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By Clarke DS

The igneous intrusions in SE Papua New Guinea are part of a sub-volcanic to plutonic, high-K calc-alkaline to shoshonitic suite. High-K andesitic to dacitic intrusives on Misima and Woodlark Islands, and shoshonitic intrusives in Milne Bay and the eastern Owen Stanley Ranges are related to late Cenozoic andesitic volcanism. Both suites are associated with epithermal Au-Ag deposits and weak porphyry Cu-Au mineralisation. All significant alluvial and lode gold occurrences in the region are spatially related to intrusions. Intrusions in the D'Entrecasteaux Islands and near Mt Suckling are core plutons related to basement doming, and appear to have no association with either volcanism or hydrothermal mineralisation. The core plutons are not petrochemi:ally distinct from the volcano-re:-ed Intrusions.

Jan 1, 1991

By Norman F. Schulz

The technical excellence of separation achieved in a mineral concentration process, or any other process where two constituents of any kind are physically separated from each other, is expressed uniquely and quantitatively by the separation efficiency (E,): E. = (R - R,), where R is the percentage of the valuable constituent and R, the percentage of the waste appearing in the concentrate. The separation efficiency meets all the usual requirements of a true efficiency and is independent of the nature of the process used. It can be visualized as the percentage of the feed material that, in effect, actually undergoes complete separation while the rest of the feed is distributed randomly and unchanged into the two separation products. The problem of quantitatively assessing the results of a mineral concentration process recurs whenever one tries to decide which of two similar test results is more desirable. Although economic factors may greatly influence the eventual choice of process, they usually have very little to do with the technical excellence of a separation and are, therefore, not a consideration at this point. Concentrate grade and recovery, used simultaneously are the most widely accepted measures of performance, and if both the grade and the recovery are greater for one case than for another, the choice is simple. However, when the results of closely competitive operations are compared, it often occurs that one set of test results shows a higher grade and a lower recovery than another and the choice is no longer obvious. Laboratory investigations of a mineral concentration process applied to a given feed material can often be designed to yield successive paired values of concentrate grade and recovery over the practical range of the course of a separation process.' These results may be plotted as some combination of the cumulative values of grade, weight recovery, or unit recovery in the concentrate. The resulting correlations for two slightly different sets of conditions can then be compared qualitatively, or even quantitatively in particular cases.2 Such comparisons are not applicable, however, to individual concentration test results. The reduction of a separation test result to a single efficiency figure, preferably as a percentage of ideality, would afford the simplest and most precise means for comparing the results of competitive tests. The effective proportion of feed material that undergoes complete separation in any real concentration process, hereafter referred to as "separation efficiency," is such a figure. It is essily calculated from the usual feed and product data and can even be evaluated, in many cases, from data in publications describing various concentration processes. Mineral Concentration Processes and Their Evaluation Mineral Concentration Processes: The overall concentration process may be subdivided into a liberation process which severs the bonds between unlike mineral species, and a separation process, which sorts the resulting liberated particles into two products—a concentrate rich in mineral A and a tailing lean in mineral A. A perfect mineral concentration process is represented by the idealized scheme shown in Fig. 1 in which liberation and separation are not differentiated. The feed mixture consists of two mineral types which are completely liberated early in the process and then sorted by type. All the mineral particles of type A which respond uniquely to the selection forces acting in the ideal separator, report to the concentrate, while all the mineral particles of type non-A report to the tail-

Jan 1, 1971

By N. H. Emmons

In the United States the value of gold- and silver-ores is everywhere reckoned in ounces troy of the metal per " short ton " (2000 lbs. avoird.) of the ore. In the case of silver, which fluctuates in market value, the New York quotation on a given day must be known, in order to determine the gross value of an ore at that date. In Mexico the assay is reported 'in kilogrammes of silver or grammes of gold per metric ton. As everyone is aware, this is practically a percentage system, 1 kg. of silver per metric toll (1000 kg.) being 0.1 per cent., and 1 gramme of gold per metric ton, 0.0001 per cent. The notation is simple and convenient for both assayer and miner.* The Mexican government purchases at an unvarying price for silver, payable in silver, and at a fixed price for gold, payable in gold if the seller insists, but preferably in silver at about the rate of exchange between U. S. gold coin and Mesican silver coin. The value of silver is $40.915 per kilogramme in Mexican silver dollars. That of gold is $675.416 per kilogramme in gold coin. The mints do not, as a rule, issue gold dollars, as the government keeps them to pay the interest on the national debt. Almost all bullion containing a considerable percentage of gold is shipped to foreign countries (principally the United States) to be refined. The Mexican mint receives for refining and coining no bullion which is not at least .950 fine in gold and silver, separately or

Jan 1, 1902

By L. M. Pidgeon

"Part I. - PrinciplesIntroductionIf the oxides of the metals are listed in decreasing order of the magnitude of their heats of formation, the list appearing in Table 1 will be formed. It will be apparent at once that this is essentially the well-known electrochemical series (Table li). The metals at the bottom of the list exhibit great stability and are attacked only by the most violent of chemical reagents. Higher on the list, the metals become more reactive until, at the head of the list, the metals lithium and calcium are unstable in moist air.The reactive metals to be considered lie above manganese. In this group will be found familiar and abundant metals such as aluminium and magnesium, the unfamiliar yet abundant metal titanium, the rare and mysterious metal beryllium, and others of greater or less interest.While these metals exhibit properties as unlike as those of uranium and magnesium, they have in common a specialized and difficult extractive metallurgy. They form compounds of great stability, and their reduction is to varying degrees difficult. The position of a metal in Table 1 (and more particu1arly its position in the free energy column) gives a numerical value to this property.Oxides with low free energies of formation may be reduced with great ease; thus silver and mercury may be produced from their oxides by simple heating in air.Lead, iron, and zinc may be separated from their oxides by heating with carbon to moderate temperatures. Magnesium oxide, on the other hand, .requires a high temperature and very special conditions before it will yield magnesium metal."

Jan 1, 1950

By C. H. Kats, E. G. Meiter, J. J. Bloomfield

"INTRODUCTION.Street manholes for underground power and signal circuits, and for sewers vary in size from small openings that a workman can enter, to large central vaults. As the gas hazards in manholes are similar to those in mines, the mine safety committee of the Bureau of Mines was asked whether the miners' flame safety lamp is a practicable device for indicating inflammable gases in manholes.The coal miner detects firedamp (methane or CH4) with his flame safety lamp by observing the luminous can or the increased length of flame that appears when 1.5 per cent or more of the combustible gas is present. The lamp flame also warns him of lack of oxygen by going out when the oxygen becomes less than 16 to L7 per cent. A tentative rule regarding the proposed use of safety lamps at manholes has been suggested previously as follows: ""Testing for gas: Do not enter manholes until you have .assured yourself that the manholes are free from dangerous gases, by testing with an approved safety lamp, by ventilation, or by other adequate methods."" The Bureau's committee recommended that miners' flame safety lamps are not suitable,testing in manholes because hydrogen gas may be formed therein. The flame safety lamps approved by the Bureau of Mines for use in aseous mines are safe in presence of methane and while not tested against the members of the paraffin hydrocarbon series (CnH2n+2) higher than methane they been have been used therein; the lamps should not be used in hydrogen or acetylene they can initiate explosions in presence of these gases."

Oct 1, 1925

By Hamid-Reza Manouchehri

Mining & Sustainability? A Paradox?! ? Production: gradually or exponentially increasing (giant mining) ? Ore Grades: gradually declining, ? Open Cut Mining: now widespread, likely to be sustained in the future though the long-term is hard to predict as new mineral deposits are likely to be deeper ? Waste Rock / Overburden: increasing rapidly, likely to be sustained in the future and closely linked to open cut mining, especially for coal and base metals ? Economic Resources: commonly increasing but some remain stable or gradually declining, future is linked closely to exploration, technology and economics; Increasing population, changing life style and technology development, ultimately result in sustainability in mining industry to hang in the balance and continue.

May 1, 2013

By Patterson B. G

The small party of mining men who formed the original Mount Morgan Syndicate in 1882, commenced operations on the top of the mountain with what were practically nothing more than small hillside quarries, It was not until 1891 that underground stoping began, and open-cut mining therefore commenced with the commencement of operations. This hillside quarrying, however, seems to have been quite haphazard and unsystematic ; there is no trace or sign of any properly devised mining scheme. This can be seen plainly enough from an examination of old mine plans; for in less than five years after stoping was; adopted, the open-cut was down into the stoped-out ground. The following table, prepared from these old plans, shows the downward trend of the excavation in the past fiftyseven years:-

Jan 1, 1939

By R. W. Watson, H. D. Peer

"This paper outlines the history of steel linings in the Autogenous Mills, the factors influencing the decision to convert to rubber lining, and the operating and cost experience with the rubber liners to date. INTRODUCTIONSherman Mine is an open pit iron ore mining, concentrating and pelletizing operation located at Temagami, Ontario. The operation employs 392 hourly employees and 81 salaried employees and has an annual capacity of over one million L.T. of high quality iron ore pellets produced from low-grade magnetite ore.The Sherman Mine is a joint venture; 90% owned by Dominion Foundries and Steel Limited (Dofasco) of Hamilton, Ontario, and 10% owned by the Tetapaga Mining Company, a wholly-owned subsidiary of the Cleveland Cliffs Iron Company of Cleveland, Ohio. The operation and management of the mine is by Cliffs of Canada, Limited. Development of the property was started in 1965, with pellet plant production beginning on March 17, 1968. At the present time, the orebodies contain sufficient reserves to sustain operations at current production levels for 30 years. Mining was originally carried out in three orebodies, the North, South, and West pits. The North Pit has been exhausted and mining operations ceased in the spring of 1980. A fourth pit, the East Pit, will be developed during the early 1980's."

Jan 1, 1981

By M S. Powell

In this paper it is argued that we are entering a new mining era, based on considerably more challenging orebodies as we rapidly expand minerals production by exploiting previously uneconomic deposits. Rapidly increasing production costs cannot be sustained by continually increasing commodity prices, as illustrated by the recent downturn in the industry.Driven by reduction in grade, requiring the processing of 50 to 100 per cent more ore to recover the same amount of metal, often accompanied by finer grain structure - necessitating a finer product size and more challenging recovery - the productivity of the mining industry is dropping in real terms. The author suggests that tweaking current technologies and relying on economies of scale will not turn the industry around, but rather a switch to considerably more efficient processing is required.It is imperative to utilise the properties of the orebody, especially in terms of variability, in revolutionising our processing techniques. Our new mantra may be æselective mass mining with flexible processingÆ. This paper explores how this may be implemented using current technology and equipment in order to change the cost profile of the new mining industry. Orebody knowledge, based on the underlying process drivers, rather than proxies, is required. These properties and their variability will need to be propagated into the orebody through utilising the geological data rather than smoothing block averages, which destroys information on variability.Novel flexible circuit designs are presented that can respond to a wide variability in ore properties and throughput requirements in order to slash processing costs and energy utilisation, while enhancing overall recovery. The success of these circuits will be based on an ability to predict at both design stage and in everyday processing the variable ore properties, which is dependent upon reliable orebody knowledge that caries variability and underlying processing properties.CITATION:Powell, M S, 2013. Utilising orebody knowledge to improve comminution circuit design and energy utilisation, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 27-36 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By G. van den Berg

Anglo Platinum has initiated a project to investigate hydraulic hoisting technology for platinum mines. Hydraulic hoisting systems are categorized as pumping, fluidizing feeder or displacement feeder systems. Positive displacement piston diaphragm pumps are most suitable for pumping systems but they require that the ore be milled prior to pumping. Fluidizing feeders are unsuitable for platinum ore applications due to the large percentage of fines in the ore. The three-chamber pipe displacement feeder has been most widely used for hydraulic hoisting and is the most suitable technology for platinum mines. Eight hydraulic hoisting applications covering pumping and displacement feeder technologies are reviewed. The key factors to be assessed in selecting a hydraulic hoisting system for a platinum mine are discussed. A typical platinum mine configuration will comprise an ore preparation plant (to produce -15 mm material through screening or crushing), motive water pump on surface, high pressure water pipeline from surface to underground, ore slurrification plant, three-chamber displacement feeder, and a high pressure slurry pipeline to surface. The water and slurry pipelines are housed in a purpose drilled borehole. A hydraulic hoisting system selection chart is presented. A 150 mm diameter system with a 25 MPa pressure rating is required to hoist 60 000 t/m from a depth of 1 200 m. Generic capital and operating costs are presented . A major potential advantage of hydraulic hoisting is that the fines, which contain a high percentage of PGMs, are captured underground and transported directly to the concentrator. Hydraulic hoisting technology is ideally suited to providing supplementary hoisting capacity to existing shafts that have reached the skip hoisting capacity limit. The main advantage for new shafts is the potential to significantly simplify underground rock handling and transportation. All hydraulic hoisting system components are proven. The only risk area is the wear rates of the valves and piping. Current estimates are based on gold ore. The business cases investigated show that the financial viability of the installations are not sensitive to wear rate. Hydraulic hoisting is technically feasible and financially attractive for platinum mines. Keywords: Hydraulic hoisting, slurry

Jan 1, 2004

By E. M. Bloomfield, C. A. Jermy

Geotechnical tests were carried out within an area of the coastal sand dune cordon of northern KwaZulu-Natal which is currently being exploited through mining of heavy mineral sands. In situ and laboratory tests were carried out to assess the density, strength and stability of the dune sand material before and after mining. The most stable combination of sand (>63µm) and fines(<63µm) material was determined as well as the effects of compaction on the strength and permeability of the sand dune samples. In situ DCP and density tests showed that the sand dunes fall in density immediately after mining has taken place. However, settlement of the dunes occurs with time and the density increases within the top 2 metres, and particularly once the dunes become re-vegetated. Laboratory tests carried out showed that the dune sand with between 20% and 30% added fines had the highest shear strength and dry density, but the permeability was shown to decrease steadily with increasing fines up to 40% fines added. The addition of fines would result in steadily increasing water flow problems. Compaction of the dune sands was shown to significantly increase the density and the shear strength, but with a corresponding dramatic decrease in the permeability of the material. Slope stability analysis carried out on models of the dunes with increasing fines showed that the addition of fines increases the stability of the dune sands up to a peak of about 30% fines added, and compaction further increases the stability of the dune sands. Despite the benefits of compaction on the stability of the tailings dunes, the loss of permeability is thought to outweigh the benefits as well as the high costs involved. The stability of the tailings dunes greatly increases after re-vegetation. Compaction and densification of the tailings dunes may be necessary within the top 2 m interval if large-scale construction takes place on the sand dunes. Petrographic analysis was performed to understand the granular interactions of the sand and fines fractions, and the effect of the in situ weathering of the dune sands on the stability of the sand dunes. As the fines contents increase the gaps between the sand grains are filled, until a point where the percentage of fines begins to push the sand grains apart.

Jan 1, 2003

One of the most important goals of the experiments is to determine deformation and strength of the rock specimen under applying triaxial load, the status of failure condition is one of the subjects which can be used in soil mechanic and foundation engineering applications. To investigate the effect of confining pressure on the triaxial test, the rock has been considered as bonded-particles method and has been simulated by particle flow code three dimensions (PFC3D). This model, does not have pre-existing microcracks and all the particles have the same size (2 mm diameter), equal to the real specimen shape. To compare the mechanical behavior and failure pattern of the numerical and experimental results at uniaxial and triaxial tests at the same conditions, the Young&apos;s modulus, Possion&apos;s ratio as well as maximum axial stress have been considered. On the other hand, the microcracks growth and change of failure pattern at the modelling of the uniaxial and triaxial tests with different confining pressures up to failure point were reported. According to the experimental work, the number density of microcracks decreases from starting value and remains almost constant up to the failure beyond that point. The comparison of the numerical and experimental results of maximum axial stress and the Possion&apos;s ratio revealed good accordance. The initial number density of microcracks at the model was zero and increased while microcracking. Also, the sudden increment around the maximum stress was observed which is because of unstable growing of microcracks near the maximum stress. In all uniaxial and triaxial tests with different confining stresses, the number density of microcracks during the failure remains almost constant which can be considered as a proof of failure occurring in the model. The experimental results indicate the similar trend as well.

Nov 25, 2010

By Xiaofei Guan, Adam C. Powell, Uday B. Pal, Srikanth Gopalan

This paper reports a solid oxide membrane (SOM) electrolysis experiment using an LSM– Inconel inert anode current collector for production of Mg and O2 at 1423K. The electrochemical performance of the SOM cell was evaluated by using various electrochemical techniques including electrochemical impedance spectroscopy, potentiodynamic scans, and potentiostatic electrolysis. The effects of Mg solubility in the flux on the current efficiency and the YSZ membrane stability during SOM electrolysis were discussed and examined through experiment and modeling. The electronic transference number of the flux were measured to assess the Mg dissolution in the flux during SOM electrolysis. A negative correlation between the electronic transference number of the flux and the current efficiency of the SOM electrolysis was observed.

Jan 1, 2014

Separation of Nickel and Cobalt from Manganese, Magnesium and Calcium by Synergistic Solvent Extraction - From Batch Tests to Pilot Plant Operation The synergistic solvent extraction system consisting of Versatic 10/LIX63/TBP was used for the separation and purification of nickel and cobalt from synthetic laterite leach solution after iron removal. Very encouraging results were obtained for separating nickel and cobalt from manganese, magnesium and calcium. In batch tests, the organic composition of the synergistic system was optimised and very large synergistic shifts for nickel and cobalt were observed. The ?pH50(Mn-Co) and ?pH50(Mn-Ni) values were as large as 1.65 and 2.40, respectively. The extraction and stripping kinetics of nickel and cobalt was&nbsp;fast. Semi continuous extraction tests were conducted with a pH profile of 5.5/6.1/6.5 in three stages at 40¦C. The nickel and cobalt extractions reached 99.9 per cent with only 5 mg/L Ni and &lt;1 mg/L Co left in the raffinate. With two stages of scrubbing, only 2 mg/L Mn and less than 1 mg/L Mg and Ca were left in the scrubbed organic solution. With two stripping stages, the stripping efficiencies of Ni and Co reached over 95 per cent. A fully continuous pilot plant operation was conducted successfully for 280 hours. With an O/A ratio of about 2 and a pH profile of 5.5/5.8/6.0/6.3 for the four stages EX1/EX2/Ex3/EX4, both nickel and cobalt were almost completely extracted. Using a pH profile of 5.4/5.0 for SC1/SC2, the manganese scrubbing efficiency was over 96 per cent. Using three strip stages, over 98 per cent Ni and 99 per cent Co were stripped with only 64 mg/L Ni in the stripped organic solution. The nickel concentration in the loaded strip liquor reached 83 g/L, resulting in a ?Ni of over 28 g/L. The chemical stability of the system was also assessed over 12 weeks under proposed industrial operating conditions. Operation exclusively under strip conditions at 50¦C and 40¦C resulted in LIX63 oxime half lives of 1.1 and 2.1 years, respectively. Operation exclusively under extract conditions at 50¦C and 40¦C resulted in oxime half lives of 0.26 and 0.82 years, respectively. Losses of Versatic&nbsp;10 were less than four per cent.

Sep 13, 2010

By A. M. S. Hussein, B. E. Rogers

Progressively more stringent emission limits and advances in regulations and enforcement have made air compliance more difficult for metallurgical facilities. This has historically required improved collection and treatment of stack sources. Fugitive emission reductions are now increasingly required to achieve compliance. Typically small in magnitude, fugitive emissions have a disproportionate impact on ground level concentrations, can be visible to the public and can influence workplace health and safety. Moreover, these emissions are complex, as they are spread over multiple sources, distance and time and are affected by various activities and conditions. As such it is difficult to find, define and monitor these emissions and their impacts. This difficulty translates to the design of emission reduction measures and operational requirements. This paper presents an overview or ‘heuristic’ of fugitive emissions management for compliance in existing operations. This includes aspects such as the impact of regulations on project scope, approaches to developing a comprehensive emission inventory, considerations for solution development and the plant’s role in sustaining compliance once achieved. Case studies are presented to demonstrate how portions of this approach have been applied.

Jan 1, 2015

The Aredor mine employs conventional dragline methods to exploit alluvial diamond deposits. Production began in 1984 with the policy of mining in the terrraces during the wet season and in the river flats during the dry season. Early production failed to meet predicted grades. A major geological programme to re-evaluate existing reserves was undertaken. This resulted in considerable increases in reserves in the river flats. Many of the reserve blocks included flats on both banks of the Baoule river. Methods of access into the river flats, normally flooded in the wet season, were developed to allow mining throughout the year. River diversions have been the main method used to provide access into the reserve blocks and, most importantly, to remove the risk of flooding. Aredor has carried out six main river diversions with a total length of 12 km. The diversions allowed production planning to be made on the basis of the economic evaluation of the reserve blocks, without undue regard to natural ground conditions or to seasonal factors. The Manning equation was used for the calculation of channel dimensions and gave satisfactory results. The practical aspects of the diversion preparations and flooding prevention are described

Sep 1, 1991

By Steven W. Wilson, Brigitte Lacouture, Nichola A. McKay

"The Aqqaluk deposit is a lead-zinc ore body that lies immediately north of the Main Pit at Teck Cominco’s Red Dog Mine in Alaska. It is scheduled to begin production in 2011 or 2012. A project team was formed in late 2004 to plan and implement an ore characterisation and metallurgical evaluation on the Aqqaluk ore. The main objective of this program was to facilitate the transition from Main Pit to Aqqaluk ore in the Red Dog concentrator. The program will develop a detailed, spatially-oriented characterisation of the Aqqaluk deposit attributes including lithology, ore geochemistry, bulk mineralogy, ore texture, as well as evaluating the metallurgical characteristics of ore hardness, process water geochemistry, bench-scale floatability and mineral liberation in ground feed, concentrate and tails. Developing correlations between the metallurgical response and the ore characterisation data is key to the emerging science of geometallurgy. This on-going program is a technical due diligence study that aims to reduce the risk associated with processing a new ore and to provide predictive information for mining and stockpile preparation.The test work to date has focused on characterising the five ore zones currently identified in the Aqqaluk deposit, namely: Siliceous (sub-divided to High Zn:Fe and Low Zn:Fe), Vein, Median Plate and Baritic ores. It identified a number of challenges associated with surface oxidation, or weathering, of the core samples and outlined a program for future sample collection and storage. Metallurgical testing evaluated ore hardness by a variety of methods. Preliminary floatability studies have been completed and some initial links have been established between the ore characteristics and the flotation response. Soluble mineral weathering products, the relative abundance of pyrite and marcasite and the iron content in the sphalerite grains have all shown a correlation to mineral recovery.As new core is received, both the ore characterisation and the metallurgical testing will be continued with a view to strengthen the existing correlations and develop new links."

Jan 1, 2007

By A Scott, A C. Torrance

Explosive properties are typically summarised on manufacturers’ technical data sheets. The values most commonly referred to are the density of the explosive, the energy it is expected to release and its velocity of detonation (VOD). This information is provided so that blasting engineers can make a considered selection of the most appropriate product for a particular blasting application. As an example, a lower VOD and energy product is better suited to blasting in soft, weathered material. In contrast, blasting hard, competent material generally requires a higher energy product which is usually characterised by higher VOD and energy.A review of the technical information supplied by manufacturers shows a wide range of methods of reporting both VOD and energy. A meaningful comparison between products from different suppliers is almost impossible to achieve based on these data. For a similar product, one supplier may have a VOD range of 2800–5800 m/s and a relative bulk strength of 1.34 while a corresponding product from another supplier is quoted with a VOD range of 3100–6300 m/s and a relative bulk strength of 192. In the end, the user has to rely on the level of confidence available from their experience using a certain type of explosive in the type of rock to be blasted at that mine.This paper reviews some of the technical and supporting information that is freely available in the literature and attempts to provide some clarity as to what the numbers that are reported actually mean for the end user. An alternative approach to the evaluation of the suitability of a given explosive to blast a given rock mass is proposed.CITATION:Torrance, A C and Scott, A, 2015. What is relative about energy?, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 447–454 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By T. Gammatikopoulos, C. Gibson, M. Aghamirian

Lithium is an alkali metal found in brine deposits in the United States, Chile and Argentina, as well as in hard rock deposits in Australia and North America. In hard rock deposits, spodumene (LiAl(Si2O6)) is the major economic lithium bearing mineral. Spodumene is typically upgraded through a combination of dense media separation, flotation and magnetic separation. One of the major spodumene concentrate quality specifications is the iron oxide content because it will help determine if the concentrate is considered ceramic grade for use in ceramics and glass or chemical grade for use in lithium ion batteries. The rejection of iron bearing silicate minerals from artificial ore composites (high amphibole and low amphibole, high quartz/feldspar) from a hard-rock spodumene deposit was investigated using magnetic separation and flotation. The efficacy of Wet High Intensity Magnetic Separation (WHIMS) in iron rejection was found to be dependent on the mineralogy of the iron-bearing silicate minerals including modal distribution, liberation and association. When magnetic separation was conducted prior to flotation, the Fe2O3 grade in the final flotation spodumene concentrate was reduced from ~3% to ~1% Fe2O3 , and improved lithium recovery in flotation was observed. It was determined that it was most effective to conduct WHIMS on the flotation feed after scrubbing and de-sliming because this reduced the required magnetic field to recover iron-bearing minerals, likely because slime coating was greatly reduced. Flotation collector screening of four different collectors with varying rosin acid content demonstrated a potential increase in selective flotation of spodumene over some iron silicate minerals with decreasing rosin acid content in the fatty acid collector. However, due to marginal difference in flotation results with the various collectors tested, further testing is required to confirm the observations made and establish a hypothesis regarding the effect of rosin acid content.

Jan 1, 2017

By Edward Keller

The work covered by this report was undertaken at the request of Dr. J. A. Holmes, late Director of the Bureau of Mines, to whom the writer had been recommended by C. W. Goodale and E. P. Mathewson, of Butte and Anaconda, Mont. The director's offer was accepted for the honor it conferred and for the useful results that might be promulgated by the work in the technical field covered. The investigation was facilitated by permission to visit various works and sometimes by contributions of important definite information on a given subject. It is therefore with pleasure that acknowledgment of courtesies is made to the following corporations and gentlemen: Anaconda Copper Mining Co., Raritan Copper Works, Baltimore Copper Smelting & Rolling Co., United States Refining Co., American Smelting & Refining Co., B. B. Thayer, A. C. Clark, A. L. Walker, William H. Peirce, E. A. Cappelen Smith, H. H. Alexander, F. C. Newton, S. Rolle, R. "V. Deacon, N. W. Pierce, O. C. Martin, C. N. Sappington, James K Thomson, W. G. Derby, and the assistants of the writer, K. 1V.McComas and W. L. Raup. In this report considerable space has been devoted to general and theoretical principles on the ground that under modern conditions the best results in any field of activity can be attained only with the aid of an adequate knowledge of the best theories underlying the practice. The diffusion of such knowledge in the field of sampling metallic metallurgical materials appears to have been quite insufficient, and it is felt that when the principles involved are once thoroughly comprehended their universal application will be appreciated, so that it will make no difference whether pig iron or gold bullion is offered for test. It may further be noted that a discussion of the theories underlying practical methods is placed under the caption "Practical Sampling," with the idea that no one can be a practical sampler unless he is capable of analyzing his methods. No attempt has been made to advocate anyone method, but it is hoped that the analyses of all of the methods in use will show where each may be correctly and safely applied. The fact that it is possible to demonstrate the theoretical correctness of several methods indicates the invalidity of the claim that any single one of them is the standard.

Jan 1, 1916