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By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By B Mead, N D. Pascoe

The Grasberg block cave rail haulage system will be located in the Grasberg mining district on the island of Papua, Indonesia at an elevation of almost 3000 m above sea level. It will transport 160 000 t of ore per day from multiple chute galleries to the underground crushing systems. The design entails approximately 23 km of rail (ballasted and embedded track), over 120 ore chutes to deliver ore to the muck cars and three unloading stations to deliver ore to the crushers. The rail haulage system is a closed-loop system entirely within the underground mining complex.The GBC rail haulage project will be a substantial operation. Trains consist of 11–12 bottom-dump 20 m3 muck cars driven by a 40 t electric locomotive, about 400 t per trainload. Approximately 12 trains and one spare are required to maintain full production at 160 000 t/d. All rail installations will be 1435 mm (standard) gauge with AS60 steel.The rail system is designed to operate in full automated mode, including chute loading, hauling and dumping. The operation will be controlled from a surface control room.The rail haulage project has been divided into five stages, which will be brought online over the course of ten years. Current development and construction are focused on Stage 1 for initial production that will occur beginning in late 2017. This stage includes 24 chutes in three chute galleries, one unloading station, workshop areas and almost 7 km of track (3 km of embedded and 4 km of ballast).This paper describes the key components of the rail haulage system and reports on the current status of development and construction of the system. This paper also overviews the challenges, successes, issues and risks of this project in areas such as scheduling, development, construction and operation.CITATION:Pascoe, N D, Mead, B and Vega, H, 2016. Grasberg block cave haulage project – design and construction update, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 585–598 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By C. M. Rule

Stirred milling using either horizontal or vertical mills has made a rapid entrance into PGM ore and tailings concentrator flow sheets. Currently there are in excess of 40 units installed in either mainstream, ?MIG? or intermediate concentrate regrind, ?UFG?, applications in PGM, primary ore treatment plants or in tailings scavenging and reclamation treatment plants. This represents a total of approximately 70 MW of installed fine grinding equipment. This paper outlines the reasons why the technology take-up has been rapid and illustrates the advantages of this technology over conventional milling. Keywords PGM, platinum group metals, UG2, Merensky, platreef, stirred milling, MIG?mainstream inert grinding, UFG?ultra fine grindingof intermediate flotation concentrates, IsaMill?, Metso SMD?, Deswik, ceremic grinding media, QEMSEM, MLA.

Jan 1, 2011

By N. A. Abdel-Khalek

The amenability of carbonate separation from a sedimentary phosphate ore using a bioflotation process was studied. Two types of bacteria were chosen for this study. The experiments were performed using oleic acid as collector. The flotation experiments were carried out using statistical designs for optimizing the main operating parameters. A concentrate containing 0.78% MgO and 30.15% P2O5 with a recovery of 92.31% was obtained from a feed containing about 2.45% MgO and 27% P2O5. These specifications could not be obtained using conventional flotation experiments under similar conditions in the absence of bacteria. Measurements of zeta potential, adsorption of bacteria and collector, as well as froth power, were performed to explain the role of microorganisms in the surface modification of both dolomite and phosphate and in their interaction with the collector.

Jan 1, 2009

By H Han, W Wang, E Zhou, D Duan

"In this study, physical, chemical and microscopic characteristics of high phosphorus oolitic hematite were studied firstly. Based on which, the new methods of high phosphorus hematite used in the æsolid state reduction by rotary hearth furnace - magnetic separationÆ and iron nugget by rotary hearth furnace direct reduction processes were discussed. The results showed that high temperature, low basicity and high carbon content is conducive for the apatite in iron ore to reduce into molten iron, but not conducive to the removal of phosphorus. Through optimisation of process parameters, dephosphorisation rate of high phosphorus hematite is more than 60 per cent by æsolid state reduction roasting-magnetic separationÆ process while that is up to 80 per cent by iron nugget process. This study can provide theoretical and technical basis on economical and rational use of high phosphorus oolitic hematite.CITATION:Han, H, Duan, D, Zhou, E and Wang, W, 2013. New methods for high phosphorus oolitic hematite dephosphorisation by direct reduction process, in Proceedings Iron Ore 2013, pp 391-396 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

The Acid Leaching Process for the Synthesis of Zeolitic Material from Paper Sludge Ash

Sep 13, 2010

By A. J. Partos

The figures and tables referred to herein are presented in the Proceedings; Volume I, Pages 309 through 316. The 61-story One Liberty Place, at 315 m., is the tallest building 1n Philadelphia and was in 1987 the ninth tallest building in the world. One Liberty Place is supported on drilled piers socketed in heterogeneous metamorphic rock. Design load distribution was 40% on the base and 60% by shaft resistance. Drilled Pier G-10 supporting a T13-A type outrigger column was instrumented. There 1s no uplift load on the exterior outrigger columns. This column supports a 53,850 kN compressive load which includes a 13,440 kN windload. Table 4 presents experimental parameters from straight line interpolation of Mohr failure envelopes for the Wissahickon schist. Lowest values of cohesion (12,200 kN/m2) and angle of internal friction (19°) are at a 50° - 59° angle of foliation. Figure 3 presents a typical pier supported by shaft resistance 1n decomposed (D) and coreable (C) rock and by end bearing (B).

Jan 1, 1991

By P Knight, P A. Walker, H E. Speight

Pushback 8 South is the final cut back on the south wall of the East Pit at BHPBilliton?s Mt Whaleback iron ore mine in the Pilbara region of Western Australia. Development of the cutback commenced in 1999 and is now nearing completion with ultimate pit walls up to 300m high. This paper describes the geotechnical, structural and hydrogeological setting of PB8S, reconciliation between design and slope performance and the slope engineering techniques used to effectively manage the risks associated with the slope. These techniques include slope buttressing and sequential mining with waste backfilling, as well as a detailed program of displacement and hydrogeological monitoring, structural mapping and slope stability assessment. During the course of slope formation recent developments in remote slope displacement monitoring using radar and remote structural mapping using terrestrial photogrammetry have also been applied. Aspects of the PB8S slope management plan are relevant to many large scale open pit slopes. This paper provides geotechnical practitioners with information which may be helpful in managing geotechnical risk in similar environments.

Jan 1, 2006

The common approach of designing blasts on a trial and error basis is quickly coming to an end. When utilizing the full scale blast environment, trial and error can quickly become cost prohibitive and leaves much to be desired in terms of true blast diagnostics. In these situations testing is usually only continued until a perceived improvement is achieved and when the gross undesirables in blast results are removed. This does does not preclude, however, that further improvements of up to 100% or more can not be obtained.

Jan 1, 1991

There are many ways to classify structural slurry wall elements, but the DFI committee chooses to limit these descriptions to three functional requirements, namely: design function, plan configuration and type of panel reinforcement. (See Figure 1 in Part IV for panel classification). Slurry walls can function as: 1. Curvilinear or linear elements for temporary and/or permanent structures to resist lateral forces transferred from the ground, water, earthquakes and various surcharge loads. 2. Load bearing elements in various plan shapes to resist vertical forces. 3. Combination Elements to resist combined forces under conditions 1 and/or 2. For example, elliptical shafts with various openings for tunnels and conduits have been commonly used for deep excavations.

Jan 1, 2005

By Mohamed Ourriban, Yongjun Peng, Qi Liu, Denis Cotnoir

"In the flotation of fine particles, there are two difficulties in general: low recovery of valuable minerals and high recovery of gangue minerals. The low recovery of fine valuable minerals is attributed to low bubble-particle collision and attachment efficiencies, while high recovery of fine gangue minerals is attributed to mechanical entrainment through water films between gas bubbles. Over the last five years COREM has been actively developing fine particle flotation techniques with the objective of improving fine particle flotation separation efficiencies.The fine particle flotation techniques with both generation of micro bubbles and enlargement of sizes of fine particles have been developed to increase bubble-particle collision efficiency. For oxidized fine sulphide minerals, both physical and chemical techniques have been developed to modify the hydrophilic surfaces and therefore increase bubble-particle attachment efficiency. To reduce mechanical entrainment of fine gangue minerals, a new approach, polymeric depression and flocculation, has been developed. By selecting the high molecular weight polymeric depressants to make fine gangue minerals hydrophilic and also enlarge their sizes, entrainment can be minimized. By combining these techniques, we are able to improve both flotation recovery and selectivity in some fine ore flotation systems.INTRODUCTIONWith depleting mineral resources and their declining grades, ores leaving grinding circuit for froth flotation are becoming increasingly finer to achieve mineral liberation. Together with the increasing demand for minerals, ever-increasing tonnages of fines are being produced. Therefore, flotation of fine particles is an immense challenge to be tackled by researchers all over the world today."

Jan 1, 2006

By John C. Machisak

INJURY DATA and experience at coal mines in the United States for 1956 and 1956 are presented in this publication with a minimum of interpretation and comment because of the numerous tables required to afford detailed comparisons between the 2 years. Corresponding table., for 1955 and 1956 follow in consecutive order under the following general sections: (A) General injury experience: Tables I to II: contain current data on number of fatal and nonfatal injuries, frequency and severity rates, and average severity. (B) Selected injury data: Tables 12 to 77: Contain current data on causes of injuries, distribution of injuries by part of body injured, clays of disability, extent of hand and machine loading, experience by employment and production groups, and various other operating factors and physical conditions related to the occurrence of injuries at the Nation's coal mines. (C) Injury experience, by States: Tables 78 to 87: Contain data on injuries, injury rates, employment, cause of injuries, aid severity rates by States. (D) Major disasters: Table 88: Contains data on major disasters, 1947 -56. A major disaster is defined as a single accident in which, or as the result of which, five or more persons are killed instantly or died at any time thereafter. (E) Historical coal-mine-fatality experience: Tables 89 to 93: Contain data on fatalities and their causes at bituminous-coal and anthracite mines from the earliest records through 1956.

Jan 1, 1960

By E. M. Shemang, C. E. Suh, L. Mbinkar, University of Yaounde

"Two iron deposits within the Precambrian mineral belt of Cameroon are described in detail for the first time: the Archean Metzimevin replacement iron deposit enclosed in Fe-enriched itabirite, and the Proterozoic granite-hosted, shear zone-related Mayo Binka magnetite deposit. In the Metzimevin de¬posit, quartz is corroded and microplaty hematite overprints martite-textured hematite. The Mayo Binka massive magnetite veins show evidence of deformation in magnetite (fracturing and microbrecciation) and in overgrowths of specular hematite (mechanical twinning and undulating extinction). The magnetite is partially replaced by hematite (martitization) and goethite. Ores from both deposits have >88% total Fe2O3 and low contents of contaminants such as SiO2, Al2O3, MgO, CaO, P2O5, and TiO2. They are also poor in Cu, Pb, Zn, V, Cr, and Ni. The Metzimevin massive hematite is characterized by an accentuated light rare earth element depletion relative to the Fe-enriched itabirite. It is suggested that the Metzimevin iron deposit is the result of hypogene leaching of gangue minerals from, and further hematitization of, an itabirite protore. Although the genesis of the Mayo Binka massive magnetite is unclear, it is spatially related to Neoproterozoic granitic rocks. The data allow some comparison of these little known, but pot¬entially economic iron deposits, with some of the world’s better investigated deposits, and are useful to the exploration efforts for iron ore currently underway in Cameroon and the Central African subregion.Deux gisements de fer de la ceinture minérale Précambrienne du Cameroun sont ici décrits en détail pour la première fois : le gisement de fer Archéen de Metzimevin formé par remplacement et encaissé par des itabirites enrichies en Fe, et le gisement de magnétite Protérozoïque de Mayo Binka as¬socié au cisaillement dans un granite. Au gisement de Metzimevin, le quartz est corrodé et de l’hématite en microplaquettes se développe au dépends de l’hématite a texture martitique. Les veines de magné¬tite massive de Mayo Binka présentent des évidences de déformation dans la magnétite (fracturation et microbréchification) et dans les surcroissances d’hématite spéculaire (Macles tectoniques et extinction ondulante). La magnétite est partiellement remplacée par l’hématite (martitisation) et par la goethite. Le minerai des deux gisements contient >88% de Fe2O3 total et un faible contenu en contaminants tels que SiO2, Al2O3, MgO, CaO, P2O5, et TiO2. Il est également pauvre en Cu, en Pb, en Zn, en V, en Cr et en Ni. L’hématite massive de Metzimevin est caractérisée par un appauvrissement relatif en terres rares légères relativement à l’itabirite enrichie en Fe. Il est suggéré que le gisement de fer de Metzimevin résulte du les¬sivage hypogène des minéraux de la gangue ainsi que de l’hématitisation poussée d’une préconcentration de fer dans l’itabirite. Bien que la genèse de la magnétite massive de Mayo Binka demeure obscure, elle est spatialement associée aux roches granitiques Néoprotérozoiques. Ces données permettent de comparer ces gisements de fer peu connus et potentiellement économiques avec certains des gisements mondiale¬ment mieux connus, aidant ainsi les efforts d’exploration présentement en cours en Afrique de l’ouest et au Cameroun."

Jan 1, 2008

By D. R. Wardrop, G. Walton

Paper presented at the UK minerals industry conference held in Leeds, UK, 3-5 April 1995. A description is given of the approach that has been taken to the design of the 550 000 000 t superquarry with a planned life of 60 years that is proposed for development. The quarry has been designed to replicate a corrie and a sea loch and will include excavated slopes more than 570 m high. The main geological and physical controls on the layout and the choice of development method are outlined. The operation phasing is presented and the geotechnical constraints on slope design and the feasibility of alternative quarry limits are discussed. Environmental considerations have been dominant in planning and the quarry has been specifically designed for closure

Jul 1, 1996

By M. Tangstad

Keywords: Pyrometallurgy, ROMA, Norway, furnace, fines The ROMA project is a competence project on optimizing the usage of raw materials and energy in the ferromanganese, silicon, and aluminium industries. One of the items investigated in this project is the effect of fines on the furnace, and the usage of agglomerates versus lumpy ore. From lab-scale experiments it is shown, that for the charge to agglomerate in the furnace, the charge has to be exposed to a temperature higher than about 1200°C, independent of fines content. The fines less than 1 mm also showed higher agglomeration properties, compared to 2?4 mm, which showed no agglomeration property at all. The properties of agglomerates are very dependent on the pre-treatment of the agglomerate. Sinter and pellets, which have been heated to high temperatures, show a better strength, but also a lower CO reactivity. Sinter also shows a lower melting temperature compared to the ore.

Jan 1, 2011

By M. Cornelius

Drilled shaft construction through caving soils has historically been an uncertain, risky and costly operation. Uncertain, because the zone of caving may or may not be known, risky because the loss of a shaft is disastrous, and costly because the Contractor, who faces these risks, must include large sums of money in his bid for shaft wall stabilization techniques. This case study will discuss the stabilization techniques that are typically used within the United States, and then present new technologies, born in Europe, that are being utilized in the States by Malcolm Drilling Company today.

Jan 1, 1999

By Arcy, D&apos Banister

The productive potential of the Nation's principal active and inactive silver mining districts has been evaluated, using 1964 economic conditions and assumed silver prices ranging up to $3 per ounce. On this basis, the United States has an estimated 5 billion ounces of unmined silver, excluding minor byproduct silver in gold deposits. Approximately 3.5 billion ounces are contained in deposits requiring higher silver prices to be economically minable, and about 1.3 billion ounces are in currently operating mines. An analysis of silver consumption trends since 1954 indicates that the free world industrial market will climb to nearly 800 million ounces annually by 1985. U.S. industrial silver consumption has been projected to increase at an annual rate of 4 percent from 1965 to 1985. Average annual increase for the remainder of the free world has been 8.1 percent since 1953; this is expected to level off with an estimated average increase of 5 percent to 1975 and 4 percent from 1976 to 1985. Foreign silver production has, in the past, exceeded consumption, but in 1960 the supply deficit was over 100 million ounces. U.S. production, also traditionally exceeding demand, was suddenly overtaken at the start of World War II. By 1964 the annual deficit, excluding coinage, was over 84 million ounces. World coinage requirements of silver are fast declining and expected to total only 25 million ounces annually by 1985. In fact, coins are already becoming major short-term source of silver to fill the large deficit between industrial demand and current production. The larger, long-term deficit (to 1985, the limit of this study) must be balanced by silver from the hoards and new production from silver deposits. Eventually all of the deficit must be met by new production or silver prices will rise until consumption is greatly curtailed. To balance free world supply and demand, it is estimated that the free world must produce an additional 85 million ounces by 1975 and 375 million ounces by 1985. The estimated proportion expected from U.S. deposits is 17 million ounces by 1975 and 74 million ounces by 1985.

Jan 1, 1969

By F. Habashi

The changes that took place in China between 1984 and 2004 are remarkable. Visits were made to universities, research institutes, metallurgical industry, and many cultural centers within five trips during this twenty years period. The impressions of these visits are outlined.

Jan 1, 2007

By H. M. Lizama, R. L. Stephens

Secondary leaching of previously leached supergene copper ore has been developed and implemented at Teck’s Quebrada Blanca operations. This consists of using in-situ leaching techniques aided by geophysics to recover the residual metal value from dormant copper leaching heaps. These heaps of crushed and agglomerated ore had been leached 20 years previously but still contained significant copper values. The concept was first tested on a 3,200 m² sector, where injection wells introduced leaching solution into a production zone and electro-resistivity tomography was used to track solution movement within. Operating parameters such as well spacing and injection flow rates were determined. Copper recovery and leaching performance were evaluated based on solution and metallurgical balances. Both ferric leaching and bacterial leaching were enhanced through strategies to pull air into the heap. Positive results from the initial test prompted a larger, industrial pilot test on a 60,000 m² area and 97 injection wells. This larger test obtained a copper recovery of 30% in 457 days. Secondary recovery is now being implemented to all of the abandoned heaps at Quebrada Blanca and has become a production stream. This technology has the potential of wide application to mature copper heap leach operations, and Teck has patented the process and associated methodologies. INTRODUCTION Heap leaching of copper secondary sulphides has been carried out at Quebrada Blanca (QB) since 1994. Ore from the supergene deposit was crushed and agglomerated prior to leaching. Originally, the ore was leached in 6 m lifts stacked one on top of another, known as a static pad configuration. However, in 1998, after the second lift, the operation changed to a dynamic pad configuration; the ore was leached in a single lift and the resulting residues were removed to make way for fresh ore (Yáñez, Ahlborn, & Ramos, 2007). The first two static lifts were left in place and served as a foundation for the dynamic lifts above. Those two buried lifts of leached residues, some 16.8 million tonnes and covering an area of approximately 0.71 million m², remained buried under the dynamic pad.

Jan 1, 2019

By C. Dixon

The Abandoned Mine by C.C. Woollacott A heap of rocks marks the abandoned mine. The veld's unpitying silence lies around. Those broken stone?a mute and mournful sign Of human enterprise with failure crowned. Time with the slow hands, shall form the scene anew, Repair the gashes in the wounded soil, And cover up the last remaining clue To a poor useless record of men's toil. And this, the mound they built when hopes were high Shall be a grave, where those hopes buried lie. Outline This paper will consider the implications of the mine closure provisions set out in the new Mineral and Petroleum Resources Development Act and draft regulations. A comparison will be made with the existing provisions in the Minerals Act and the mine closure policy of the Department of Minerals and Energy (DME). The paper will consider whether the new provisions have the potential for improving performance as regards mine closure. It will investigate the financial provisions requirements and the implications of these provisions for mines. It will consider the requirements for a social and labour plan and consider the implications of this for mine closure and its funding. It will analyse the provisions relating to the transfer of environmental liabilities to third parties and consider the implications of this in the case of closing mines and non-closing mines. Finally it will consider the issue of residual legal liability of a mining company once a closure certificate has been issued.

Jan 1, 2003