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By Y. Jeannin, C. Mannerskantz, F. D. Richardson

Measurements have been made between 1040o and 1300°C of the equilibrium between chromium metal, chromic oxide, hydrogen, and water. The results are in closer agreement with those deduced from recent thermal data than with previous equilibrium values. They lead to the equation ?G°1300°-1600°k = -266,700 + 59.95T ± 300 cal for the formzation of a mole of Cr2O3 from chromium and oxygen. This same equilibrium has been studied with the chromium dissolved in solid iron, and in this way chromium activities in Fe + Cr alloys have been established. The measuremenis show that there are substantial positive deviations from ideality in the a, and b phases, and support the results obtained by Heymer and Kubaschewski against those of other investigators. There is considerable disagreement between existing data,1"4 which is referred to in the Discussion, concerning the equilibrium Cr2O3 + 3H2 = 2Cr + 3H2O [l] There is also substantial disagreement between the results obtained by various investigators5-' who have measured the activities of chromium in its alloys with iron by the Knudsen effusion vapor pressure method. In view of the importance in metallurgical processes of chromium and its oxide and of Fe-Cr alloys, an attempt has been made to study equilib- rium [I] both with pure chormium metal and with chromium alloyed with iron, at temperatures between 1040o and 1300oC. From the work with the alloys it is possible to derive the chromium activities from the relationship

Jan 1, 1963

By J. H. Swisher

ALUMINUM deoxidation equilibrium in liquid iron has been the subject of many investigations. Sawamura and Sano1 have written a critical survey of the literature on this subject and consider the data of Gokcen and chipman2 and of d'Entremont, Guernsey, and Chip-man3 to be most reliable. The present study was undertaken to check these data and extend the range of the experiments to lower temperature. The experiments consisted of equilibrating 5-lb Fe-Al melts with recrystallized alumina crucibles at 1580o + 10oC, and analyzing the solidified ingots for aluminum and oxygen. The electrolytic iron stock contained 100 ppm C, 20 ppm 0, and trace amounts of other elements.* The stock was induction-melted in an atmosphere of 5 pet H2 in argon to attain equilibrium between the melt and the crucible. The temperature was measured with a W, 3 pet Re/W, 26 pet Re thermocouple, and the equilibration time was about 4 hr. The results are plotted in Fig. 1 along with curves obtained from the data of d'Entremont et al.,3 for temperatures of 1740o and 1910°C. The present results can be seen to be consistent with the previously reported high-temperature data. The reaction under study is AI2O3 (s) = 2 Al (l) +3O [1] for which the deoxidation product is Kl =[% Al]2 [%O]3 [2] (Under the experimental conditions, Fe A12O4 is not expected to form.4) In Fig. 2, values of log Ki evaluated at 0.01 pet A1 are plotted against 1/T, using data obtained from all the investigations under consideration. The deoxidation product is plotted in Fig. 2 in lieu of the equilibrium constant (normally obtained by evaluating Ki as pet A1 — 0) because of the uncertainty involved in extrapolating to zero concentration. The difficulty in extrapolating the present data for 1580°C is apparent curvature in the log Ki vs pet Al plot. This curvature is perhaps due to the uncertainty in oxygen analysis for concentrations of 5 ppm and below. For the data of Gokeen and chipman,' the extrapolation is also difficult because of experimental scatter. The line drawn through the points plotted in Fig. 2 may be represented by the equation:

Jan 1, 1968

By I. Yochev, A. A. Metelev, A. I. Stepanenko, A. P. Purgin, M. G. Viduetsky, A. M. Panshin, I. F. Garifulin

"Pneumatic flotation cells of the KFM series are intended for the concentration of various ores and technogenic materials. The key operating principle of the flotation cells of the KFM series is the preparation of pulp for flotation in blowers, offering the advantage of an accelerated flotation process. This report summarizes the results of the implementation and operation of the KFM cells at Russian and foreign projects in 2015-2017. Their implementation and operation experience confirms the high efficiency of the KFM cells, especially in the head of lead and zinc, copper and zinc, copper flotation processes:head KFM-600 flotation cell recovers from lead-zinc ore (Rudmetal AD modular concentrator) up to 52% of lead in a single stage (without cleaner flotation) in a salable lead head concentrate (grade of up to 76% lead, 2.3% zinc). The flotation cell proved to be much more efficient than originally expected in the process description in the design phase,flotation of the copper-zinc ore from Novo-Shemur deposit (PJSC Svyatogor concentrator) in two KFM- 1400M cells recovers up to 65% of copper and 70% of zinc in a high-grade bulk copper-zinc concentrate containing 12-13% of copper and 25-28% of zinc. The installation of the two cells with a combined capacity of 26 m3 allowed to decommission 4 chambers of FPM-16 pneumomechanical flotation cells with a combined capacity of 64 m3 with a simultaneous increase in the overall ore concentration performancea single KFM-1400M cell (13 m3) recovers from the copper ore from Volkov deposit (PJSC Svyatogor concentrator) 53% of copper into a high-grade copper head concentrate (33.0% of copper). For comparison, the pneumomechanical cell FPM-16 (capacity 16 m3) has almost half as low copper recovery (29.6%) into the head concentrate at a lower grade (30.5% copper)."

Jan 1, 2018

By Sumit Kumar Hota, Asadi Srinivasa Reddy, Ashok Kumar, Ashish Agrawal, Surajit Sinha, Chandra Sekhar, Tanay Ray

In the current operational facility, due to the absence of pellet pot facility, a novel technique is developed wherein the pellets with different chemistries prepared in the laboratory are fired in an induration machine by placing them inside the Inconel baskets. In the present work, an attempt is made to optimize the pellet chemistry by developing 2 different flux pellets, viz., low basicity (B2: CaO/ SiO2 = 0.3) pellets and high basicity (B2 = 0.9) pellets, and further the effects of basicity on pellet properties, viz., cold compressive strength, porosity, swelling index, reducibility index, and softening and melting temperature, are studied. It has been found through the multiple operational trials that increasing the basicity of pellets from 0.3 to 0.9 results into improving the pellet strength, reducibility and softening and melting characteristics. Moreover, it is observed that the swelling index of pellets is reduced by adding 2% MgO to both low and high basicity pellets that could be attributed to the formation of high melting point slag in pellets. The operational trial has resulted into producing pellets that exhibit a combination of desired mechanical strength, improved swelling, and reducibility index and possess better softening–melting behavior among all the chemistries studied.

May 24, 2022

By M. J. Santofimia

Multiphase microstructures in steel have been intensively studied in the past years, but combining non-equilibrium phases still offers a great potential for further development of these steel grades. Thus, improved combinations of mechanical properties can be obtained with microstructures formed by bainite or martensite, in combination with retained austenite. In particular, microstructures on the basis of carbon-depleted martensite and retained austenite can be produced by the very promising production process named quenching and partitioning (Q&P). Originally, the Q&P process aimed to avoid the formation of bainite during the heat treatment. However, the process does provide the possibility for (carbide-free) bainite formation during the partitioning step, i.e. in the presence of martensite. This article evaluates this approach, considering thatthe formation of bainite from austenite is strongly influenced by the preceding formation of martensite. Although the accelerating effect of the presence of martensite during bainite formation has been observed, it is not yet fully understood, and experimental and theoretical studies are being performed in order to come to a more effective exploitation of these processes for the formation of multiphase microstructures.

Jan 1, 2013

The initial results from the magnetic sludge factorial experiment are summarized in Table 2. For ease in interpretation, the four separate variables are identified with the high and low level of each variable being represented by + and - signs, respectively. Several important relationships are evident from an inspection of the data in Table 2. For example, magnetic sludges were obtained only during the high temperature tests. Secondly, with the exception of the last three tests, there is an apparent correlation between settled sludge volume (or sludge settling rate) and pH. Data on residual dissolved Fee+ indicate that precipitation of iron is not complete at pH 8 and room temperature (Tests l, 3, 5, and 7). Although this should not have any confounding effect on the interpretation of the results, it is an important observation and an aspect of the overall conceptual process that should be considered carefully during any future developmental work. The data in Table 2 were employed more fully in the calculation of main effects and variable interactions for each of the three responses. For this purpose, a codified design and calculation matrix, shown in Table 3, was used. The derivation of this matrix is described by Davies.(9) The positive or negative signs for variable interactions are obtained simply by algebraic multiplication of the signs for each single variable involved in the interaction. Main effects and interactions are calculated by multiplying the response values (Table 2) by the corresponding effect elements (Table 3), summing the results, and dividing by 8. The resulting value for single variables is also sometimes designated as an average effect, since (for tyro levels of the variable only) it is simply the difference between the average response of all trials carried out at the first level of the variable (factor) and that of all trials at the second level. If the effect of one variable changes at different levels of another variable, the two variables are said to interact. The relative importance of the variable or variables in question is reflected by the absolute magnitude of the calculated main effect or interaction. First-order (two variable) interactions are usually easy to interpret in terms of the physical behavior of the system. Second- and higher-order interactions are often less readily understood, and are often of minor significance in any case. The main effects and interactions derived from the present study in the manner described above are listed in Table 4. It should be noted that

Jan 1, 1971

By Thomas L. Hurst, Joseph M. Hans

INTRODUCTION The former Shiprock uranium mill site is located in the community of Shiprock, New Mexico on a bench overlooking the San Juan River (Figure 1). Both Shiprock and the mill site are located on the Navajo Indian Reservation. The millsite was built on a 230 acre tract leased by Kerr-McGee Oil Industries Inc., and was operated from 1954 to 1963. The Vanadium Corporation of America and its successor, the Foote Mineral Company, operated the mill from 1963 to 1968. During its entire operation, 1.5 million tons of uranium ore was milled producing 3711 tons of U308 concentrate. The concentrate was sold to the former U.S. Atomic Energy Commission. During milling, the ore was crushed to less than 35 mesh and acid leached to remove the uranium. The [ ] uranium was removed from the leach solution by solvent extraction and the uranium-barren solution discharged into raffinate ponds for disposal by evaporation and seepage. The spent solids (tailings) were pulped with waste process fluids and water and discharged to form tailings piles. The piles (ponds or dams) were formed with earth starter dikes. The dike walls were then raised using the separated sand fraction of the tailings. The slime fraction of the tailings were discharged into the diked impoundment. Since only a small fraction (0.25 percent) of the ore was U308 by weight, most of the ore bulk ended up as wastes. The wastes contain nearly all of the uranium decay chain radionuclides present in the ore except uranium. The mill was shut down and partially dismantled before the lease expired in 1973, and the remainder of the mill and wastes reverted back to the Navajo Tribe which was the lessor. Portions of the millsite were then occupied by the Navajo Engineering and Construction Authority (NECA), a Tribal enterprise. In order to provide skilled personnel to operate earth moving equipment, a heavy equipment operator's training school was also established on the mill site and was administered and operated by NECA. The U.S. Public Health Service (PHS) and the Environmental Protection Agency (EPA) staff conducted radiation surveys at four inactive uranium millsites on the Navajo Reservation in April 1974. During the survey of the Shiprock millsite, it was noted that the training activities were degrading its radiological condition. Both the PHS and EPA staff recommended that the training activities be directed towards the decontamination of the millsite area and interim stabilization of the tailings piles. Preparation for the decontami-nation work commenced in April 1974. In order to facilitate the decontamination activities, the mill site and its contaminated environs were divided into major areas labeled A through F (Figure 2). Areas A, B, and C are defined as "perimeter" areas and contained surface contamination ranging from 5 to 30 cm deep. Area D defined the mill yard and contained contamination ranging from 0.6 to 5.5 m deep. Area E contained several raffinate ponds which were removed to line up the northwest boundary of the tailings pile and to move the contaminants away from the vicinity of the occupied on-site buildings. The contaminants in Area E ranged from 0.05 to 5.5 m in depth. Area F encompassed the ore storage yard. Contaminants in this area ranged from 0.18 to 1.2 m deep. Washes, however, had contamination exceeding 7.6 m deep.

Jan 1, 1981

By M. Sadegh Safarzadeh, Erich U. Petersen, Jan D. Miller

"An acid bake-leach process has been identified and applied to an enargite concentrate containing ~67% enargite, ~27% pyrite (FeS2), ~3% chalcopyrite (CuFeS2), ~6% sphalerite (ZnS) and <1% galena (PbS). Sulfuric acid baking of this concentrate resulted in the transformation of the concentrate to S8, As2O3, HFe(SO4)2•4H2O, PbSO4 and Cu0.47Fe0.53SO4•H2O, all of which (except elemental sulfur and PbSO4) dissolved during the water leaching step. More than 87% Cu and 64% As were leached when the concentrate was baked at 200°C over 7 hours with a sulfuric acid-toconcentrate weight ratio of 1.6. Less than 1% As was lost to the gas phase. Upon water leaching, copper and arsenic were released to the solution at a concentration ratio of about 3:1. While enargite is very reactive during the sulfuric acid baking reaction, other minerals present in the concentrate such as pyrite, galena and sphalerite may compete to consume sulfuric acid and be converted into their corresponding sulfates at different concentrations. The purpose of this paper is to present our examination of the effects of sulfuric acid dosage on the extractions of Cu, Fe, and As from an enargite concentrate. In this regard, the concentrate, baked concentrate, and the leach residues are characterized and a conceptual process flowsheet is described for the acid bake-leach process for the treatment of enargite concentrates.IntroductionThe occurrence of enargite (Cu3AsS4) in copper ores and concentrates and the associated high arsenic values in such resources have resulted in a considerable decrease in the demand for such concentrates because they cannot be economically processed using conventional technologies such as smelting and atmospheric leaching (Mihajlovic et al., 2011). The exploitation of such problematic deposits through competitive treatment strategies will be required in the near future in order to satisfy increasing pressure from the copper market. It is expected that the average arsenic grade of copper concentrates will increase from 0.17% in 2008 to 0.22% by 2020 (Mayhew et al., 2011).The acid bake-leach process was identified and selected for further investigation at the University of Utah as a possible process alternative for the treatment of enargite concentrates. In this case, adequate amounts of concentrated sulfuric acid are mixed with the concentrate and heated to the temperature range 150-300°C. Subsequently, the baked material is leached with water/weak acid and the acid bake-leach process has some advantages over conventional sulfuric acid leaching (Prater et al., 1970). During water leaching, metal sulfates are released in solution, after which they can be extracted through well-known hydrometallurgical techniques. While the majority of the work has been done on oxide minerals, there is a limited record of sulfuric acid baking of sulfides."

Jan 1, 2015

NEW MEMBERS. The following list comprises the names of those persons who became members during the month of November, 1913: Members. AGNEW, JOHN L., Gent Supt Copper Cliff, Ont., Canada. BAILEY, FRANK C, Mine Mgr 2003 IV. Snito Ave., Spokane, Wash. BATEMAN, ALAN M., Min. Geol.,Care L. C. Graton, Geological Museum, Cambridge,Mass. BERRYHILL, JAMES G.,Atty.-at-law and Mining Investor, Equitable Bldg., Des Moines, Ia. CARTLIDGE, OSCAR, Mgr. State Mine Rescue Station, 452 W. South Grand Ave.. Springfield. Ill.

Jan 12, 1913

By J. H. Nead, T. S. Washburn

The grades of commercial steel produced in large quantities can be divided into two general types from the standpoint of ingot structure— killed and rimmed. Killed steel covers a wide variety with carbon contents throughout the range of the steel portion of the iron-carbon diagram, and is the kind usually used in the production of alloy steels. Rimmed steel is normally associated with carbon ranges under 0.30 per cent, and is not yet used to a large extent in alloy practice, except for copper-bearing steels. The fundamental differences in the ingots made from these two types of steel can be seen most easily in Figs. 1 to 7. In the rimmed ingot shown in Figs. 5 to 7 there are two lines of blowholes approximately parallel to the sides and coming together at the top and at the bottom of the ingot. In the killed ingots shown in Figs. 1 to 4 these blowholes arc not present. Also, the rimmed ingot has no well defined pipe, as compared to the deep pipe on the killed ingots, and there are numerous primary blowholes in the lower portion of the rimmed ingot perpendicular to the surface and close to it, while on the killed ingot there are no blowholes of this type in the ingot shown in Fig. 1 and very small ones close to the surface in the lower part of the ingot shown in Figs. 2 to 4. With respect to the presence of pipe and the primary blowholes, the structure may vary considerably in both rimmed and killed ingots, therefore these cannot be considered as fundamental structural characteristics differentiating the two types of steel. The primary blowholes in rimmed ingots may be much less pronounced than those in the ingot shown in Figs. 5 to 7—as, for example, in "ingot iron," which is entirely free from them. An example of a cross section of an ingot-iron ingot is shown on page 16 of "Research and Methods of Analysis of Iron and Steel," published by the American Rolling Mill Co. (1920). A certain amount of pipe may also occur in rimmed ingots, though it is normally porous and not as well defined [or deep as in the killed ingot of Fig. 1. In killed ingots, primary blowholes may occur near the surface, as in Figs. 2 to 4, and

Jan 1, 1937

By Luz Marina Acharte L., Luis Quispealaya A., Enriquez Donaires A., Daniel Lovera D.

La presente investigación modela los impactos de los polvos de perforación en frentes mineros empleando una metodología integral: producción, ambiental, seguridad y salud ocupacional en la pequeña minería Cambalache ubicado en la Provincia de Angaraes, Departamento de Huancavelica a una altura de 3,250 m.s.n.m, que realiza actividades de explotación subterránea de minerales polimetálicos (Ag, Pb, Zn , Fe y Cu), con un análisis mineralógico de la roca: Magnetita ( tamaños hasta de 0,775 mm), Hematita (0,250 mm), Ilmenita (0,250 mm), Rutilo ( 0,175 mm), Leucoxeno ( 0,185 mm de longitud), Pirita ( 0,045 mm), Calcopirita ( 0,055 mm ) y Goethita.

Sep 1, 2017

The Inspiration ore body was discovered in 1904, but a 50 -ton mill which was installed to treat the disseminated copper ore promptly failed. The original Inspiration Company then gave way to another, backed by W. B. Thompson, which found 40 million tons of ore. Control then passed to Anaconda and Dr. Louis Ricketts took charge as consulting engineer. He abandoned the mill which had been partially built (to the horror of the stockholders, according to Ira Joralemon) and spent a year and a million dollars experimenting with flotation. The first large copper flotation plant was then built and recovered 50% more copper than had previous gravity mills. It, of course, was highly successful and became the forerunner of modern flotation plants. Inspiration is distinctive as the first fully integrated copper operation in the United States. Three open pits supply sulfide ore to the concentrator and mixed oxide and sulfide ore to a vat leaching and electrowinning operation with leached ore also being sent to the concentrator. In addition, dump leaching supplies pregnant solution for scrap iron precipitation. Inspiration has recently constructed an electric furnace smelter and also operates a continuous casting and rod plant. The new primary 54 in. by 74 in. crusher was completed in May of 1972 and replaced an older unit which was in the way of the open pit development. It is of an unusual circular silo design, which permits dumping 0 from two positions at 90 from each other. One dumping position is in line with the spider, while the other one is at right angles to it. A box girder splitter bar hung from the dump pocket slopes downward from the second dump point, across the spider cap, and rests on the pocket bottom. This splitter was not installed to protect the spider but was to aid in distributing wet ore being dumped from that position to both sides of the spider. Any large amount of wet ore delivered to one side of the spider would pack in the crusher cavity. The round front wall of the 18 ft (5.5 m) deep pocket is unlined except for light steel protection at the top, while the straight back wall is lined with vertically positioned railroad rails. Water for dust suppression is sprayed into the pocket through a cluster of 3/8 in. (9.5 mm) nozzles at the back of the pocket. About 5 gpm (19 lpm) of water are mixed with 90 psi (620 kPa) air upstream of the nozzles to produce a fog which is effective in dust control. The 54 in. Allis -Chalmers gyratory is placed mostly within one quadrant of the circular enclosure to permit placement of the apron feeder on the lowest level in the proper position to transfer crushed ore to the

Jan 1, 1978

By N. Mashaba, D. Botha

The aims of this study were to investigate the factors that influence women’s engagement in technical positions in the South African mining sector and to determine what could be done to promote their successful participation. A convergent parallel mixed-methods research design was used to ascertain the factors that facilitate, inhibit, and influence engagement. In the quantitative phase of the study, questionnaires were circulated to women employeees; and in the qualitative phase, semistructured interviews were conducted with employer representatives, most of whom were human resource personnel. Confirmatory factor analysis showed that the three-factor structure (vigour, dedication, and absorption) of the Utrecht Work Engagement Scale fit the sample data reasonably well. Although there is room for improvement, respondents demonstrated acceptable levels of engagement in their work. On the other hand, the qualitative findings showed that employee engagement is impacted by unfavourable working conditions, work–life balance, and the mining industry’s male-dominated work culture. The findings showed that employee engagement should be elevated to a core human resource function. To increase the participation of women in mining, human resource professionals are encouraged to collaborate with mine supervisors, managers, and employees to develop programmes that promote employees’ absorption in and dedication to their work.

Nov 6, 2023

By Janet A. Johnson, T. B. Borak, K. J. Schiager

INTRODUCTION In the USA regulatory agencies have adopted standards which limit the allowable concentration of ore dust in underground uranium mines. The American Conference of Governmental Industrial Hygienists has recommended a threshold limit value for silica which has been incorporated by reference in federal regulations (30 CFR 57, 1980). The Nuclear Regulatory Commission (NRC) imposes a limit for unprocessed uranium ore dust for reducing the exposure to radioactive materials from inhalation (10 CFR 20, 1980). We have been unable to trace the origin of the NRC standard. Recently the ICRP has published recommendations concerning limits for intakes of radionuclides by workers (ICRP, 1979). We have used the ICRP methodology to compute the dose equivalent commitment to lung and bone from inhalation of insoluble ore dust particles. These values were used to derive air concentrations which would provide an acceptable risk for induction of cancer from occupational exposure. Results indicate that the new ICRP recommendations are more restrictive than the present NRC standard for radioactivity. The maximum allowable dust load in a uranium mine will depend on both the concentration of silica (% Si02) and the amount of radioactivity (% U308). We have combined these standards assuming complete independence of biological hazard to establish boundary conditions for which either silica or radioactivity is the dominating factor. Based on these results we present suggestions for analysis of ore dust to insure compliance with regulatory agencies. SILICA STANDARDS Silicosis, fibrotic lung disease produced by crystalline silica, was one of the earliest forms of occupational disease to be recognized. The pulmonary lesions caused by silica, silicotic nodules, consist of concentrically arranged bundles of collagen fibers. Fusion of these nodules results in progressive massive fibrosis causing alveolar changes which decrease ventillation and blood flow in the lungs. The current Threshold Limit Value (TLV) for mineral dusts is based on the concentration of free silica. The standard is expressed in three forms: (1) Particle count standard: TLV (mppcf) = 300/(% Si02 + 10) (mppcf = million particles per cubic foot) (2) Respirable dust mass standard: TLV (mg/m3 ) = 10/(% respirable Si02 + 2) (3) Total dust standard: TLV (mg/m3 ) = 30/(% Si02 + 3) The most commonly used form is the respirable mass standard. These curent values for silica are based primarily on epidemilogical studies of Vermont granite workers and other occupationally exposed workers. In the Vermont study no cases of silicosis were seen in individuals employed after dust control measures were initiated and whose subsequent exposure averaged less than 5 mppcf. The free silica concentration in the airborne dust to which these workers were exposed averaged 25%. A concentration of 10 mppcf of granite dust is considered equivalent to 0.1 mg/m3 quartz. The current respirable mass standard allows a maximum silica concentration of 0.1 mg/m3. NIOSH (1974) has recommended a reduction in the current TLV for respirable silica to 0.05 mg/m3 Si02. As with radiation standards, the TLV for mineral dusts is an upper limit for time weighted average exposure. It is recommended that dust concentrations be maintained as far below the TLV as current practices permit. RADIATION STANDARDS The present concentration limit for airborne natural uranium in ore dust prior to chemical separation is 75 µg of uranium per m3 of air ( 10 CFR 20, 1980) . This corresponds to 1.85 Bq/m3 (50 pCi/m3 ) of natural uranium, U-238, U-235 and U-234 and is equivalent to 3.7 Bq/m3 (100 pCi/m3 ) of gross alpha activity with radioactive equilibrium between the long-lived alpha progeny through Ra-226. The origin or basis for this standard is not known, but it presumably includes both chemical toxicity and risk from somatic radiation injury. Recently the International Commission on Radiological Protection (ICRP, 1977) has published recommendations concerning the objectives and criteria for limiting radiation exposures. This was followed b y a revision of the limits for intake of radionuclides by workers (ICRP, 1979; ICRP, 1980). The previous compilation of limits for internal emitters was published by ICRP Committee II in 1959 (ICRP, 1959). The latest version includes new methodlogy outlined in ICRP Publication 26 (ICRP, 1977) as well as extensive metabolic information

Jan 1, 1981

By Robert M. Abrams

The passage of an ore pulp through the froth flotation process results in the production, after dewatering, of a finely divided concentrate. This intermediate product Trust be handled to the subsequent process facilities via belt or other types of conveyors, hoppers, bins, chutes and feeders, and the like, Flotation concentrate has some difficult handling properties such as abrasiveness, sluggishness when moist, or fluidity when dry, These characteristics require the judicious selection or design of handling, equipment and the careful engineering of transfer chutes. It is to the special problems which are encountered in this in-plant handling procedure that this paper is addressed. The methods which will be discussed have beer. developed primarily for the handling of the froth flotation concentrates of copper, leap, and taconite iron ores, but may be generalized, with appropriate modifications, to apply to other similar metallic concentrate&. The controlled feeding of concentrates frog: bins and hoppers is covered alone with some of the techniques used to transfer concentrates from ore belt conveyor to another. Methods are described for unloading belt conveyors into bins or storage tiles, The reclamation from storage piles is also treated.

Jan 1, 1976

By L. W. Beckstead

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

Jan 1, 1993

By L. A. Haas

The U.S. Department of the Interior, Bureau of Mines, is conducting laboratory-scale research to determine the leaching conditions required with sulfuric acid for extracting base metal from a matte with a copper-to-nicke1 ratio of 6 to 1. The matte (61.2 pct Cu, 9.7 pct Ni, 3.3 pct Fe, 21 pct S, and about 5 oz of precious metals per ton) was prepared from Duluth Complex (Minnesota) ore by smelting a bulk sulfide flotation concentrate. The matte was pre leached at atmospheric pressure to selectively extract 20 pct of the nickel, and the undissolved solids were referred to as the matte residue. Autoclave leaching of the matte residue extracted over 95 pct of the nickel, copper, and iron. The remaining leach residue in the autoclave contained over 100 oz of precious metals per ton. The leaching conditions required were matte grind of greater than 98 pct minus 325 mesh, pulp concentration of 50 g/1, total pressure of 100 psig with oxygen at 150° C or 400 psig with air at 180° C, 50 g/1 H2S04, 700 rpm agitator speed, and 6 hours of leaching time. The leaching time was decreased to 2 hours by increasing the agitator speed to 1,300 rpm. Sulfuric acid concentration greater than 50 g/1 did not appreciably enhance the base metal extraction, but with less acid, residues with higher iron concentrations were obtained. Supersaturation of the liquor was prevented at 150° C by using a 1ixiviant that contained 30 g/l Ni and 20 g/1 Cu.

Jan 1, 1979

By Joseph A. Mann

"Soil mixing is a proven solution for the construction of ground improvement adjacent to sensitive structures due to the use of computer controlled and monitored equipment. The precision to which geometry is controlled can often be the deciding factor for selection of soil mixing versus other means of ground improvement. Protection of existing structures and utilities requires careful planning, investigation, and may demand supplemental control measures. During planning, it is imperative that the process be fully developed by the project team to ensure proper control measures are budgeted; otherwise, the success and safety of a project can be jeopardized.This paper discusses case histories of soil mixing performed in close proximity to large, sensitive, and critical utilities, including foundation support for an existing earthen dam impounding a water reservoir in upstate NY, a buffer zone for a tunnel drop shaft at an operating wastewater treatment plant in Washington DC, and a containment system to protect a major water line along an interstate widening project in Baltimore, MD. The function of soil mixing as ground improvement varied; however, the selection and application of the technique versus other means of ground improvement was specific due to the controlled manner of construction with soil mixing. Through the three case studies, small diameter (3- ft [0.9 m]), large diameter (8-ft [2.4 m]), and mass mixing applications will be discussed. Furthermore, the case studies offer insight into how the geotechnical community in the Mid-Atlantic utilizes soil mixing technology for widely varying applications. Additional topics of discussion will be the design and construction of the ground improvement systems, procedures and execution of the work to protect existing structures, and the QA/QC methods, procedures, and results from the projects.INTRODUCTIONWet soil mixing, also known as the Deep Mixing Method, is a ground improvement technique that improves the characteristics of weak soils by mechanically mixing them with cementitious binder slurry. To construct columns, a powerful drill advances a mixing tool with radial mixing paddles located near the bottom of the drill string. The binder slurry is pumped through the drill string to the tool as it advances and additional soil mixing is achieved as the tool is withdrawn.Wet soil mixing is applicable in a wide range of soils. Soil mixing has been performed at depths more than 100 lf (30 m). Soils vary widely in their ability to be mixed, depending on the soil type, strength, water content, plasticity, stratigraphy, and texture. Almost any soil type, including organics, can be treated with certain soil mixing technologies, although they may require high binder contents and/or pretreatment to achieve the required results. In cases when the target soil is very soft, including some very weak clays and peats, 100% of the soil may require treatment. This complete coverage of soil mixing is referred to as mass mixing."

Jan 1, 2015

By P. Schwarzkopf, J. H. Brophy, J. Wulff

A constitutional diagram has been proposed for the tantalum-rhenium alloy system. Rhenium dissolved in tantalum up to 48 wt pct, and the maximum solubility of tantalum in rhenium is 5 wt pct. Intermediate x and s phases were found. X-ray diffraction, metallography and melting points were employed. SEVERAL investigators have described the phases present in the tantalum-rhenium alloy system, but no detailed account of the phase diagram prior to the present investigation appears to be available. Greenfield and Beck&apos; reported the solubility of 48 to 54 at. pct Re in tantalum, a s phase at 41 at. pct Ta, and x phase isomorphous with a Mn from 37 at. pct Ta to less than 25 at. pct. They employed as-arc melted specimens and heat treatments at 1200 °C lasting 72 hr.2 Duwez3 reported the existence of a a phase at the composition Re3Ta2. Niemiec and Trzebiatowski4 observed a lattice parameter a = 9.69A, for the x phase at a composition Ta7Re22, but no s phase. Savitiski and Tylkina5 have shown a hardness maximum at the composition 40 at, pct Ta and a somewhat lower plateau from 35 to 10 at. pct. Knapton6 indicated results similar to those of Greenfield and Beck.&apos; The present investigation describes the phase relationships in the tantalum-rhenium alloy system determined by melting point measurements, X-ray diffraction and fluorescence, and metallography. EXPERIMENTAL PROCEDURE Commercial 99.7 pct Ta powder from Fansteel Metallurgical Corp. and 99.9 pct Re powder from Chase Brass and Copper Co. were used for alloy preparation. The powders were weighed to desired nominal compositions and then pressed without binder into compacts weighing 5 or 10 g. The compacts were arc-melted six times on alternate sides on a water-cooled copper hearth using a nonconsumable tungsten electrode in an atmosphere of titanium gettered helium, at a pressure of 500 mm Hg. In most cases this procedure yielded alloy buttons of satisfactory homogeneity. A series of alloy specimens were analyzed wet chemically for use as standards for X-ray fluorescence. All specimens were analyzed by fluorescence and the results showed that the compositions differed by less than 3 pct Re from the original as-weighed nominal composition.* On the basis of self-

Jan 1, 1961

By S. Y. Tshipeng, A. T. Kaniki

"RÉSUMÉ La lixiviation acide d’un minerai oxydé cupro-cobaltifère en milieu non réducteur et en milieu réducteur en utilisant le Fe2+ et le métabisulfite de sodium (SMBS) a été étudiée et les résultats ont été comparés. Les tests de lixiviation ont été effectués à température ambiante, dans une pulpe à 30 % solide, avec une vitesse d’agitation de 600 tr/min, un potentiel redox de 350 mV pour les tests en milieu réducteur, et un pH de 1,8. Les tests de lixiviation avec usage du SMBS comme réducteur ont donné la meilleure cinétique de lixiviation du cuivre. Pour le cobalt, c’est l’usage du Fe2+ qui a donné la meilleure cinétique de lixiviation. L’augmentation de la dose du fer dans la solution au-delà de 5 g/L a influencé négativement les rendements de lixiviation du cuivre et du cobalt. ABSTRACT Acid leaching of a copper-cobaltiferous oxide ore in a non-reducing environment and in a reducing environment using Fe2+ and sodium metabisulfite (SMBS) was studied and the results compa¬red. The leaching tests were performed at room temperature, in a 30% solids pulp, with a shaking speed of 600 rpm, a redox potential of 350 mV for the tests performed in the reducing environment, and a pH of 1.8. The leaching tests using the SMBS as a reducing agent produced the best copper leaching kinetics. As for cobalt, the best leaching kinetics was achieved using Fe2+. An increase in the amount of iron in the solution above 5 g/L negatively influenced copper and cobalt leaching yields. INTRODUCTION La République Démocratique du Congo (RDC) est un pays reconnu dans le monde pour son sous-sol qui regorge d’énormes quantités de cuivre et de cobalt (USGS, 2014; Palisade Research, 2016). La technique qui semble être la plus recommandée pour extraire à la fois le cuivre et le cobalt des minerais oxydés cupro-cobaltifères est la lixivia-tion acide en milieu réducteur. Cette technique est appliquée par la majorité des sociétés minières qui exploitent ces minerais dans l’ex-province du Katanga en RDC, à savoir: Tenke Fungurume Mining, Mutanda Mining, Kamoto Cop¬per Company, Sicomines, Gécamines, Boss Mining, Chemaf, Ruashi Mining et Somika. L’usage d’un agent réducteur est indispensable pour l’obtention d’un bon rendement de lixiviation du cobalt. En effet, plusieurs chercheurs s’accordent sur le fait que le cobalt trivalent de l’hétérogénite (CoO·2Co2O3·6H2O) est insoluble à l’action de l’acide seul; il ne devient soluble qu’après réduction en cobalt bivalent (Mwema, Mpoyo, & Kafumbila, 2002; Mulaba-Bafubiandi, Ndalamo, & Mamba, 2007; Ferron, 2008; Seo, Choi, Kim, & Tran, 2013; Stuurnam, Ndlovu, & Sibanda, 2014)."

Jan 1, 2018