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George Edward WebbeR, a member of the Institute since 1906, died in San Francisco, May 29,1922. Born in Dixmont, Me., in 1852, he went to California in 1857 where he lived until 1883 and received a thorough training in precious metal mining in the employ of his father and J. B. Haggin. In 1883, he went to the Black Hills, South Dakota, as superintendent of the Father DeSmit mine; later the Deadwood Terra also came under his management. These mines are now part of the Homestake Mining Co. In 1889, the Rothchilds asked Hamilton Smith to recommend an engineer to take the place of the late Henning Jennings, who had resigned from the El Calloo Mining Co., of El Calloo, Venezuela, in order that he might proceed to South Africa. Mr. Webber accepted the position, remaining there until 1893 when he returned to the United States on his way to Johannesburg, South Africa, where he had been appointed manager of the Crown Reef Mining Co., Ltd. Early in 1896, he became general manager of the Rand Mines, Ltd., and served in this capacity until 1911. During the latter part of this period, Mr. Webber was active in the greater amalgamations of H. Eckstein & Co., and it was largely on his advice that H. Eckstein & Co. carried out the negotiations preceding the formation of the Crown Mines, Ltd. On his retirement from South Africa, he returned to live in Ross Valley, near San Francisco, Cal. The Rand Mines, Ltd., of which Mr. Webber was general manager, was the largest company controlled by Wernher Beit, the leading London house interested in South Africa mining, whose local agents in Johannesburg were H. Eckstein & Co. This firm is now known as the Central Mining & Investment Corpn., and after the departure of Henning Jennings, Mr. Webber was the firm's most valued and confidential advisor and was very highly thought of in mining circles in South Africa. He was a ma of most upright and straightforward high character, plain spoken and sincere; and his recognized high character was one more qualification which aided his great success.

Jan 1, 1923

By D. A. Larson

This is the second in a series of Bureau of Mines reports which de-scribe experiments designed to devise an efficient, economic mechanical fragmentation technique for hard rock. The previous study described the development of the ripper cutting technique to successfully cut hard rock with 3-to 6-in-wide drag cutters. This study was conducted to test the ripper cutting technique at full scale using 9-in-wide drag cutters. The experiments were conducted with a special test device that cut shallow, 6-by 6-ft openings in 9,200-lbf/in2 compressive-strength concrete blocks. The tests showed that ripper cutting can excavate large-scale openings of the type required by the mining industry at a production rate that outperforms the conventional drill-blast technique. Moreover, the system can produce openings of various sizes and shapes, is simple to oper-ate and maintain, and produces very little dust or noise during operation. The system appears to meet all of the criteria for a successful hard-rock mining machine.

Jan 1, 1987

By M S. Tuohy

Since the independent review of fatal accidents in the mining industry carried out by Dr Sean Brady (the ‘Brady Review’), High Reliability Organisations (HROs) and their operating principles have been discussed across the Queensland mining sector. HROs are organisations said to consistently avoid serious failures, reliably delivering their organisational objectives despite operating in environments characterised by high levels of inherent risk and/or operational complexity. HRO concepts have previously found application in other similarly hazardous industries to mining. However, as Hopkins (2009) pointed out, HROs ‘are very elusive creatures that inhabit the realm of theory more than the real world’. The Queensland Resources Council (QRC) with support from the Commissioner for Resources Safety and Health initiated a review of how HRO principles can be translated to the mining industry. The review was conducted by Noetic, an Australian management consulting company that specialises in hazardous industries, including mining and metals, and has prior experience applying HRO theory.

May 5, 2022

By K. Zabrusky

Estimation of high grade epithermal vein deposits can present challenges in correctly representing the continuity of grade within an undulating structure. One solution to this problem is to use dynamic anisotropy to follow local variations in the strike and dip of veins. Dynamic anisotropy (D.A.) allows the search parameters to change block by block throughout the model to account for the local variations. Presented here is a study of a vein that changes orientation down-dip. A traditional resource estimate using one search ellipse orientation was performed. The resulting estimated grades did not accurately reflect those of the sample data. Using a search ellipse with an average value for the two distinct dips caused the grade estimation to be less representative than desired. Rather than performing domain-based analysis with different search ellipses, the issue was corrected by using Datamine Studio 3?s D.A. function. The 3D vein trend was used to estimate the dip and dip direction of the vein in each block. The resulting D.A.-estimated model follows the two dip trends of the vein. This software makes a previously tedious task faster, and allows for more accurate estimations.

Feb 27, 2013

By E. S. Shedd

Misch metal was electrowon from a treated bastnasite concentrate in a REF3-LiF-BaF2 electrolyte at 950° C. The metal was prepared and tapped from a 1,000-ampere cell, and nearly 100-percent utilization of the rare-earth oxide was achieved. Prolonged operation in a smaller cell (40 amperes) indicated that the composition of the misch metal product was reasonably constant, and no significant increase of impurities, with the exception of carbon, was noted.

Jan 1, 1970

By A. Karzhau, B. Aldamzharov, Y. Tussupbekov

"This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of the mine 24 hours a day. The mine auxiliary ventilation has been designed with the maximum length of ducting systems with extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal. The mine needs additional investments for re-designing ventilation network and therefore needs to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation network optimisation processes at the operational metal mine in Kazakhstan. Mine ventilation surveys were undertaken for validation of the computerised ventilation model. The mine ventilation network modelling has been undertaken to be compliant with Kazakhstan mining regulations. Results of the mine ventilation optimisation processes were analysed in terms of financial risks for investments into reconstruction of the ventilation shafts, extension of the ventilation network, variable speed fans and automated regulators. Savings were resulted from improving ventilation system efficiency by reduce airflow in areas without mining activity and redistributing existing ventilation capacity. Based on savings derived from the optimised ventilation network an additional NPV has been received. Stochastic modelling use Monte Carlo simulation producing representation of possible outcomes for the additional NPV considering electricity pricing as a major variability. As result probability distribution for the optimised ventilation network demonstrates the additional NPV in a range of 1.6 - 2.2 million US$ in the corresponding defined value of 90%. Moreover, there is only 5% chance that the additional NPV to be less than 1.5 million US$. Conducted risk analysis for ventilation network optimisation processes gives considerations on volume of investments into the mine ventilation system.INTRODUCTIONCosts for mine ventilation are typically increasing when mines expand to deep levels. Some underground mines in Kazakhstan still use continuously operating fans at its maximum capacity. This is not helping to save energy and costs that are in many cases very significant. This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of a mine 24 hours a day. The mine auxiliary ventilation has been designed with maximum length of ducting systems and extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal from the mine workings (Sabanov, 2019). The mine needs additional investments for re-designing ventilation networks and therefore need to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation networks optimisation processes at the operational mine in Kazakhstan. Mine ventilation surveys for validation of the computerised ventilation model have been undertaken"

Jan 1, 2019

By J. Brent Hiskey

Thiourea is a reagent capable of complexing gold and silver. With the proper oxidant, thiourea can effectively dissolve metallic gold and silver. A thiourea leaching system possesses potential advantages over conventional cyanidation including acidic solutions instead of alkaline, cationic complexes in contrast to metal-cyanide anoionic complexes, and faster leaching rates with acidic solutions of thiourea. Generally, the rate of metal dissolution in thiourea solution is fastest when ferric ion is employed as an oxidant. Acidic thiourea solutions for certain applications could offer a nontoxic and noncorrosive option for processing gold and silver. This paper provides information regarding the fundamental chemistry of the acidic thiourea system, and reviews the dissolution of both silver and gold in acidic solutions of thiourea. The recovery of gold and silver from acidic thiourea solution is discussed.

Jan 1, 1981

By L. K. Weaver

The Bureau of Mines has started an engineering study of oil recovery by miscible-phase displacement. As a preliminary part of the investigation a literature and patent study was made to prevent duplication of work already done, to become acquainted with those actively engaged in research of this type, and to learn where the recovery method has been attempted. This bibliography gives results of this study. The bibliography includes 254 references to U.S. and foreign patents and published literature relating to miscible-phase displacement processes of oil recovery. It is divided into the following sections: (1) U.S. patents, (2) foreign patents, (3) literature references, (4) author index, and (5) publication and source reference. The patents and literature references are listed alphabetically by author by intervals of years. References to February 1963 are included.

Jan 1, 1963

Smarter Gravity Separation Using the Kelsey Centrifugal Jig With relatively coarse grained and easy-to-process ore sources rapidly dwindling, smarter techniques are becoming more necessary for the continued success of mineral processing operations, while increasing environmental pressures also require more environmentally friendly solutions. Gravity concentration has long been employed to pre-concentrate minerals using water, reducing or eliminating the need for less environmentally friendly processing techniques such as flotation. Conventional gravity recovery devices however, operate in a 1 g environment with limited efficiency in the separation of fine particles of heavy and gangue minerals. More recently developed centrifugal concentrators operate in a multi-g environment and enable finer heavy minerals to be recovered, but also exhibit limited separation efficiency due to the separations being based on solid particle mass differentials. Further enhancement of gravity separation has been achieved using the Kelsey Centrifugal Jig (KCJ), which also operates in a multi-g environment, but separates minerals on the basis of specific gravity differentials, using 'ragging' media and a pulsation mechanism, which maximises the effect of differential particle accelerations. This enables the efficient separation/recovery of minerals down to a particle size of ~6 ¦m and/or separation of minerals with a low specific gravity differential. This paper presents details on the principals of operation of the KCJ, recent improvements in the technology and recent developments in modular KCJ plants, as well as some KCJ application areas, which offer smarter, environmentally friendly processing opportunities for the mineral processing industry.

Sep 13, 2010

Radioisotope x-ray and neutron techniques of analysis can be used for the determination of a wide range of elements in minerals. The most important applications of these techniques to the mineral industry are for onstream analysis of slurries and solids, and for geological assay.The radioisotope x-ray techniques are the more developed ones and two installations for on-stream analysis are in routine use in mineral processing plants in Australia. Neutron techniques are being developed, and should prove valuable for geological assay of large numbers of mineral samples, for borehole logging, and for specific on-stream analysis when the cost of the equipment is warranted.

Jan 1, 1970

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By M. Rudolph, D. H. Hoang, N. Kupka

"Froth flotation has been subjected to modelling since the 1930s and there is a large variety of approaches that intend to portray the entire process. Among these, three fundamental models could be considered exhaustive: the Pyke model [1], the Yoon model [2] and the Schwarz and Koh model [3]. The first model makes use of the generalized Sutherland equation collision model, the Dobby-Finch attachment model and its own stability model, the second model considers an extended DLVO approach for the attachment and stability whereas the third model makes use of computational fluid dynamics. Using a scheelite ore as a test case, a specific batch flotation test was repeated ten times and analyzed for chemical assays, mineralogical studies and froth properties. The detailed data are used to feed the three flotation models aforementioned in order to compare the outcomes of the three approaches, their capacity to reflect reality and to predict it. This article intends to pave the path towards reliable process models incorporating in depth mineralogy information, e.g. surface liberation, particle size and particle shape. KEYWORDS Froth flotation fundamental models, scheelite, automated mineralogy, critical assessment. INTRODUCTION In mineral processing, only three approaches aiming at describing froth flotation fundamentally could be considered exhaustive: Pyke’s collision and sliding approach [1], Yoon’s kinetic approach [2] and Schwarz and Koh’s approach [3]. The models will hereafter be referred to as Pyke model, Yoon model and Schwarz and Koh model. The first model makes use of the generalized Sutherland equation collision model, the Dobby-Finch attachment model and its own stability model, the second model considers an extended DLVO approach for the attachment and stability whereas the third model makes use of computational fluid dynamics. All three approaches rely on the widely accepted idea that flotation is a kinetic reaction dependent on a flotation rate constant kp, as expressed in Eq. [4]:"

Jan 1, 2018

By Zheng Shuangying

The vibration effect of the buried oil pipeline caused by tunneling blasting affects the pipeline safe transportation, tunnel construction schedule and its construction cost. The dynamic response characteristic of the buried pipeline under blast loads is the foundation of designing blasting scheme as well as the prerequisite to pipeline safety. Taking the blasting engineering of Xiannvyan tunnel passing through a buried oil pipeline of the high-speed railway from Xi’an to Chengdu as the background, according to the space relation between tunnel and pipeline as well as the charge structure of V-cut blasthole, the 3D non-linear dynamic finite element model of tunnel-pipeline was established in the light of the equivalent explosive. The vibration response of the ground surface over pipeline resulting from V-cut blasting was simulated by LS-DYNA finite element program. Comparing with the measured vibration velocity waveforms in site, the accuracy and reliability of the numerical model are well verified. On the base of the simulation method, the coupling dynamic response characteristic between buried oil pipelines and soil was further studied and the Von-Mises stress distribution of pipeline wall was obtained. The safety of buried oil pipeline was quantitatively analyzed through the inside stress data of pipeline given by the simulation calculation above. The analytic results show that the maximum vertical vibration velocity of pipeline on the interface between pipeline and soil is 1.62~5.07 times as large as that of soil. When the buried pipeline is located over the tunnel and 13 m (42.6 ft) away from tunnel, the maximum vertical vibration velocity of ground surface caused by V-cut blasting with charge weight 17 kg (28.6 lb) is 3.17cm/s (1.25 in/s) and the vertical peak vibration velocity of pipeline top is about four times as the ground location directly above the pipeline. Meanwhile, the peak dynamic stress of the buried pipeline caused by blasting vibration is only 2.35% of the allowable stress of steel pipeline. The maximum superimposed stress of the blasting dynamic stress and the hoop stress produced by the pipeline internal pressure is 242.56 MPa (35 180.36), which implies that the buried oil pipeline is completely safe under the conditions of the given tunnel blasting scheme.

Jan 1, 2015

By R. E. Head

[This paper embodies the data obtained by a detailed statistical, microscopic analysis of the composite mill feed concentrate, and tailing from the Magne and Arthur mills of the Utah Conger Co. for 1 month in 1929. It is one of a series of papers by the same authors5 dealing with the mineralogical characteristics of the ore and concentration products from various well-known plants. The work was done in the cooperateive microscopic service laboratory at Salt Lake and represents the results of cooperation between the United States Bureau u of Mines, the Utah Copper Co. and the University of Utah.]

Jan 1, 1935

By R. E. Head

This paper embodies the data obtained by a detailed statistical, microscopic analysis of the composite mill feed concentrate, and tailing, from the Magne and Arthur mills of the Utah Conger Co. for 1 month in 1929. It is one of a series of papers by the same authors dealing with the mineralogical characteristics of the ore and concentration products from various well-known plants. The work was done in the cooperative microscopic service laboratory at Salt Lake and represents the results of cooperation between the United States Bureau of Mines, the Utah Copper Co., and the University of Utah.

Jan 1, 1935

By Eugene F. Eisemann

This paper is concerned principally with the competition between fuels for industrial use in the United States. Total energy consumption in the U. S. is allocated between industrial and other uses The industrial volume is divided into two classifications (1) where-there is competition between fuels and (2) where there is no competition. The pattern of fuel consumption by electric utilities was studied because there, is competition between fuels and because it is the fastest-growing industrial market. Electric utilities fuel consumption is discussed by geographic area. The growth of each fuel in several areas is investigated. The analysis of these areas, includes the study of the relationship between prices where fuels are in, competition:- (1) on the East Coast where price competition occurs between oil and coal, and (2) in the southeastern states where competition exists between coal and gas..

Jan 1, 1961

By Paul D. Merica

One of the most useful, and at the same time least commonly used, methods of thermal analysis for the determination of transformations in metals and alloys consists in the recording of the time intervals required for successive increments of temperature change during heating or cooling, the temperature of the furnace containing the specimen being altered at a uniform rate. The curve obtained by plotting these time intervals as a function of the mean temperature of the specimen durnig the interval is called the inverse-rate curve. It is probably due to the fact that no simple and convenient method has apparently been available for the measurement of the successive time intervals that this method has not been so generally used as, for example, the differential method, for which several types of automatic or semi-automatic apparatus have been designed. Whenever this method has been used, the intervals have usually been measured with the use of a chronograph; its operation as practised at the Bureau of Standards1 is as follows: The temperature of the specimen is measured by a thermocouple and a dial potentiometer. The operator sets the potentiometer at successive values of the electromotive force differing by equal increments, usually 0.02 millivolt, and records the exact instant on the chronograph, by pressing a contact key, at which the galvanometer coil passes through its null position. Two-second intervals are also recorded on the chronograph record, and the number of seconds elapsing between successive signals is afterward counted from the record and plotted as a function of the electromotive force or of the temperature. This is an admirable method and by it most minute thermal arrests may be detected, but it requires a good chronograph, which is generally difficult to obtain, and the time of one operator during the recording and, subsequently, to read the intervals from the record; the latter operation often takes from one to two hours. There is here suggested a simple, convenient, and, it is believed, equally accurate method of recording the successive time intervals, by which the

Jan 1, 1921

By B. Rai

Aluminum is produced by the Hall-Héroult process, which involves passing a current through a molten mixture of alumina dissolved in cryolite (Na3AlF6) to reduce the alumina to aluminum. This process occurs in an electrolytic cell comprising a steel shell lined with a carbon cathode and refractory bricks. After a certain period of operation, the carbon cathode deteriorates and is removed from the cell along with the refractory bricks. These removed materials are referred to as spent pot lining or ?SPL?. Typically, 30 Kg of SPL waste is generated for every ton of Aluminum produced. Globally aluminum smelters produce more than 500,000 tons of SPL each year. Even though SPL has a high fuel value (8000 to 9000 BTU/lb) and contains valuable metals, it is classified as a hazardous material due to its cyanide content and leachable fluorides. Various disposal/recycle techniques have been developed over the years for SPL but none of them is satisfactory. Thus, large quantities of SPL are still being dumped in landfills. In some countries, even the land filling of SPL is not permitted. The safe disposal of SPL thus remains a challenging problem to the aluminum industry and the search continues to find methods of converting SPL into a non-hazardous waste and/or recovering the valuable materials contained in the SPL. We have developed a process to produce hydraulic setting cements (eco-cement) containing a novel fluoride rich cementitious phase, from SPL by an energy efficient, low temperature clinkering route. A raw mix blend of SPL, lime/limestone, calcium sulfate and tailings sands based on a set of design criteria developed by us, is clinkered in a two stage process; 2 hours soak at 800 °C followed by clinkering at 1000 °C for another 2 hours, to produce a soft clinker. The SPL eco-cement is made by grinding the clinker to a fineness of 5000 cm2/g Blaine with 1% by weight borax as a set retarder. A good quality cement having as high as 20 % by weight of SPL in the raw mix (28 % by weight in the cement clinker) is thus produced. The physical properties of the SPL cements are found to be comparable to the Indian standard (IS) specifications for Ordinary Portland Cements (OPC). We have produced the cement in a few Kg batches and tested the reproducibility of results from batch to batch for further scale up. Once commercialized, this process has a potential of converting SPL into value-added cement products. Keywords: cement, spent pot lining, clinkering, fluoride, aluminum, ordinary portland cements

Sep 1, 2012

By Gus Van Weert

Solvent extraction of iron from zinc refinery liquors has been investigated before, mainly on the MEHPA / HC1 stripping combination. Leakage of chloride into the zinc electrolyte and the costs of handling the ferric chloride eluate have prevented this option from being adopted. This work explores the stripping of the ferric loaded organic phase with nitric acid to take advantage of the possibility of regenerating up to 6N nitric acid, and clean spheroidal hematite production by autoclave treatment of ferric nitrate solution. The work was carried out with DEHPA (di 2-ethyl-hexyl phosphoric acid) in kerosine. Degradation testing proved the DEHPA to be very stable, when exposed to nitric acid solutions up to 6N and temperatures up to 70" C over a period of five days. Ferric iron was removed completely froma synthetic hot acid leach liquor, irrespective of the DEHPA concentration. Stripping the ferric turned out to be the problematic step. Although the stripping efficiency was considerable better at 50 than at 20° C, it decreased with increasing DEHPA/Fe ration in the organic phase. Eluates of no more than 5-10 g/l ferric nitrate were produced. surprisingly, nitric acid was found to be a less efficient stripping agent than sulphuric at equivalent normality and its efficacy levelled off above 4N. It was concluded that DEHPA does not appear to be the S.X. reagent to take advantage of the low cost autoclave decomposition of ferric nitrate opportunity in zinc hydrometallurgy and that another S.X. reagent, also resistant to nitric acid, but based on weaker compound formation is needed.

Jan 1, 1998