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Leaving the relatively simple life on a uni student and venturing off in to the outside working world is a big step. There are two paths to follow, the æContikiÆ style graduate program where you pick up your new lease vehicle and set forth on two to three years of planned activities, or can pack everything into one bag, stick a copy of æthe rough guideÆ in your pocket, and set off on your own graduate adventure. I took the path less travelled.The less followed method of developing yourself isnÆt for everybody. Parts of the journey are difficult and sometimes it feels hopeless, but itÆs all a learning experience. Along the way you might not get to all the graduate ætourist stopsÆ like dragline, truck-shovel, or drill and blast. But youÆll see most of them and pick up a lot more of the local culture such as people management, operational knowledge, quality control, and the black arts of project financial management.

Jan 1, 2004

By Russell F. Palmer

"THE PRACTISING MINING ENGINEER has little time for the expression of opinion on abstract questions concerning the definition and future of mining engineering. The following comments indicate some of the thoughts and suggestions of those who have found mining engineering a very satisfying field of endeavour. The future of mining engineering has been under discussion for perhaps twenty years. It is significant that the University of Saskatchewan program was in- 88 traduced at the time that some traditional mining schools were questioning the need for mining engineers. One Canadian university, and some American schools, actually abandoned the field at this time, some ten years ago.Five years ago, Mining Magazine sponsored an international debate on training for the future, and concurrently, the CIM undertook a major survey of the mineral industry's manpower needs. The current CIM discussions should prove more fruitful, as they are directed toward mining engineering alone, rather than the broad, and confusing, spectrum of all professionals and semi-professionals of the mineral industry."

Jan 1, 1972

By Yutaka Shiraishi

Two methods were developed to measure the viscosity of the slag from the glassy to the liquid states in an ascending temperature process. The first method employs rotating parallel plates within which is placed a solid cylindrical sample in line with a sphere. The second method uses a tube which is pushed into a sample held in a rotating crucible. Both methods were tested using standard glasses, including boron trioxide and NBS standard glass #711. The main difference between these two methods is in sample preparation. Comparative measurements between two methods were performed using a sodium boro-disilicate glass containing 5%mol of alumina. The results showed good agreement in the mid-range of glassy viscosities, 105 to 107 Pa.s. In the liquid state, there were no significant differences between the two methods.

Jan 1, 2003

By Yutaka Shiraishi

Two methods were developed to measure the viscosity of the slag from the glassy to the liquid states in an ascending temperature process. The first method employs rotating parallel plates within which is placed a solid cylindrical sample in line with a sphere. The second method uses a tube which is pushed into a sample held in a rotating crucible. Both methods were tested using standard glasses, including boron trioxide and NBS standard glass #711. The main difference between these two methods is in sample preparation. Comparative measurements between two methods were performed using a sodium boro-disilicate glass containing 5%mol of alumina. The results showed good agreement in the mid-range of glassy viscosities, 105 to to7 Pa.s. In the liquid state, there were no significant differences between the two methods.

Jan 1, 2003

By Bo R. Arvidson

The strongest magnetic roll separator available on the market enhances certain industrial minerals purity to levels not reached previously. Exceptionally strong forces can also be utilized for processing relatively fine particles at feed rates several times greater than for prior art electromagnetic induced roll separators. Since the unique machine does not work with an air-gap, coarser particles than what has been considered practical can be separated at rates up to 20 tph per single machine unit and 40 tph for a double machine unit. Examples for all these three application areas will be given. While the roll separator can be used only for dry processing presently, the other new separator can be used for both wet and dry applications. It is a medium-intensity, high- gradient magnetic separator employing a filamentary matrix. It was primarily developed for recovery of weakly magnetic iron ore fines, but it is finding applications for cleaning of industrial minerals such as wollastonite, calcite, clay and talc. Its main advantage is a simple design resulting in a very lm cost machine both with regards to total capital investment and operation. A significant technical advantage exists also. The metallurgical performance is similar to conventional electromagnetic separators, while the cost per ton is about half of such separators.

Jan 1, 1985

By Brian Hart, Hong Zhong, Liuyin Xia, Kyle Douglas

"The application of hydroxamic acid as flotation collectors is fast becoming of significant interest for the selective flotation of oxide minerals along with upgrading oxidised based metal ores and minerals. In the hydroxamate molecule, both O and N atoms in the -CONHOH group are identified as coordination atoms in many complexes and their characteristic structure determines the strength of the chelation complex between the hydroxamate and metal ion. The chelating performance of hydroxamates with transition metals including Cu, Ni, Zn, Fe, Cr, Mn and Co, the alkaline-earth metals (Be, Ca) and heavy metal (Pb) has been studied. Cytec has developed and patented a range of alkyl hydroxamates as flotation collectors for oxidised base metal sulphides, metal oxide minerals and precious metals. This paper introduces two new structures of hydroxamate collectors (Cyclohexyl and 4-tert-butylbenzoyl hydroxamates) and evaluates their characteristics relative to alkyl and benzoyl hydroxamates. Laboratory small-scale flotation tests were conducted with ilmenite and wolframite. The use of 4-tert-butyl benzoyl hydroxamate resulted in a best flotation performance for ilmenite. A relatively low cyclohexyl hydroxamate dosage was required in the flotation of wolframite to obtain comparable results to the alkyl and benzoyl hydroxamates. It is also shown a significantly improved flotation performance with 4-tertbutyl benzoyl hydroxamate in the flotation of ilmenite plant ore in comparison to the oleic acid or fatty acid collectors typically used for ilmenite flotation.INTRODUCTIONHydroxamic acid has been reported as an effective collector in small-scale flotation tests for a wide variety of oxide minerals and ores including copper (Lee et al., 1998, Lee et al., 2009), iron (Quast, 2000), tin (Sreenivas and Padmanabhan, 2002; Wu and Zhu, 2006), tungsten (Marinakis and Kelsall, 1987; Xia et al.,2004), and rare earth metal minerals (Ren et al., 1997; Pradip and Fuerstenau, 1991; Assis et al., 1996; Pavez et al., 1996; Zheng et al., 1996; Chen et al.,2005).As part of our previous research project, a selected suite of hydroxamic acid collectors including fatty, aromatic and modified hydroxamates (naphthoyl, cyclohexyl, benzoyl, octyl, 4-tert-butyl benzoyl, salicylic hydroxamates) were tested in order to examine their potential for recovery and selectivity of rare earth minerals. In these tests it was shown that the flotation performance with the hydroxamate collectors was significantly better than the phosphoric acid collectors on no slimes removed samples (Brian, et al., 2014; Liuyin, et al., 2014 and unpublished internal technical reports). Furthermore the tests revealed that for most of the hydroxamate reagents tested, their dosage requirements to achieve recoveries at the same levels was less than that for the phosphoric acid collectors. Recovery of both La and Ce was significantly better for the salicylic hydroxamate and that Nb, Y and Zr recovery were favoured by the octyl and benzoyl hydroxamates (Brian, et al., 2014; Liuyin, et al., 2014). Comparison among the hydroxamate collectors showed that the flotation and recovery performance of cyclohexyl (CHH) and 4-tert-butyl benzoyl (TBH) was not as good as the other hydroxamates for rare earth recovery."

Jan 1, 2015

By J. W. Francis

The paper about to be presented deals with the relatively straightforward problem of mining two seams of coal separated by an interval. This may seem to be a simple problem but within a twenty-mile radius of Consolidation Coal Company's Burning Star # 4 Mine at least seven (7) different techniques of two-seam mining have either been tried or are now in use. This great variety of techniques should point out that there is no single simple solution; a conclusion often reached when dealing with mining problems. As a background for the techniques presently in use at the Burning Star # 4 Mine, we will briefly outline the other five (5) stripping techniques used in the area: 1. Captain Hine 6360 Range Diagram (Figure 1) 2. Leahy Mine Range Diagram (Figure 2) 3. Burning Star # 3 Mine 5860 Range Diagram (Figure 3) 4. Burning Star # 3 and # 4 Mines Dragline in Pit Range Diagram (Figure 4) 5. Captain Hine 2570/5561 Range Diagram (Figure 5) Each of these techniques was chosen because it provided the best and least expensive alternative for mining the coal under the particular restraints encountered. The Mine Engineers at our Burning Star # 4 Mine, Mr. Dennis Mahan and Mr. William Weaver, will now detail two additional mining techniques, out-lining our experience to date.

Jan 1, 1977

By Larry M. Southwick

Considerable process and equipment development and analysis were conducted during design of Freeport Sulfur's laterite nickel facilities in the late 1950's. These covered ore concentration at Moa Bay, Cuba, and concentrate refining at Port Nickel, Louisiana. Both facilities utilized acid pressure leaching and product refining steps that were similar to those later used elsewhere in the laterite processing industry. Unfortunately, developments in Cuba removed the Moa Bay facilities from Freeport control. This deprived the project of a start-up that would have allowed thorough resolution of difficulties and a quicker attainment of production capacity. This meant lost opportunity to positively prove and complete the technology when optimum resources and staff were available. This paper reviews the original rationale of how design was developed via laboratory, bench scale and pilot studies by Freeport Sulfur. Some later operating problems will also be discussed. This case is unusual, but the lessons learned may be instructive in the application of difficult and complicated technologies, as well as the need to understand and retain access to the original development and design activities Keywords: Moa Bay, Freeport Sulfur, Port Nickel, acid pressure leaching, pilot plant, pachuca, Chemical Construction, pressure letdown, cobalt, titanium

Jan 1, 2004

By T. B. Armstrong

"The energy situation is extremely fragile and there is a critical need for energy conservation in the short term and development of new technology in the long term. To keep tabs on energy flow, separate electrical metering is being installed in major industries: The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production. IntroductionThe current energy situation is best described as being extremely fragile. The supply of hydrocarbons is being exhausted, while energy needs remain the same in the industrialized nations and are increasing in the Third World.The petroleum era is expected to come to an end in this century, as there will no longer be enough. Gas, coal and electricity will replace some oil as it becomes scarcer during the transitional period before new sources are gradually developed.There is a critical need for energy conservation in the short term. Conservation will, to a great extent be justified by savings, as it will be less expensive to recover energy that is being wasted than to acquire additional energy.The gradual development of new technology is expected to increase energy self-sufficiency in the long term. New energy sources are attracting a growing interest and are becoming more economically feasible and marketable. These rediscovered energy sources include peat, wood, solar, domestic waste and windmills.The sudden increase in energy prices has forced users to look at the way energy is used and find ways to save it. Major industries have found that using separate meters for each department spurs energy responsibility. Energy trim programs now rank in importance with profitability, productivity and with quality.Energy has become such a crucial cost element that better data are essential for planning. Data acquisition varies with plant size and number of utilities. Meters can be read manually or monitored by computerized energy management systems.The relationship of electrical energy input and production levels for a typical uranium mine will be used to demonstrate two steps to energy management. The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production.The direct benefit of saving energy is the money saved."

Jan 1, 1980

By R Justa Cámara, L M. Pinillos Lorenzana

Barcelona Metro Line 9 (Line 9) is an essential infrastructure, improving the transversal transport in Barcelona by connecting the greater part of the currently existing radial lines. The purpose of Line 9 is to create a peripheral line surrounding the city of 1.6 million people (metropolitan area exceeds five million), and includes 52 stations, of which a total of 32 are deep shafts and 19 transfer stations in a densely populated area with several geographical constraints.The tunnel has an inner diameter of 10.90 m, bored with an earth pressure balance (EPB) tunnel boring machine (TBM) in soil sections with precast concrete lining type universal ring and segments 40 cm thick.A single tunnel solution was adopted, driven by functionality and costs savings. The internal tunnel section is divided into two levels to allow for one way rail transit including the platforms in each station on each level in order to optimise the space created by the TBM excavation and comply with the operational needs. The division in two areas of the tunnel was carried out through the construction of a horizontal slab resting on the tunnel segments with a constant thickness of 40 cm at the central part of the slab and supporting side walls.CITATION:Justa Cámara, R and Pinillos Lorenzana, L M, 2014. Two transit levels in one tunnel tube – an innovation in Barcelona Metro Line 9, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 387–398 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By Wolfgang Guetter, Wolfgang Weber

The upgrading project for two hydropower plants in Slovenia, Europe, requires two new pressure tunnels: Plave II (5.9 km Flysch) and Doblar II (3.6 km Limestone). It is the worldwide largest application of the Double-Shield TBM technology combined with an extremely thin-walled monoshell honeycomb lining system. The first drive has been completed by end of June 2000 achieving best performances of 50m/d and almost 1,100m/mth. In September 2000 the TBM started the second tunnel without any required adjustments due to geological or other reasons. Another 52m/d have been bored already within one production day of two 8 h shifts achieving a maximum penetration rate of more than 6m/h. The completion of the Doblar drive can be expected by February 2001.

Jan 1, 2001

By Armin D. Ebner

A hetero-f1occulation model is developed to calculate the net force exerted on a paramagnetic nanoparticle that is moving in the proximity of a magnetite particle that is fixed or supported within an open flow and magnetically energized packed bed. Important variables investigated include the size and surface charge of the particles, the magnetic field, the ,flow velocity, the electrolyte concentration and the magnetic susceptibility of ,the nanoparticle. The ?influence: of other naturally occurring forces, such as hydrodynamic and DLVO(Le., van der Waals and electrostatic) forces on the magnetic attractive forces between the magnetite particle and the nanoparticle are also examined. This model reveals the feasibility of using magnetite particles instead, of stainless steel as the energizable element for high gradient magnetic separation.

Jan 1, 1999

By E. D. Gálvez, L. A. Cisternas, R. Acosta-Flores, F. A. Lucay

Froth flotation processes are carried out in flotation cells that are grouped into banks, and these banks are interconnected, forming a flotation circuit. A literature review shows the existence of papers related to flotation circuit design based on mathematical programming. However, due to the complexity of solving the mathematical model in most of the work, it is considered that a small number of species is present in the feed to the circuit, which differs from practice. In addition, simple bank models are generally used. This paper presents a methodology for designing mineral concentration circuits that overcomes the problems mentioned. It allows the use of more suitable cell or bank models and the inclusion of several species. The methodology is based on two phases. The first phase identifies the set of optimal structures using discrete values of stage recoveries, solving several mixed integer linear programming (MILP) problems. In the second phase, the optimal design for each of the structures obtained in the previous phase is determined using a suitable model for the recovery at each cell or bank, which results in a mixed integer nonlinear programming (MINLP) model. The design of a copper concentration plant with eight species and the design of a zinc concentration plant with three species and five size fractions by species are used to validate the proposed methodology. The structure of the cells in the rougher and cleaner banks deliver structures that are novel.

By R. Y. K. Yang, O. Olajide, E. H. Cho

"This study focuses on partial leaching of coal pyrite with sodium sulfite in order to increase its degree of liberation from the coal matrix, which might help to increase the removal of coal pyrite, together with trace elements of arsenic (As), selenium (Se) and mercury (Hg) by froth flotation. The U.S. Bureau of Mines’ two-stage reverse flotation process was applied to further reject coal pyrite in the second stage, where pyrite was floated while coal was depressed. The results show that the leached coal flotation with sodium sulfite produced lower pyritic sulfur content in the clean coal by approximately 25% compared to the flotation of unleached coal. The synergism mechanism between the leaching and the two-stage flotation is delineated. The rejection of trace elements (As, Hg and Se) after the two-stage flotation process with unleached coal followed approximately that of pyritic sulfur (64.6%). However, after the two-stage flotation with leached coal, the rejection of Se was as high as that of pyritic sulfur (77.9%), while the rejections of As and Hg were 20% lower than that of pyritic sulfur, or followed approximately the same rejections obtained through the two-stage flotation with unleached coal.IntroductionSO2 is a reducing reagent, but when it is combined with oxygen in an acidic solution, it will become a strong oxidizing reagent for leaching coal pyrite. Sundaram et al. (2001) found that the reagent could leach coal pyrite to produce ferric iron, which, in turn, catalyzed the oxidation of SO2. Similarly, Cho et al. (2009) found that the reagent could leach pyrite as well as mercury from coal. In the present study, coal was preconditioned by partially leaching coal pyrite in water-coal slurry with sodium sulfite at 48° C. The partial leaching of coal pyrite may enhance its degree of liberation from the coal matrix and give a chance to be rejected more effectively during the flotation.The two-stage U.S. Bureau of Mines’ reverse flotation process (Miller, 1975) was used in this study. The first stage is conventional flotation, in which coal is floated while material with higher ash and pyrite content remains in the cell. Kerosene is used as a promoter to float the coal and MIBC (methyl isobutyl carbinol) is used as a frother. In the second stage, the floated coal from the first stage is depressed using a starch-based reagent, while coal pyrite is floated using potassium amyl xanthate. In this second stage, the liberation of coal pyrite from the coal matrix seems to be an important factor for successful rejection of coal pyrite."

Jan 1, 2015

By J. A. Brown, F. P. Kerr, M. A. Archer

Extensive testwork on Global Atomic’s Dasa Project in Niger has been conducted at SGS Minerals laboratory in Lakefield, Ontario. Much of this testwork has focused on a two-stage counter-current leach as a means of improving leach efficiency and decreasing reagent consumption. The ore contains uranium mineralization requiring relatively high sulphuric acid concentrations during leaching to achieve acceptably high extractions. Excess free acidity in the leach solution may be utilized by recycling the high acid solution and contacting it with fresh ground ore. The resulting pregnant leach solution (PLS), now higher in uranium concentration and depleted in free acid, is forwarded on to uranium recovery by solvent extraction. The first stage residue proceeds to the second stage leach where fresh acid and oxidant are provided to leach the remaining uranium. This paper presents the results of the testwork conducted on both a Master Composite and Variability samples.

Jan 1, 2020

By W R. Honeywell

Two series of laboratory flotation tests were carried out to study the treatment of Faraday ore by two-stage flotation. First the acid-consumers and then the uranium minerals were floated. On ore ground in the Faraday mill grinding circuit, and using Faraday mill water in the flotation step, a recovery in the flotation concentrate of from 91 to 93 percent of the uranium was obtained in 45 to 50 per cent of the weight at a grade of 0.24 per cent U30s. The saving of acid in leaching the uranium-bearing flotation concentrate, as compared to leaching the whole ore, ranged from 30 to 46 per cent. The average over-all extraction resulting from flotation and leaching of the flotation concentrate was 88.1 per cent. The best collector for the acid-consumers was sodium oleate. Acintol F A-1 emulsion was the best collector for the uranium minerals, and, in the uranium flotation step, an acid pH in the slurry was found to be necessary. The reagent costs were estimated to be 22 to 26 cents per ton of ore.

Jan 1, 1963

By A. L. Mackie

Arsenic is a highly toxic element and a known human carcinogen. It is often a contaminant in mine water discharges, particularly from gold mining and roasting operations, and has been found at concentrations as high as 4 g/L. Increasingly stringent controls on the concentration of arsenic allowed to be discharged into the environment cannot always be met by current treatment processes and thus enhanced technologies are required to achieve low treated effluent concentrations at reasonable costs. A novel two-step physicochemical treatment process was developed and tested at the bench-scale using synthetic mine water solutions (SMWs) in order to reduce arsenic concentrations in treated effluent to exceed current discharge regulations. The process includes chemical coagulation with ferric sulphate and ballasted flocculation for enhanced solid-liquid separation. The SMW contained 59 ± 2 mg/L arsenic with a slightly alkaline pH. The effect of arsenic speciation on the process was tested by using either arsenite, As(III), or arsenate, As(V), in the SMW solutions. Tests were performed at cold temperature (3 ± 1 °C) to ensure the process? efficacy year-round in northern climates. Arsenic in As(V) SMW solutions was removed to a final total concentration of 0.0054 ± 0.002 mg/L, almost half the current recommended allowable drinking water concentration of 0.010 mg/L. As(III) was removed to a concentration below the Metal Mining Effluent Regulations? (MMER) current limit of 0.50 mg/L, to 0.32 ± 0.06 mg/L, with double the coagulant dose used in Stage 1 for As(V) removal. Oxidant addition at a concentration of 5 mg KMnO4/L in Stage 2 resulted in final total arsenic concentrations of 0.017 ± 0.002 mg/L and pH adjustment with lime in Stage 2 resulted in 0.0415 ± 0.0007 mg As/L in treated As(III) SMW. Both of these modifications to the process reduced arsenic concentrations to below proposed new MMER limits (i.e., 0.10 mg/L). This two-stage treatment process was shown to reduce arsenic concentrations to well below current treatment guidelines while reducing or eliminating chemical oxidant demand.

Aug 1, 2013

By P. Karimi, G. Mollataheri, A. Amini, H. Abdollahi, A. Mirzaeian

"Chah-Palang tungsten deposit is one of the two tungsten indexes in Iran. Based on ore characterization techniques (including optical mineralogy and XRD analysis), quartz, chlorite, atacamite, azurite, malachite, ferberite, illite, goethite and calcite were distinguished as major minerals. The representative sample was analyzed qualitatively by using X-ray fluorescence and chemical assay. Results showed that grade of tungsten oxide and copper were 1.76 and 7%, respectively. Also precious metals such as gold with grade of 1 ppm was reported in the sample. By using the mineralogical studies on the prepared sample, the liberation degree of tungsten minerals in –0.6 mm size fraction was obtained 80%. Several tabling tests were carried out on the original sample which at the optimum conditions and as a result, the grade and recovery of tungsten in concentrate reached to proximately 23 and 59%, respectively. To increase the grade of tungsten in the final concentrate, acidic leaching were performed on the tabling concentrate and its appropriate results indicated that the grade of tungsten increased to 27%. Also more than 90% of copper was extracted as a byproduct in the fallowed process.INTRODUCTIONTungsten, a unique hard, heavy heat-resistant metal, is distributed erratically in economic concentrations throughout most continents, occurring mostly as small irregular deposits that are not readily adaptable to large-scale, low-cost mining methods. The diverse physical characteristics and mineral composition of tungsten ores also present problems in beneficiation, requiring complex hydro- and pyrometallurgical treatment to yield high-purity concentrates suitable for the production of ferro alloys, carbides and super alloys in the steel industry (Hausen, 1989; Kholmogorov & Kononova, 2005). Tungsten is most frequently used as tungsten monocarbide, which has hardness close to diamond, in cemented carbides. The most well-known use of tungsten is in lighting filaments, and other important applications include improving steel properties and use in tungsten alloys, electronics, and chemical products (Hausen, 1989; Kholmogorov & Kononova, 2005)."

Jan 1, 2016

By Tomasz Niedoba

Coals occurring in nature are characterized by various properties. According to Polish norms, the classification of coals taking into consideration their calorific value contains coals from type 31 (flaming coal) until type 43 (meta-anthracite). The samples of coal of type 31 were taken for investigation. To this purpose, the approximations of ash contents and density distribution functions were done for this coal type (determination of marginal distribution functions). They were the basis for application of so-called Morgenstern distribution functions to approximate the function Z=L (?), where Z is ash contents and ? - coal density. Then, the comparison between obtained results was performed and the results were statistically evaluated. They are the basis for the further research of other types of coals.

Sep 1, 2012

By J. C. Motz

Introduction The variations in crushing plant design seen during the past 30 years provide a valuable source of information and experience from which one can select the optimum arrangement for a specific requirement. The purpose of this paper is to examine some specific equipment configurations as they relate to operating efficiency, plant availability and final product size in the form of mill feed. Due to the varied ore properties, physical space limitations, size of ore body and financial resources, there is no single crushing or grinding flowsheet that is best. Since energy costs have risen dramatically during the recent past, it became obvious that significant savings are possible by selecting equipment and a circuit which offers low power requirements balanced with low wear metal and installed equipment costs. Relatively clear and measurable are the metal and maintenance costs and the ever increasing cost of energy. Metal costs in the form of crusher liners have not quite doubled in the last 20 years, while power costs have increased by a factor of four or five, depending on the region considered. Estimates made only a few years ago on the increased cost of energy are obsolete in view of the recent unstable conditions. Since power is therefore becoming the largest cost area, it behooves us to investigate potential cost saving methods in the form of lowering the KWH/Ton of crude. Crushing To reduce run-of-mine to mill feed requires reduction in a series of stages. The physical size and power requirements of a crusher capable of reducing hard rock vary depending on the application. Extended experience in most crushing plants show that a minus 19MM (3/4 in.) rod mill feed can be and is being produced in three stages of crushing. In the majority of cases, the last crushing stage is closed circuited using anywhere from 19MM (3/4 in.) to 14MM (9/16 in.) slotted openings on the sizing deck. In some copper concentrators, the mill feed is all passing 13MM (1/2 in.) using three stages of crushing and 14MM (9/16 in.) square openings on the closed circuit screens. It was felt by some that crushing to this fine size is only possible if the ore is friable and is relatively soft in comparison to the harder ores. Our results show that some of the copper ores are as hard as taconite and are crush-

Jan 1, 1982