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The methods employed for the addition of uranium to steels and cast iron in the Physical Metallurgy Division are outlined. It has been found that the segregation of? uranium is a major problem in aluminum and/or silicon killed steels and that the cooling rate is important in al-lowing the segregate to settle to one area, which can subsequently be discarded. The unsegregated areas were found to be extremely low in total oxygen, acid insoluble aluminum (AizOg) and, in some cases, slightly lower in nitrogen. Uranium recoveries, based on analyses of the unsegregated areas, were quite constant, at about 45 per cent, regardless of the type of uranium master alloy used for the addition. No segregation of carbon, manganese, silicon, sulphur, phosphorous, chromium or copper was observed in these steels. In commercial heats, the recoveries ranged from 7 to? 29 per cent for ladle additions to oxygen converted steel. When the steel was fully deoxidized prior to the uranium addition, recoveries ranged from 49 to 66 per cent, regardless of the steelmaking process.

Jan 1, 1964

By Shaohua Ju, Feng Jiang, Chuanhua Li, Jinhui Peng, Libo Zhang

A T-junction microchannel is proposed to overcome the existing problems in conventional solvent extraction of indium from zinc hydrometallurgical process. Experiments were performed to separate In3+, Fe3+ and Zn2+ from sulfate solutions using Di-(2-ethylhexyl)phosphoric acid (D2EHPA) as the extractant. The effect of parameters such as pH, microchannel size and contact time on the extraction behavior of In3+, Fe3+ and Zn2+ was investigated. At contact time of 2.52 s and initial solution pH of 0.48, the extraction efficiency of In3+, Fe3+ and Zn2+ was 99.13%, 8.31% and 2.45%, respectively and separation factor of In3+/Fe3+ and In3+/Zn2+ were almost 1229.54 and 4864.98, respectively. Then, the loaded organic phase was stripped in T-junction microchannel. At contact time of 33.6 s, 99.64% of stripping efficiency of In3+ and stripping equilibrium were reached, while concentration of Fe3+ and Zn2+ was lower than 0.01 g/L with good separation of In3+ from Fe3+ and Zn2+.

Jan 1, 2016

By S. E. Maddigan

MANY experimental and theoretical investigations have been made on the processes involved in annealing and recrystallization of metals. Most of these, however, have treated the time element as constant and heat-treatments used have been for relatively short times. Aside from the early work by Carpenter and Elam,1 the few that have investigated the effect of time on the processes have usually done so for fairly short time periods. During recent years the growing interest in long-time creep properties of metals and alloys under stress has made increasingly important a better knowledge of the recrystallization during prolonged heating. The recrystallization temperature of 70-30 brass given by Clark and White2 in 1932 was somewhat above 400°F. (204°C.) but under 500° (260°C.), whereas in our laboratory recrystallization has been observed in creep specimens at 204°C, the lowest temperature of Clark and White's range. Since the creep properties are quite different above and below the recrystallization temperature, erroneous conclusions can readily be drawn by the application of short-time recrystallization data to creep problems. It was thus considered of both practical and theoretical value to make long-time annealing studies near the minimum recrystallization temperature. In this way useful additions could be made to data on recrystallization temperatures and, since the reactions in this range are relatively slow, additional insight could be obtained into the factors involved in annealing. Several comprehensive reviews of recrystallization processes have been written,3-5 therefore only a brief review of the generally accepted facts will be given here.

Jan 1, 1940

By I. H. Gundiler

Sandia National Laboratories (SNL) is tasked to support The Department of Energy in the dismantlement and disposal of SNL designed weapon components. These components are sealed in a potting compound, and contain heavy metals, explosive, radioactive, and toxic materials. SNL developed a process to identify and remove the hazardous sub-components utilizing real-time radiography and abrasive water-jet cutting. The components were then crushed, granulated, screened, and separated into an aluminum and a precious-and-base-metals fraction using air-tables. Plastics were further cleaned for disposal as non-hazardous waste. New Mexico Bureau of Mines & Mineral Resources assisted SNL in investigation of size- reduction and separation technologies.

Jan 1, 1995

By J. Hanna, J. Jung, I. J. Anazia

Electrothermal production of elemental phosphorous (P4) generates substantial amounts of highly toxic phossy water and sludges. Because of their high phosphorous contents and lack of reliable processing technology, large tonnages of these hazardous wastes have accumulated from current and past operations. To date, the sludges are stored in ponds or above ground in special containment vessels at various locations around the country. Two samples of electric furnace phosphorous sludge (obtained from the Tennessee Valley Authority (TVA) in Muscle Shoals, AL and FMC Corporation in Pocatello, ID) were investigated using physical separation techniques to recover the elemental phosphorous and to render the residual material nonhazardous. The techniques tested include sizing, froth flotation and dispersed air oxidation. Flotation and sizing produced concentrates containing 69% - 90%P4, with recovery in the range of 60% - 87%. Dispersed air oxidation of the fines and tailings products was effective in the removal of an additional 10% - 15% of the P4 in the sludge.

Jan 1, 1993

By Ibezim J. Anazia

Electrothermal production of elemental phosphorous (P4) generates substantial amounts of highly toxic phossy water and sludges. Because of their high phosphorous content and lack of reliable processing technology, large tonnages of these hazardous wastes have accumulated from current and past operations. To date, the sludges are stored in ponds or above ground in special containment vessels at various locations around the country. Two samples of electric furnace phosphorous sludge obtained from the Tennessee Valley Authority (TVA) in Muscle Shoals, Alabama and FMC Corporation in Pocatello, Idaho were investigated using physical separation techniques to recover the elemental phosphorous in the sludge and to render the residual material non-hazardous. The techniques tested include sizing, froth flotation and dispersed air oxidation. Flotation and sizing produced concentrates containing 69 -90% P4 with a recovery in the range of 60 -87%. Dispersed air oxidation of the fines and tailing products was effective in removal of an additional 10 -15% of the P4 in the sludge.

Jan 1, 1991

By Takashi Sakaguchi, Akira Nakajima

Some attempts were made to recover and remove heavy metal elements, such as uranium, copper and chromium, from aqueous systems by using several phosphate compounds. The phosphorylated derivatives of polysaccharides, chitin phosphate, chitosan phosphate and cellulose phosphate, have high abilities to accumulate heavy metals, such as uranium, copper, manganese, cobalt, cadmium, zinc, nickel and chromium, compared with non-phosphorylated polysaccharides. In order to obtain basic information on the recovery and removal of uranium that may be present in industrial effluents, mine waste water and seawater, the adsorption of uranium by chitin phosphate and chitosan phosphate was investigated in details. The adsorption of uranium by chitin phosphate and chitosan phosphate was much greater than copper, cadmium, manganese, zinc, cobalt, nickel, magnesium and calcium. The adsorption of uranium was very rapid and was affected by pH of the solution, temperature. granule radius and co-existence of carbonate ion. The amounts of uranium adsorbed on the adsorbents increased linearly as the external uranium concentration increased. Uranium adsorbed on chitin phosphate easily desorbed with diluted sodium carbonate solution. On the other hand, uranyl and cobalt ions were separated from each other by using chitin phosphate. Along with these fundamental research, some attempts were made to recover uranium from seawater.

Jan 1, 1993

By Nakajima A

Some attempts were made to recover uranium from aqueous systems by using various biosubstances including microbial biomass. Of hundreds species of microorganisms tested, extremely high uranium absorbing ability was found in Pseudomonas stutzeri, Neurospora sitophila, Streptomyces albus and Streptomyces viridochromogenes. In parallel to this research, various kinds of biomass and commercially available crude drugs were screened to measure their abilities to adsorb uranium. Of these biosubstances tested, high uranium adsorbing ability was observed in chestnut skin, Comi fructus, Geranii herba and Chaenomelis fructus. From these results it may be considered that the uranium adsorbing ability of these biosubstances is due to the following; the uranyl ion is co-ordinated by the polyphenolic compounds such as tannin existing in the biosubstances. Along with these fundamental researches, we have developed a new adsorbent, the immobilised persimmon tannin (kakishibu tannin), for uranium and thorium. This adsorbent can accumulate 1.71 g of uranium per gram of the adsorbent (14 mEq/g), and the uranium adsorbing capacity of this adsorbent exceeds by several times that of commercially available chelating resins (2 - 3 mEq/g). We attempted to recover uranium from uranium refining waste water by using the immobilised persimmon tannin. As a result, the immobilised persimmon tannin can recover trace amounts of uranium from uranium refining waste water containing 10 - 15 ppb uranium with high efficiency. Almost all of the uranium adsorbed onto the adsorbent can be desorbed by washing with 0.05 - 0.1 M HCl or H2SO4. Thus, the immobilised persimmon tannin can be used repeatedly in the adsorption-desorption Cycle.

Jan 1, 1993

By T. K. Kim

Iron, cobalt, and nickel in waste ferric chloride solutions were separated from each other by solvent extraction. The iron was extracted from the cobalt and nickel by the use of methyl isobutyl ketone, and then the cobalt was separated from the nickel by the use of a high-molecular-weight quaternary amine dissolved in a solvent of high aromatic content.

Jan 1, 1980

By Batric Pesic

Solvent extraction of nickel and cobalt from the real leach solutions was studied by using a novel extraction system. The extraction system was a mixture of 2,6-bis-[5-n-nonylpyrazol-3¬y]] pyridine (BNPP) and dinonylnaphthalene sulfonic acid (DNSSA). The BNPP represents a novel molecule, which was designed and synthesized in our laboratory. Real solutions for the study were prepared by leaching of two different cobalt-nickel bearing concentrates, cobaltite [CoAsS] concentrate from Blackbird Mine, Idaho, and siegenite [(Ni,Co)3S4] concentrate from Missouri Lead-Belt. The major strength of the system was the selectivity of extraction toward nickel and cobalt from iron containing solutions. Other metals such as Mn, Mg, Ca, Al and As were not extracted. Due to the irreversibility of copper extraction, its removal from solution was required prior to the extraction of nickel and cobalt. Zinc was not as strongly extracted (i.e. its stripping was readily achieved) as was copper . Flowsheets for the recovery of nickel and cobalt were also developed.

Jan 1, 1997

By A. Onozaki

The Iijima Refinery was put into operation in 1972 and today has an annual production capacity of 156,000 tons of SHG electrolytic zinc. At the Iijima Refinery, the recoveries of gold and silver are 98.0% and 98.2%, respectively. Gold contained originally in the zinc concentrates reports in the Cu-Pb-Ag residue produced in the residue treatment process. Silver contained in the concentrates is distributed between the Cu-Pb~Ag residue and the copper cake produced in the purification process. The Cu-Pb-Ag residue, which contains gold and silver, is sent to the Oyachi flotatfon plant of the Kosaka Smelter; about 82% of the gold and about 85% of the silver are recovered in the copper flotation concentrate.. In this paper, the distribution of gold and silver in the Iijima. Refinery is discussed and operational results of the Oyachi flotation plant are presented.

Jan 1, 1990

By R. Q. Honaker

Dense medium processes are used to clean over 50% of the coal that is fed into preparation plant facilities worldwide. The dense medium is formed using ultrafine magnetite which is recovered after the cleaning step using drain-and-rinse screens and magnetic separators. Despite the high efficiency of the recovery system, magnetite losses of around 0.5 to 1.0 kg/ton are common. Considering the recent escalation of magnetite to around $300/ton, an alternative source of magnetic material from coal combustion bottom ash has been investigated. A relatively low cost recovery system has been developed. Dense-medium tests using the bottom ash magnetics revealed a high level of efficiency as indicated by probable error values around 0.03 over the range of medium densities investigated. The results from both the recovery and coal cleaning efforts will be discussed in detail.

Jan 1, 2009

By E. B. Viljoen

I. R. M. Chaston* As a gravity concentration man in the midst of so many eminent flotation experts I feel like Daniel. However, I would respectifully disagree with the opening remarks by Mr Viljoen. First, he suggested that gravity concentration of cassiterite was not effective in sizes below 30 microns. In a paper to the Institution of Mining and Metallurgy in 19621, I gave some figures for gravity concentration of deslimed material both for cassiterite and for wolfram which indicated that careful shaking table concentration of fine deslimed material could give recoveries of 75 to 90 percent in the size ranges above 13 micron. These were figures obtained from normal plant operation. It would be interesting to know if any tests were carried out on gravity concentration of the efficiently deslimed feed to this flotation plant to see what the recovery would be. Secondly, Mr Viljoen suggested that gravity concentration plants were expensive. He gives the cost of a flotation plant at R300 000 to treat a feed of 400 t.p.d. to the desliming section or 160 t.p.d. to the flotation cells. From his operating cost figures it would appear that the flotation cost is approximately R2 to R4 per ton of feed to flotation and the total operating cost about R2 per ton of feed. A table plant to treat this tonnage of deslimed feed would consist of about 30 tables and should cost considerably less to install and very much less to operate than the flotation plant. Third, Mr. Viljoen suggests that the concentrate grade from a gravity concentration operation would be unacceptably low. The paper quoted abovel showed that once a primary fine concentrate had been made, giving a suitably deslimed material, further gravity concentration could give concentrates of over 40 per cent Sn with high recoveries. Tailing from this dressing stage was naturally returned to the feed. To give these high grades, sulphides had been removed by flotation but this necessary flotation step is simple and cheap to operate and could be restricted to the final stage when the quantities involved are small. I would agree that flotation has a part to play in fine tin recovery but this role would only seem to become fully effective at sizes below 10 micron and would seem to require, as Mr. Viljoen has suggested, a mechanism for sizing which is effective down to, say, 1 micron. In the meantime it would be interesting to try tabling the flotation tailing to see if it was possible to recover the plus 20 micron cassiterite which is not recovered well by flotation. One point which has been made during discussion is the suggestion that cassiterite particles have a fully hydrated surface at normal pH. I wonder if this layer could materially affect the overall specific gravity of a very small cassiterite particle and, if so, whether gravity concentration of the very fine cassiterite would be improved if carried out in acid conditions, which would inhibit the formation of this hydrated surface. REFERENCES 1. CHASTON, I. R. M. 'Gravity concentration of fine cassiterite'. Trans. Inst. Min. and Met. Vol. 71, p. 4, 1961 and 1962, and discussion of this paper. Dr. R. P. King Mr Viljoen has drawn attention to the difficulties associated with the presence of very fine particles in the flotation pulp and he has indicated that, on the Union Tin plant, they must sacrifice 30 percent of the cassiterite in the minus 7 /Lm material which they separate and lose. He has also indicated that the rate of flotation of cassiterite remained high at these small sizes. The problem arises because the very small particles stabilize the froth making it impossible to handle. These stable froths cannot be easily broken and therefore cannot be pumped and processed. It seems that further fundamental research work on the properties of stabilized froths is urgently required. We have found another difficulty associated with the presence of very fine particles in phoscorite slurries; these particles can be easily floated but with very poor selectivity. We believe that it is possible that a purely physical mechanism such as entrainment is responsible for the collection of fine particles in the froth. We have also demonstrated that the presence of fine particles in the pulp adversely affects the rate of flotation of the larger particles. For esample, we have found an increase in recovery of + 100 um material from 2 percent to 17 percent in a standard laboratory batch flotation test on phoscorite when the - 37 um was removed from the feed. Such interactions between particles are not universally recognised as an important aspect of the complex flotation mechanism. They should appear in any quantitative formulation of the flotation rate process.

Recovery Efficiency Study on Ilesha Placer Gold Ore by Flotation Using Locally-Sourced Frothers and Collectors

Sep 13, 2010

By George J. Tamaro

In 1967 the Port Authority of New York and New Jersey undertook the construction of the World Trade Center (WTC) diaphragm walls, installation of the lateral support system and the excavation of the site, commonly referred to as the "Bathtub". The work took two years to complete. In 2001 the City of New York undertook the re-excavation of the site after the terrorist attacks. The recovery work will take eight months to complete. This paper describes the initial work on the WTC "bathtub" in the late 1960s and the recent work during the recovery.

Jan 1, 2002

By Robert M. St. Louis, Janet M. Edgecombe

The Mercur Mine's autoclave is an alkaline pressure oxidation system, operating at 220°(428°F) and 3061 kPa (444 psig) (total). Up to 725 metric tons (800 st) of ore, grading 2.02 g per mt (0.059 oz/st) Au and containing 1 to 2% sulfide sulfur, are treated daily in this circuit. A mineralogical and chemical investigation was initiated to locate the cause of lower than anticipated gold recoveries. Greater than 70 wt. % of reduced sulfur is oxidized; therefore, the autoclave is operating at specification. The low recoveries were attributed to increased concentrations of organic carbon in refractory ore. Activated carbon concentration in the attached CIL circuit was increased from 15 to 40 g per 1, leach temperature held below 21°C (70°F), and carbon flow changed to counterflow. These measures resulted in gold recovery improving from 75 to 82%.

Jan 1, 1990

By G. Sedelnikova, D. Kim, N. Ibragimova

"The object of this study was three samples of sulfide old tailings, which were obtained after Cu- Zn ores processing. Fine metal dissemination in sulfides as well as fine mineral intergrowth determine the refractory nature of material and low metal extraction using simple methods: gravitation 1-6%, flotation 32-48%, cyanidation 10-37%, and cause the need to use sulfide bio-oxidation to liberate gold and to recover it by heap leaching. The gold is the main valuable metal in the tailings with content 1-1.5 g/t. Heap bio-oxidation with mesophilic and moderately thermophilic bacteria was used for pre-treatment of tailings before gold heap leaching from bio-oxidation solid residues. The tailings have a low permeability < 0.3-1.2 m/day, which is caused by migration of fine particles, decreases recovery rates, and is poorly suited to heap bacterial leaching. Therefore, preliminary gold sulfide tailings with permeability < 0.3 m/day were pelletized using copper containing flotation tailings of converter slags and calcium oxide. The most gold extraction was obtained for pelletized tailings. The base metals (copper and zinc) were precipitated from bacterial solutions. Experiments have demonstrated effective bio-oxidation of all third gold sulfide tailings samples and enough high gold extraction 60-82%.INTRODUCTIONIn Russia, as everywhere, there is a trend toward deterioration in the quality of mineral resources, including non-ferrous and precious metals, manifested in lower grade and higher complexity of ores. The proportion of refractory ores is increasing, characterized by fine dissemination of metals in sulfides, intergrowth, and poor liberation of minerals during concentration, resulting in loss of base and precious metals to flotation tailings.Over the operation period, huge reserves (hundreds of million tons) of ore processing waste have been accumulated at mining and concentration operations. Because of the imperfections of the previously used ore concentration technologies, the content of precious and base metals in flotation tailings is comparable with the grades of the ores being currently processed. Therefore, old tailings can be viewed as an additional reserve of non-ferrous and precious metals. The works (DEN, 1997; Lyman, Johnson, & Guerney, 2000) demonstrate that, tailings containing sulfides (sulfur, copper, zinc, and other metals) are a potential environmental hazard since they produce acid drainage water containing toxic components. Processing those tailings would require designing an effective technology for recovering useful components and protecting the environment."

Jan 1, 2016

By F Cesnik

The past few years has seen a concerted effort to improve the recovery of gold along with its co-product copper from the æhigh-gradeÆ Ridgeway concentrator at Newcrest Mining LtdÆs Cadia Valley Operations. A number of initiatives were identified and implemented, culminating in a sustained improvement in recovery of approximately three to four per cent for both metals. The key projects that contributed to this improvement included the installation of a 355 kW Metso stirred media detritor (SMD) to create a regrind circuit, installation of froth crowders on the scavenger circuit, installation of a Falcon SB2500SE gravity concentrator as part of the regrind circuit to improve gold recovered to bullion on-site and the optimisation of the grinding circuit to reduce the 80 per cent passing size (P80) of the primary grind from 160 to 121 ¦m for the same throughput via process control, cyclone and ball charge optimisation. This paper will focus on the supporting test work, where applicable, that was used as the design basis for the project, but more importantly the plant data before and after the changes to clearly identify the improvements to plant recovery.

Jan 1, 2007

By H. M. Breza

Recovery of phosphate concentrate at Texasgulf’s Aurora Mine has received increased attention as existing mine and plant production capacity limits are reached. Improved mining techniques and mineral processing of waste streams offer the potential to increase phosphate production without major expansion. Innovative processes are being identified to maximize pit mining recovery and to minimize waste stream losses. This results in more effective use of the phosphate. Texasgulf Inc., is a wholly owned subsidiary of Elf Aquitaine. The Aurora phosphate rock mine and refining facilities are located in North Carolina on the south shore of the Pamlico Sound. Phosphate ore is recovered by open- cast mining techniques. Draglines and other earthmoving equipment are used to uncover and mine the phosphate ore. The mined ore then is slurried and pumped about 8 km (5 miles) to the processing plant.

Jan 1, 1994

By H. M. Breza

Recovery of phosphate concentrate has received increased attention as the existing mine and plant production capacity limitations of the Texasgulf Inc. Aurora, North Carolina Mine are reached. Improved mining techniques and mineral processing of waste streams offer a potential to increase overall phosphate production without major expansion of the Mining facility. Innovative processes have been or are being identified to maximize pit mining recovery and to minimize losses in waste streams, resulting in more effective use of the valuable phosphate natural resource. This paper will describe the various continuous recovery improvement programs being implemented to achieve this goal.

Jan 1, 1994