso contains 83.5% MoS2 (50% Mo), which is less than the international standard product that contains > 90% MoS2 (54% Mo).This low value is due to the presence of talc, which is a naturally hydrophobic mineral that is also floated with the molybdenite during the flotation process. Separability studies were done in a Partridge-Smith cell using dextrin and quebracho as a molybdenite depressor in solution at different concentrations and pH values to produce a technical-grade concentrate (i.e., 90% MoS2). These basic studies were accompanied by measurements of the zeta potential and contact angle, which was determined by the capillary rise method. The results of these floatability studies were used in bench flotation tests, which were performed in a Denver laboratory flotation cell. A molybdenite concentrate of 93.4% MoS2 was obtained when dextrin was used at a concentration of 100 g/t."

Jan 1, 2014

By Yuting Xue, Khaled Mohamed

Rib control is a critical component of underground coal mining operations, and comprehensive rib support design is essential for ensuring the safety of miners. However, there is currently no standardized approach for rib support design in the United States. To tackle this issue, the National Institute for Occupational Safety and Health (NIOSH) has developed a standalone application, Design of Rib Support (DORS), for rib stability assessment and rib support design under development load. The rib support design in DORS starts with the rib support density, which is determined from the unsupported rib factor of safety (RibFOS). The relationship between rib support density and unsupported RibFOS was obtained from the collected field data. The objective of this research is to validate DORS rib support design by analyzing the extent to which the proposed design capability was utilized to support the coal ribs using numerical simulations. A series of numerical models with varying overburden depth and rib height were built and the ribs were supported with the DORS-recommended rib support design. The bolt load was analyzed under different conditions and the results show that bolt load varies with the overburden depth, the location of the bolt and bolt length. The load in the bolt(s) increases with the reduction in unsupported RibFOS. The bolt(s) start to yield when the unsupported RibFOS falls below a threshold value of 0.8 and secondary support is necessary.

Jun 25, 2024

By B. W. Clyburn

"Coals from the Corporation's Lingan and No. 26 mines are classified as high-volatile A bituminous, with a very low inherent ash and containing all essential properties to provide metallurgical coal, with the exception of a high sulphur content. Only through proper cleaning at low specific gravity can the pyritic sulphur fraction of the coal be liberated to an extent where final sulphur meets coking standards. This first section of the paper deals with the operation of a Dutch State Mines heavy-media cyclone circuit to achieve an actual separation in the region of 1.30 specific gravity. A discussion of the parameter of control which must be applied along with an appraisal of the results obtained is included in the paper.Introduction Very high fluidity and unusually low ash are the outstanding characteristics of coal from Devco's mines in the Harbour seam of the Sydney coal field. However, sulphur content, mainly pyritic, ranges from 1 to 2.50Jo. To obtain the low and consistent sulphur level required in metallurgical coking coals, preparation at low specific gravity is required. Metallurgical coke is used as the fuel in an iron blast furnace. Its sulphur content largely determines the quantity of slag required to remove it and therefore influences furnace productivity and unit costs. Typical data on the various coal properties of importance in the production of coke are included in Table 1."

Jan 1, 1985

By E. E. Thurlow

The Geology Div. of Northern Pacific Railway Co. has adapted to its exploration efforts field procedures for geochemical sampling and semi-quantitative analysis which have been developed through research by the U. S. Geological Survey. The sampling is done in conjunction with reconnaissance geologic mapping on aerial photographs or during field checking of maps which have been prepared largely by photo-geologic interpretation. Stream sediment or soil samples are taken and readily soluble combined heavy metals are extracted with ammonium citrate followed by colorimetric titration with dithizone at the sample site. The effectiveness of this approach has been demonstrated both in positive and negative results, thereby assisting in the elimination of large areas of land from further immediate attention and in providing targets for followup exploration. An occa sional sample is submitted for spectrographic analysis as a check on the possible presence of other elements of economic significance. Field laboratory procedures are also being implemented for determining molybdenum, mercury, and other metals which may be valuable in themselves or useful as pathfinder elements in outlining mineralized areas. A lead-zinc occurrence in Oregon and a copper deposit in Montana were readily detected and outlined by stream sediment and soil sampling and a heavy metal anomaly that lacks obvious interpretation has been found in northwestern Montana. Detailed mapping and additional sampling are followed by other exploration techniques if warranted. Northern Pacific's program of reconnaissance mineral evaluation of company-owned lands and adjoining areas was outlined in a paper presented at the American Mining Congress meeting in Las Vegas four years ago. That report included policies and procedures with respect to mineral development on Railway Co. lands and may be referred to for information on these matters.1 In a sense this paper is a progress report on the use of geochemical prospecting in Northern Pacific's continuing mineral evaluation program. At this time there are positive results but as yet there is no real success story in the form of a minable ore body. GEOCHEMICAL PROSPECTING PROCEDURES The decision to use geochemistry as a reconnaissance exploration tool must be followed by research into sampling and analytical methods and into the interpretation of results. Various analytical methods were tried and evaluated by the Geology Div., including hot and cold acid extraction with and without fusion of the sample, panning and magnetic separation followed by dithizone testing of the magnetite, and cold aqueous extraction of metallic ions followed by indicator tests. Samples were first sent to the laboratory for analysis; later field laboratories were established and now virtually all testing is done at the sample site. It was concluded that, where applicable, the quickest dependable methods for reconnaissance use are the dithizone field tests for readily extractable heavy metals. This involves the simultaneous treatment of a sample, usually of soil or stream sediment, with two immiscible liquids; one a cold aqueous extractant (ammonium citrate) and the other a solution of dithizone in toluene or xylene.2 Dithizone is a colorimetric indicator that reacts quantitatively with copper, lead, zinc, and several other metals to form colored complexes in organic solvents. Results may be expressed in ml of 0.001 pct dithizone that are required to reach the blue end point, or calibration of the test results can be carried out with standard solutions of zinc; conversion factors would then be expressed in ppm, zinc equivalent. Calibration of dithizone solution with zinc standards indicates that 1 ml of 0.001 pct dithizone is equivalent to approximately 0.25 ppm of zinc. An authoritative report on dithizone field tests authored by H. E. Hawkes in the June-July 1963 issue of Economic Geology* states "The appeal of these techniques has been the simplicity and speed of the determinations as compared with more tedious laboratory determinations of trace metals in weathered rocks, soils, or plant materials. This obvious appeal, however, has often been more than matched by misuse of the method in such a manner that the results are

Jan 1, 1964

By J. E. Hilliard

AS is well-known, the length per unit volume, Lv, of any lineal feature (such as grain edges in a polycrys-talline specimen) can be estimated from a count of the number of point intersections with a random plane of polish. Specifically:' Lv = 2NA in which NA is the expected number of intersections per unit area of the plane. As with all other measurements in quantitative metallography, the estimation of Lv is subject to a sampling error, the magnitude of which will depend on the number of observations. The analysis is usually performed by counting the number of intersections falling within a closed test figure of known area applied to a succession of different areas on the plane of polish. The test figure can be drawn directly on the screen of a projection microscope or else superimposed on the image of the structure by means of a reticule in a focusing eyepiece of an ordinary microscope. The use of micrographs is very inefficient unless, as in the present study, these are also needed for some other purpose. Consider an analysis in which n applications are made of a test figure within which the expected number of intersections is N. The expected total number of intersections counted is then given by NT = nN If (N) is the standard deviation in N for one application of the figure and if the latter is applied to non-overlapping regions so that the N's are independent in the statistical sense, then the standard deviation in Nt is given by: a(NT) = a(N)vn Eliminating n from Eqs. [2] and [3] we obtain: Thus, k will be a constant for a given specimen and test figure. From Eqs. [ I] and [4] we find As is to be expected, therefore, the fractional error in the estimation of Lv is inversely proportional to the square root of the total number of intersections counted. It is very desirable to have at least a rough estimate of k before starting an analysis, since it will then be possible to decide in advance by use of Eq. [5J how many counts should be made in order to achieve the desired accuracy in Lv. Without such information the investigator, unless lucky, will either do needless (and tedious) work or else obtain a result with too large an error. It is very easy to calculate k for the limiting case of N — 0, i.e., when the test figure is vanishingly small. In this case there is a negligible probability of two or more intersections falling within the figure. There are therefore only two possible outcomes: a count of one with a probability of, say, p and a zero count with a probability of (I-p). Thus and Therefore: We might expect k to depend on N and therefore on the size of the test figure. Such a dependence was observed2 in the analogous case involving the estimation of grain-boundary area per unit volume by a count of the number of intersections of a test circle with traces of the boundaries on the plane of polish. In this case the theoretical value of k at the limit N — 0 was 1.41. It was also predicted that k would display an initial sharp decrease with increasing N. Experimentally, k was observed to decrease to a value of 0.62 at N - 6 and thereafter remain approximately constant. Furthermore, the value of k for N > 6 was found to be insensitive to the form of the structure provided no pronounced anisotropy or grain-size variation was present. In order to determinewhether k would exhibit a similar dependence on N in the estimation of grain-

Jan 1, 1967

By Arthur Hockman, Howard W. Jaffe, Howard F. Carl

THE problem of establishing practical techniques for evaluating natural corundum ores arose from the desire to improve the existing purchase specifications for crystal corundum procured by the Federal Government for stockpiling. Such purchase specifications had been developed for a large number of products, including minerals and ores considered to be critical and strategic because of the country's dependence on foreign sources of supply. It will be apparent to anyone considering the subject of purchase specifications that the determination of whether or not a submittal meets the specifications is a prime consideration. Consequently satisfactory specifications for minerals should not only incorporate physical, chemical, and mineralogical requirements that are practical but should also establish the means by which both buyer and seller can be assured that materials meet stated requirements. The Munition Board's Material Purchase Specification for corundum, dated February 1954, covered two grades of crystal corundum intended primarily for abrasive use. It established limits of size and of composition: Grade A, at least 95 pct +3 mesh and 92 pct alumina (Al2O3); Grade B, at least 95 pct +7 mesh and 90 pct alumina. It also specified that "of each lot of material, not less than 95 percent shall have a hardness equal to or exceeding nine (9) on Moh's Scale." These specifications were unsatisfactory for several reasons. They did not allow the purchase of lower grades of corundum being used by industry. Neither did they specify exactly what a corundum ore actually was, except that it should contain so much alumina and have a certain scratch hardness. Nothing was said about the abrasive qualities of the material that was to be purchased primarily for abrasive use. During the period 1946 to 1948 supplies of the two grades defined in the then current specifications continued to decrease. Early in 1949 industry recommended that the finer sizes of corundum known as "crystal corundum concentrates," ¼-in. to 16-mesh material, be added to the stockpile specifications and that lower grades of crystal corundum also be included. In the spring of 1949 the National Bureau of Standards and the Bureau of Mines were asked by the Munitions Board to cooperate in developing more practical specifications for grading natural corundum ores. The National Bureau of Standards indicated that it would investigate physical test methods for determining abrasive characteristics of corundum, and the Eastern Experiment Station of the Bureau of Mines at College Park, Md., stated that it would consider the mineralogical aspects of the problem. Members of these two agencies, in consultation with representatives of the Office of Materials Resources of the Munitions Board and of the Federal Supply Service of the General Services Administration, and with the cooperation of industrial suppliers and consumers of abrasive corundum, developed, in a fairly short time, more practical specifications for six grades of corundum ores. The new specifications, issued January 1951, were for corundum, natural crystals, as by this time high-purity synthetic corundum was available and widely used. Nevertheless, for certain specific applications the natural product was preferred over the synthetic material. It has been extensively used in forming snagging wheels. It breaks into more or less cubical particles, whereas artificial corundum fractures into splintery shapes. These produce irregular cuts in glass, while the more symmetrical shape produces a more uniform ground surface. For this reason natural corundum has been preferred for certain types of lens grinding. Pure corundum is clear, colorless alpha alumina (A12O3), but it is rarely found in nature in this form. The most important gem varieties include the blue and yellow sapphires and the ruby. Minute impurities of titanic oxide impart a blue color,

Jan 1, 1955

By George W. Leney

The iron ore industry might justifiably claim to have originated the science of geophysical prospecting. Their use of geophysical methods was probably the first application of these scientific techniques to the search for economic materials although the exact origin is undoubtedly lost in the unrecorded observations of. some prospector experimenting with compasses or bits of lodestone. Since 1879 when Thalen published "On the Examination of Iron Ore Deposits by Magnetic Measurements" these methods have been a standard supplement to other prospecting techniques. However, many of the great ore bodies of the Precambrian iron ranges were not particularly suitable for geophysical work. The ores were oxidized and enriched through leaching of silica so that they were nearly non-magnetic and other physical properties which distinguished them from the wall rocks were largely destroyed. With the gradual exhaustion of these high grade direct shipping ores, and ores that could be beneficiated by simple means, the iron ore industry turned initially toward the importation of similar high grade foreign ores. At the same time, new processes for concentrating lower grade materials were being developed and older ones improved upon. These concentrates' when pelletized produced such important savings in blast furnace hot metal costs that they now command a premium in the iron ore market and have weakened the competitive position of the remaining high grade ores. The ore bodies now being sought are usually unoxidized and frequently they are also metamorphosed.

Jan 1, 1965

By John J. Sammarco, Janet S. Flynt

This report (Independent Functional Safety Assessment Guidance 7.0) is the last in a nine-part series of recommendations and guidance addressing the functional safety of processor-controlled mining equipment. It is part of a risk-based system safety process encompassing hardware, software, humans, and the operating environment for the equipment’s life cycle. Figure1 shows a safety framework containing these recommendations. The reports in this series address the various life cycle stages of inception, design, approval and certification, commissioning, operation, maintenance, and decommissioning. These recommendations were developed as a joint project between the National Institute for Occupational Safety and Health and the Mine Safety and Health Administration. They are intended for use by mining companies, original equipment manufacturers, and after-market suppliers to these mining companies. Users of these reports are expected to consider the set in total during the design cycle. • 1.0 Safety Introduction (Part 1).—This is an introductory report for the general mining industry. It provides basic system/software safety concepts, discusses the need for mining to address the functional safety of programmable electronics (PE), and includes the benefits of implementing a system/software safety program. • 2.1 System Safety (Part 2) and 2.2 Software Safety (Part 3).—These reports draw heavily from International Electrotechnical Commission (IEC) standard IEC 61508 [IEC 1998a,b,c,d,e,f,g]and other standards. The scope is “surface and underground safety-related mining systems employing embedded, networked, and nonnetworked programmable electronics.” System safety seeks to design safety into all phases of the entire system. Software is a subsystem; thus, software safety is a part of the system’s safety. • 3.0 Safety File (Part 4).—This report contains the documentation that demonstrates the level of safety built into the system and identifies limitations for the system’s use and operation. In essence, it is a “proof of safety” that the system and its operation meet the appropriate level of safety for the intended application. It starts from the beginning of the design, is maintained during the full life cycle of the system, and provides administrative support for the safety program of the full system. • 4.0 Safety Assessment (Part 5).—The independent assessment of the safety file is addressed. It establishes consistent methods to determine the completeness and suitability of safety evidence and justifications. This assessment could be conducted by an independent third party. • Safety Framework Guidance.—It is intended to supplement the safety framework reports with guidance providing users with additional information. The purpose is to assist users in applying the concepts presented. In other words, the safety framework is what needs to be done and the guidance is how it can be done. The guidance information reinforces the concepts, describes various methodologies that can be used, and gives examples and references. It also gives information on the benefits and drawbacks of various methodologies. The guidance reports are not intended to promote a single methodology or to be an exhaustive treatment of the subject material. They provide information and references so that the user can more intelligently choose and implement the appropriate methodologies given the user’s application and capabilities. The guidance reports comprise parts 6 through 9 of the series and are listed below: [< 5.1 System Safety Guidance (Part 6).—This guidance supplements 2.1 System Safety. < 5.2 Software Safety Guidance (Part 7).—This guidance supplements 2.2 Software Safety. < 6.0 Safety File Guidance (Part 8).—This guidance supplements 3.0 Safety File. < 7.0 Independent Functional Safety Assessment Guidance (Part 9).—This guidance supplements 4.0 Independent Functional Safety Assessment.] [ ]

Jan 4, 2006

By A. G. Metcalfe, A. Siede

The hardening of annealed copper during fatigue testing appears to be independent of the applied stress and to occur largely within the first 4000 cycles. Copper hardened by fatigue is more resistant to annealing than copper hardened by static means, and softens by a multistage process. Creep studies at 300°C show that copper hardened by fatigue also has greater resistance to creep than statically hardened copper. Some theoretical explanations are discussed. RECENT interest in the behavior of metals subjected to fatigue stems from observations of phenomena which could not be explained by current theories of the structure of metals. In 1953 Bullen et al.l showed that the hardening caused by fatigue at low stresses did not cause severe bending of the lattice as did hardening by static methods. This was later confirmed by Wood, &apos;and Broom and Ham.3 Clarebrough et al4 showed that as the stress level used in fatigue hardening was increased, bending of the lattice was introduced, eventually approaching that of metals deformed by conventional cold-working methods. The type of hardening introduced by fatigue has other interesting characteristics. Kemsley5 showed that copper fractured by fatigue at low stresses resisted annealing as compared to copper hardened an equal amount by static means. The work of Clarebrough etal, shows that the manner of energy release during annealing differs for fatigue-hardened and statically hardened metal in both magnitude and temperature range of release. In addition to hardening of the lattice, fatigue has been shown to cause softening in materials previously hardened by static methods,5 and also in age-hardened alloys.7 The superposition of a small fatigue stress on the static loading of high-temperature alloys has been shown to increase their resistance to creep.&apos; The present work was an attempt to determine the relationship between fatigue hardening, resistance to annealing, and the creep properties of pure copper. EXPERIMENTAL PROGRAM The program described here was designed to answer three questions: 1) Can stress-cycle combinations well below the failure values give significant increases in hardness? 2) Does hardening, under different stress-cycle combinations, give increase resistance to annealing? And if so— 3) Does superior resistance to annealing confer superior creep properties over statically hardened material? Three-quarter hard OFHC copper was annealed in vacuo at 550°C for 1 hr to produce a strain-free material having a fine grain size and uniform hardness. (The standard deviation of the mean hardness of the annealed samples was less than + 1.9 Dpn). Annealing was performed after machining the fatigue specimens so as to avoid work-hardened surfaces. Jigs were used to avoid distortion. The fatigue properties of the annealed copper were determined using an accurately aligned Baldwin-Sonntag SF1-U machine operating at 1800 cycles per min. This limited the minimum number of cycles which could be examined to about 4000 stress reversals. Since failure resulted in specimen damage, hardness increase with number of cycles was measured by testing specimens fatigued to different fractions of the failure values. The results are shown in Fig. 1, and are superimposed on the S-N curve of the copper in Fig. 2. The resistance to annealing of fatigue-hardened material was compared to that of material hardened by uniaxial tensile straining. Specimens hardened to equal values by both methods were progressively annealed for 1 hr at temperature in vacuo. After each annealing treatment the hardness was measured using a Leitz Durimet hardness tester with a 500-g load. At least five tests were made for each hardness determination. The standard deviation for individual specimens varied from 0.26 to 1.48 hardness points. The results are shown in Figs. 3 to 5. The discontinuous nature of the annealing process has been brought out by drawing idealized curves through the data points. This can be justified by the consistency of the results as well as by general agreement with results of Kemsley.5 The creep resistance of fatigue-hardened and

Jan 1, 1960

By Thomas C. Wilder

ALTHOUGH there is an abundance of thermodynamic information available in the literature today, no experimental data on the standard molar properties of formation of zinc oxide have ever been reported for the temperature range 419.5" to 908°C. This is the temperature range for which zinc metal is a liquid. so that the formation reaction is: Data have been reported for Reaction [I] where zinc is a solid or a gas. Maier. Parks, and nderson&apos; measured the electromotive force of the cell: in order to obtain the free energy of the reaction: over a limited temperature range. With the known properties of H2O, those of ZnO were calculated. These workers found that t;j of ZnO is -75,930 ± 150 cal per gfw and Mtf is-83,000 300 cal per gfw at 25°C. akolkin made similar measurements to obtain ilF7 and AH; of -75,870 and -82,540 cal per dw respectively at the same temperature. Truesdale and arirlg and Kitchener and 1gnatowicz4 measured the equilibrium of the reactions: respectively, in order to determine Ff of ZnO from the known values of CO, CO2, and H/X The thermodynamic properties of ZnO, presumably taken from the above investigations, are tabulated by Kubaschewski and vans&apos; and Wicks and Block" The latter reference also gives values for the range where zinc is a liquid. These numbers are probably acquired from the known extremities of the range plus Cp data of the components. In the present work the electromotive force of the cell: was measured in order to calculate the properties of ZnO from the well known values of NiO for the temperature range where zinc is a liquid. In Cell [B] the cathode reaction is: and at the anode: The overall reaction is: for which the free energy change is: where is the electromotive force in volts of cell [B] after correction for the Pt-Ta thermocouple. The standard molar free energy of formation of NiO is very well known as a function of temperature. Kiukkola and agner&apos;s data are used in most tabulations. The error is probably less than I00 cal per gfw, The latter authors were the first to use solid mixed oxide electrolytes in galvanic cells. Their work with NiO and other oxides was performed with cells similar to Cell [B] using a variety of reference electrodes of fixed oxygen potential. There may be a small composition range for which either NiO, ZnO or both exist. If nonstoichiometry does exist for these oxides, then the properties described herein are those for the metal saturated compositions. The simple cell design was the same as described elsewheres except that the molten metal electrode was about 125 g of pure zinc including a small (0.25 to 0.50 g) amount of ZnO. The mixture of pure nickel and NiO powders making up the reference electrode was contained in a tube of the ZrOl + CaO electrolyte which was immersed in the molten zinc. A platinum contact was made to the reference electrode; tantalum to the zinc. A refractory cement plug was necessary at the open end of the electrolyte tube in order to prevent zinc vapor from contaminating the reference electrode. Experimental details such as the vessel containing the cell, furnace, temperature control and measurement, and electromotive force determinations of the cell have all been outlined previously. The temperature was varied randomly in the range between the freezing and boiling points of zinc. After each temperature change, time was allowed for the drift at the new temperature to be less than +;"c whereupon a series of three cell readings was taken at least 10

Jan 1, 1970

By J. A. Jr. Carpenter

Expanding world economic prosperity, probable peaking of conventional petroleum production in the coming decades and environmental concerns require efforts to increase the efficiency of, and the development of alternatives to, petroleum-based fuels used in automotive transportation. In January 2002, the U.S. Department of Energy (USDOE) and the U.S. Council for Automotive Research (USCAR) of the "Big Three" traditionally-USA-based automakers, DaimlerChrysler, Ford and General Motors, formed the Freedom Cooperative Automotive Research (FreedomCAR) Partnership to fund high-risk, high-payoff research into advanced automotive technologies with the potential for eventually lowering the USA&apos;s nearly complete (95%) dependence on petroleum, 55% imported, for its transportation fuels. FreedomCAR replaced the Partnership for a New Generation of Vehicles (PNGV) which existed from 1993 to 2001. In January 2003, the energy-supply industry joined with the FreedomCAR Partnership forming the FreedomCAR and Fuel Initiative (FC&FI) to develop the technologies needed for the mass production of fuel-cell vehicles using hydrogen as fuel and the infrastructure needed to deliver the hydrogen to the consumer. The goal is to enable the automotive and energy-supply industries to make commercialization decisions in 2015 on large-scale introduction of such vehicles and on developing the infrastructure beginning in 2020. In addition to work on fuel cells and hydrogen infrastructure, the FC&FI conducts research on technologies with the potential for nearer-term energy efficiency and environmental benefits, such as new engine concepts, lightweight materials, and engine-electric hybrid propulsion systems. Lightweighting is seen as crucial for shorter and for longer term benefits because hybrid and fuel-cell powertrains are expected, at least in the early stages of adoption, to be heavier and more costly per unit of power than the standard internal-combustion, gasoline- or diesel-fueled powertrains they replace. Research and development (R&D) on casting of aluminum automotive components were major parts of the PNGV and R&D on casting of magnesium automotive components have been and are so under FreedomCAR. This presentation will review those efforts.

Jan 1, 2006

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer&apos;s opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By Marston G. Fleming

E. C. Peterson (Anaconda Copper Mining CO., Darwin, Calif.)—A study of this quite comprehensible and interesting paper by Dr. Fleming brings to mind several observations in the practical application of alkalinity and related factors to the actual practice of lead mineral flotation. Soda ash has long been a widely used and a very helpful alkaline conditioning agent in the flotation of galena from the usual run of lead-zinc ores. Soda ash is one of the common "standard" conditioning agents tried in any laboratory investigation of lead-zinc ores because it has so often proved helpful to galena flotation. However, the use of soda ash in the actual flotation of oxidized lead ores is certainly not widespread. In the flotation of certain lead-zinc ores from the Park City district of Utah, it was found in the usual cyanide-zinc sulphate-xanthate circuit that soda ash had an effect of producing a very condensed and flat froth regardless of the many frothers tried and that substitution of lime for soda ash to produce the same pH (8.0 to 8.4) improved the froth condition. However, the flotation of coarse particles of galena became so critical that some would pass through into the tailing, but caustic soda as a substitute for either soda ash or lime produced a very desirable froth condition in the same pH range and greatly improved the metallurgical outcome. Milling was carried out in typically "hard water" from watersheds of limestone and other calcareous rocks and the ore also apparently contributed magnesium and calcium salts to the pulp. Certain lead-zinc ores of Mexico have shown their greatest flotation response in circuits conditioned with sodium bicarbonate, and such actual mill use was also believed to be related to the water necessarily used in milling. On other lead-zinc ores the optimum has been obtained in actual milling treatment by use of caustic soda in both circuits of the operation. In flotation of oxidized lead ores in which sulphidi-zation is employed by addition of sodium sulphide, very high pH&apos;s exist in the flotation circuit. With the usual oxide lead ores, demanding 5 to 15 lb of sodium sulphide per ton of ore, pH&apos;s in the circuit (closed and open water-circuits) may range from 9.5 to 12, and for ores that contain much anglesite or a high percentage of iron oxide minerals, sodium sulphide consumption may reach as much as 30 lb per ton of ore and the resultant pH will be correspondingly higher. Yet in such flotation treatment employing either xanthate or oil (high-sulphur crude oil or diesel oils) as the collector, satisfactory to excellent grades of concentrate and lead recovery are obtained. Although there have been great studies and accomplishments in the investigations of flotation fundamentals and flotation theory, and although an investigation to determine the effect of all factors entering into the actual plant flotation of sulphide and oxide ores must become very complicated, as Mr. Fleming has mentioned in his caution regarding the drawing of general conclusions for other ores, it seems that it would be of greater interest if the techniques and investigations could, with time, approach the conditions of actual practice and lead the way to improved and more efficient flotation plant performance. Marston G. Fleming (author&apos;s reply)—The gulf between fundamental flotation research and plant operation is, perhaps, less wide than Mr. Peterson suggests. For example, the work described in my paper developed from an extensive investigation of a complex lead-vanadium ore from southwest Africa. This investigation started as a standard ore-testing problem but, at almost every stage, results were obtained which could not be interpreted in terms of previous experience. A program of fundamental research was therefore undertaken and was closely interlocked with the ore testing. This coordinated investigation resulted in a flotation process which has now been proved by two years of successful plant operation, and although the grade and mineralogical constitution of ore as well as smelter requirements have altered more than could have been anticipated, our knowledge of the fundamental character of the problem has made it possible to meet each change in conditions with much more confidence than would have been the case had we neglected this aspect of the investigation.

Jan 1, 1954

By R. R. Angel

Drilling rate is a parameter that should be considered in determining the volume requirements for air and gas drilling. The use of past methods which ignore the effects of the solids content upon the pressure and velocity of the annulus flow stream can result in undercalculation of the required volurne by as much as 50 per cent. A vertical-flow equation is presented for determining volume requirements. This equation includes the effect of the solids that are transported up the annulus in the flow stream by incorporating the drilling rate as one of the parameters. The effect of down-hole temperature on required circulation rates is also ana-lvzed. A simple approximate method of determining volume requirements is presented. This method is more accurate than the methods used in the past. Hole cleaning difficulties are analyzed for a recent air drilling job where past methods indicated that - excess air was being used. Sample curves of calculated boitom-hole pressures are presented for air and gas drilling in several hole sizes. INTRODUCTION In certain areas the use of air or natural gas as a circulating medium for drilling oil and gas wells is becoming a common practice. Large increases in penetration rate and bit life are achieved through the use of these media in preference to mud or water. Drilling rates as high as 90 ft/hr have been obtained in shales. The importance of maintaining adequate circulation rate is generally recognized; however, much disagreement exists among drilling operators as to what constitutes "adequate" circulation rate. In quarry drilling, where annular velocities can be accurately determined, an annular velocity of 3,000 ft/min of standard air is required for best results in rocks having approximately the same density as those commonly penetrated in drilling oil or gas wells&apos;. Although this standard air velocity has proven satisfactory for quarry drilling, some oil and gas well drilling operators believe that an equivalent annular velocity of more than 4,000 ft/min is required; others believe that as little as 2,000 ft/min . is sufficient. Much of this disagreement results from determining the required circulation rates with methods which fail to incorporate the drilled solids in an equation which is applicable to vertical flow. Hughes Tool Co. Bulletins No. 23&apos; and 23-A3 present data for determining circulation rates based on the Weymouth formula. These data do not include the drilling rate as a parameter and, therefore, neglect the effect of the solids being transported up the annulus. In spite of this apparent defect and the fact that the Weymouth formula is not applicable to vertical flow, the Hughes data have well served the drilling industry in many areas and are important and timely contributions to the science of air and gas drilling. The Hughes data purposely omit a correction for increasing down-hole temperature. At slow drilling rates this effectively compensates for the use of a formula which is not valid for vertical flow; however, volumes determined by the Hughes method are not sufficient to support rapid drilling rates at moderate and great depths. For example, Phillips Petroleum Co.&apos;s Cauthorn "D" No. l in the Vinegarone field for Val Verde County, Tex., was air drilled from 1,500 to 9,300 ft using a compressor delivering 1,400 cu ft/min. The 8 3/4-in hole was drilled with 5-in. drill pipe, and drilling rates as high as 90 ft/hr were obtained between 7,000 and 9,300 ft. No water or caving hole was encountered. At 7,728 ft it was necessary to wash-out 60 ft of cuttings to reach bottom after a trip to change bits. At 8,130 ft a twist-off occurred and the drill col-lars were stuck in drill cuttings. These difficulties indicate that the 1,400 cu ft/min of air was not sufficient to keep the hole clean. Hughes data indicates that 1,180 cu ft/min is sufficient to produce an annular ve-

Jan 1, 1958

By Robert D. Pehlke, Robert G. Blossey

The solubility of nitrogen in liquid iron alloys containing chromium and nickel has been measured in the temperature range 1550° to 1700°C at nitrogen pressures to 1 ah. The solubility surface has been determined for the iron corner of the liquid metallic ternary to 21 pct Cr and 11 pct Ni. The nitrogen solubility increases markedly with increasing chromium concentration. The effect of nickel and the bilinear effect of chromium and nickel are less pronounced. The solubility data are presented, and interaction coefficients are determined from them. THE solubility of nitrogen has been determined in liquid Fe-Cr-Ni binary and ternary alloys. The work in this laboratory has been directed toward a relatively limited composition range, that is to 20 pct Cr and 10 pct Ni. Humbert and Elliott1 studied the entire ternary; but their investigation was, understandably, limited in this region. The region considered here is of the greatest commercial interest. Knowledge of the solubility in these alloys is important for accurate control of the nitrogen levels in various stainless alloys. Also, the solution behavior is of theoretical interest because of its deviations from ideal dilute solution rules at moderate chromium levels. The nitrogen solubility in the iron binaries has been reviewed extensively by Pehlke and Elliott,2 particularly with respect to the dilute solution interaction parameters. More recently, there has been some interest in accurate descriptions of solute interactions in non-dilute solutions. Most of the investigators have adopted a power series fit to accommodate any non-linearity of the data. The Taylor series expansion is adopted here, using Scimar&apos;s notation3 for the interaction coefficients. An alternate method, instead of requiring higher-order expansions, is to restrict the range of the composition variables. A planar solubility surface may be fitted to the data and its range of validity determined. If the axes of this region are displaced from the usual origin—the solvent pure iron—a change in standard state is required for the solution reaction. The standard state for nitrogen in iron is commonly taken as the hypothetical 1 wt pct solution of nitrogen in pure iron. However, for an alloy of appreciable solute concentration, e.g., 13 pct Cr or 18 pct Cr-8 pct Ni, it would be convenient to use that base composition as the solvent and redefine the solution reaction in terms of that particular solvent. A brief derivation follows, defining the free energy of solution in an alloy solvent. Following the treatment by Darken and Gurry,4 for a nitrogen pressure of 1 atm: N (pure component) = N (1 wt pct in solvent) [l] AG=RT ln100/ wt pct N in solvent Taking this reaction for an alloy solvent less the same reaction for a pure iron solvent the relation between nitrogen in pure iron and an alloy is: N (1 wt pct in Fe) = N (1 wt pct in alloy) [2] AG = RT[ln(wt pct N in Fe) - ln(wt pct N in alloy)] Adding ½ N2(g) to each side and subtracting N (1 wt pct in Fe) = ½N2(g), the total free energy for solution of nitrogen in an alloy, referred to nitrogen at infinite dilution, is given by: ½N2 (g) = N (wt pct in alloy) L3] AG°3 = -RT ln(wt pct N in alloy )PN =1 atm Interaction parameters have their usual meaning when ?G°/RT is differentiated with respect to the proper composition coordinate. If desired, the dilute alloy treatment may be utilized with the new reference composition as the solvent. This has. indeed, been done by Small and pehlke5 in their analysis of nitrogen solution in liquid 18-8 alloys. The Sieverts&apos; technique was used for this investigation on an apparatus previously described.6 The materials used in this program were: Ferrovac-E iron, 99.95 pct, Crucible Steel Co.: chromium, 99.95 pct, Union Carbide Metals Co.; nickel, 99.9 pct, International Nickel Co. Recrystallized alumina crucibles were used with no evidence of crucible-melt reaction. RESULTS The locations of the ternary compositions investigated are shown in Fig. 1. The data of Turnock7 and small8 have been alluded to in other papers.5&apos;6 There were no departures from Sieverts&apos; law observed for

Jan 1, 1969

By Basil S. Fee

INTRODUCTION Probably the most important family of chemicals used in mineral processing today is a category of basic commodity chemicals loosely denoted as pH regulators. Typical chemicals which are referred to as pH regulators include lime, magnesium hydroxide, soda ash, caustic soda, ammonia, sulfuric acid, and hydrochloric acid. These chemicals are often used in very significant amounts in almost all of the major mineral processing operations such as flotation, hydrometallurgy, etc. (in dosages up to ten pounds per ton of feed ore treated). While cheaper in cost per unit weight of chemical than more specific chemicals such as collectors, frothers, extractents, etc., the overall cost to the mill operator is generally higher with pH regulators per ton of ore processed than with any other given processing chemical. For example, a rough rule of thumb in sulfide mineral flotation is that the cost of lime is double that of the collector(s) used. The symbol pH is used to designate hydrogen ion concentration. When acids, salts, and bases are dissolved in water, individual molecules are dissociated into their constituent radicals or ions. The strength of an acid or a base increases with the extent of such dissociation or ionization. An alkaline solution is one in which the number of Qydroxyl ions (OH ) exceeds the hydrogen ions (H ). In an acid solution, the hydrogen ions predominate. In either case, both ions are always present as water itself ionizes to a limited extent, that is: [ ] The pH scale is logarithmic and the pH value is the negative of the logarithm (base 10) of the molar concentration of hydrogen ions per liter of solution. For example, a pH of 5 means that the molar concentration of hydrogen ions per liter is 0.00001 (1 x 10-5). Likewise, a pH of 9 means that the molar concentration of hydrogen ions per liter is 1 x 10-9. Normally, the relationship between hydrogen ion and hydroxyl ion concentration is based on the relationship: Concentration of H ion x concentration of OH- ion = constant. Eq. (1) In dilute and/or moderately concentrated solutions that are normally used in mineral procygsing processes, the constant at 25°C is 10-14. On the pH scale, the value of pH equal to 7 represents the hydrogen ion concentration of a neutral solution. pH values lower than pH 7 indicate increasing acidity and higher values than 7 indicate alkalinity. Table 1 shows the nature of the pH scale at 25°C. Temperature affects the extent of ionization of dissolved acids, salts, bases, and water so that the hydrogen ion concentration (hence pH) of a solution is also affected by temperature. As a means of demonstrating this dependency, Table 2 shows the change of the exponent of the constant (base 10) in Eq. 1 and the pH value corresponding to neutrality as a function of temperature. It is also important, for example, that alkalinity or acidity expressed by pH not be confused with total alkalinity or total acidity. For example, total alkalinity is commonly determined by titration with a standard acid solution (usually HC1). pH is a measure of the hydroxyl ion concentration of an alkaline solution, whereas titration is a measure of an alkaline solution&apos;s acid neutralizing capacity. Thus, if one takes a series of various alkaline solutions prepared using different chemicals but all of exactly the same pH (and temperature) and then subsequently carries out a titration on each solution with a standard acid, it would be observed that the various alkaline solutions would neutralize entirely different amounts of acid. As an example, 0.08 grams of caustic (NaOH) , 6.32 grams of soda ash (Na2CO3), and 8.17 grams of ammonium hydroxide (NH4OH) all have a pH of 11.3 at 25°C. One liter of each of the above solutions neutralizes 0.073, 4.348 and 8.496 grams of HC1, respectively. Therefore, depending on the specific use of any given pH regulator, special tests need to be run by the mill operator to determine factors such as: the specific technical goals to be accomplished by

Jan 1, 1986

By J. M. Dean, R. L. Grayson

Introduction With the continued downward pressure on coal selling prices, mine management must make its operations as efficient as possible. The problem is the same one throughout the mineral industries, streamlining operations to maintain competitive min¬ing costs (Born, 1991). There has been a great deal of research into better using production manpower. But the story is different for operations support personnel. Traditionally, planning man¬power requirements are done annually, with modifications to the operating plan occurring generally quarterly. On the other hand, the demands on manpower are often dynamic, being responsive to changing productivity levels, mine conditions (including water influx, bad roof, poor floor) and unanticipated project requirements (responses to machine breakdowns or a new section setup requirement because of a drop in coal quality in another area). Such changes may impact the manpower structure significantly. At times they may cause serious labor relations problems concurrently. Realignments and the generation of instability in the workforce because of active job switching (bidding) are a result. Most responses by coal operators are reactive rather than proactive, a situation that could be mollified by more accu¬rate and detailed planning of predictable manpower require¬ments. A manpower planning tool is needed that takes into consideration such things as changing productivity levels and conditions as manhour requirements for different work areas are planned, rather than blindly allocated through traditional staffing modes. Mine Management Support System The Mine Management Support System (MMSS) is an integrated decision support system that was developed using Sun Microsystems hardware according to the contract be¬tween the US Bureau of Mines and West Virginia University. The software used in the development of MMSS included X¬ windows (Scheitler and Gettys, 1986) for the user interfaces, LASER (Raman, 1985; Hayhurst, 1990) for the expert sys¬tem and graphics components, and the C language for general programming requirements. Sun DOS was used to allow the use of IBM-compatible applications software on the Sun networks. Sun OS is the operating system. It can be viewed as an enhanced version of UNIX. One of the components of MMSS that was developed to address the many specific objectives outlined in Grayson et al. (1990) is the work scheduling system. MMSS accom¬plishes work scheduling by using a mathematical (goal) programming approach to minimize the amount of switching of workers from their regular jobs to perform other jobs (Grayson and Nutter, 1991). Manpower support program The main objective of MANSUPP is to plan a detailed manpower structure for a mine by incorporating a production schedule, condition-modified work standards and manage¬ment preferences to predict required manhours for various support activities, for example, belt moves. This is accom¬plished by gathering and processing planning information the same way a mine planner would. The program is menu¬driven with the selection of a particular menu item locating the user in the appropriate area within the spreadsheet to review/add information. All of the interactions between productivity, section conditions and management prefer¬ences and their effect on the amount of support work required must be considered. Information regarding the productivity of a particular section, the geometry of the section, work standards for individual activities and section conditions must all be gath¬ered for input into the program. In the initial stages of development, it was decided that the program be developed in an environment that mine planners were familiar with in order to accommodate the use of the program. Lotus 1-2-3 or a compatible (Quattro Pro) met this environmental need. Other reasons included ease of calculation due to built-in functions, powerful command language and macro capabili¬ties. Each of the major components of the program defined are summarized. Production schedule This area allows the user to input a production schedule for the entire mine over 12 periods. Space has been provided for five sections, advance or retreat, and two longwall sections. There are several calculations occurring in this area that do not apply to the determination of manpower, but serve as an aid in production planning. Section description The section description area of the spreadsheet is shown in Table 1. The first three items are calculated in the production schedule area of the worksheet and then transferred to this location. The remaining information is input by the user to be used in later calculations.

Jan 1, 1996

By Terkel Rosenqvist

J. Chipman (Massachusetts Institute of Technology, Cambridge, Mass.)—The fact that the experimental work has been applied to copper rather than iron and that the paper is presented to the Iron and Steel Division, I regard as rather significant. It shows the unity of metallurgy and the fallacy of trying to cut it up by metals. This result for the solubility of sulphur in molten copper correlates with Professor Schuhmann&apos;s finding that the published data on the other side of the copper-sulphur miscibility gap are also in error. I should like to ask the author to say a little bit more about the sulphur capacity of the slag. T. Rosenqvist (author&apos;s reply)-—I hope that Dr. Chipman will find the derivation of the expression for sulphur capacity more clearly explained in the printed version of the paper than in the oral discussion at the meeting. I feel that this quantity, which actually is the ratio of two activities, can be measured more easily than the individual activities. Even if the ratio CaO/CaS is chosen as the standard state, the expression can be used for any slag, even for slags completely free of lime, and it represents a way to put the desulphurizing power of all slag constituents into one bag. Some doubt has been expressed as to whether oxygen ions really exist in calcium oxide and in molten slags. From a thermodynamic view point that question is of minor importance. The term oxygen ion activity, or any activity for that matter, is defined rigorously by the equation: activity = exp p/RT, where p is the change in free energy connected with the transfer of one mol of ions from the standard state into the slag. Whatever happens to the ion in the slag is of no concern to the thermodynamicist. Regardless of whether the ion is "free" in the slag or not, or whether it is present in a very small amount, its activity can always be expressed, and for a thermo- dynamic calculation that is all we need. However, ionic activities will only be of some real value if they are simple functions of the slag composition, or can be measured easily. Concerning the real nature of the oxygen in the slag, my feeling is that the oxygen atom has a rather multiplex nature depending on how strongly it is tied by covalent forces or polarized by the other atoms or ions present. The oxides of iron, cobalt, and nickel differ from calcium oxide and blast furnace slags as to the amount of free electrons that can give rise to electronic conductivity. In slags we know that the conductivity is mostly ionic. The fact that reversible emf&apos;s can be obtained with oxygen electrodes in certain salt melts, indicates a significant amount of oxygen ions in these melts. But extended work, e.g. polaragraphic studies and measurements of transference number; are needed to obtain quantitative information about the real structure of the slags. D. E. Babcock (Republic Steel Corp., Youngstown, Ohio)—-With reference to the ion, it might be well to remember Dr. Moses Gomberg. All of his life he had no use for the ionization theory and he contributed greatly to the field of chemistry on the assumption there was no such thing as ions. I do not think we have to worry about whether the oxygen is ionic or not. I think one thing specifically should be brought to your attention and this I think is one of the important contributions of Dr. Rosenqvist. He pointed out what we know as oxygen potentials or what is described as oxygen potentials. I have used this concept for a long period of time and I want to state that if this concept is properly applied, it vitiates much of what we have in the literature, or makes our usual ideas regarding oxidation seem primitive. That one thing is more valuable than almost all the rest of the discussion as a fundamental basis on which to build a reasonable in-

Jan 1, 1952