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By Jeremy Bosman, Johan de Korte

Traditionally the feed to coal preparation plants was de-slimed at 0,5 mm. The +0,5 mm size fraction was processed via dense-medium cyclones and the -0,5 mm size fraction was beneficiated with froth flotation, water-only cyclones or spirals. With the introduction of large diameter cyclones, the aperture sizes of de-sliming screens have progressively increased to about 3 mm in the belief that the larger cyclones are not able to effectively beneficiate coal which has a particle size below the ?breakaway? size. The efficiency of separation in the large dense-medium cyclones is improved by presenting a coarser feed to the units. At the same time, the capacity of the de-sliming screens is increased due to the larger screen apertures and in effect, the throughput capacity of the dense-medium section is increased. The 3 x 0,5 mm size fraction, which was previously processed in the dense-medium cyclones, is now added to the fine coal size fraction and processed by water-only methods such as Teetered Bed Separators (TBS) and spirals. Whether this practice is the most economical way to process the 3 x 0,5 mm size fraction requires further thought since the best processing route for the 3 x 0,5 mm size fraction will depend on a number of factors, amongst which are: o The washability of the coal and the amount of near-density material o Processing efficiency o Specification of the required product o The price paid by the customer for the product o The cost of processing the coal Breakaway size The ?breakaway? size is defined as the particle size below which the Epm of the smaller particles starts to decrease significantly. Bosman1 provides an approximate breakaway size versus cyclone diameter (Figure 1). It should be kept in mind though that the Epm of the small particles is a function not only of the cyclone diameter but also of feed pressure, medium density, medium viscosity, particle shape and density, top-size of particles in the feed, etc.

Jan 1, 2006

By M. J. Adam

The Bureau of Mines is investigating the feasibility of partitioning complex sulfide minerals to simpler sulfide phases by sulfidation reactions. Such partitioning can afford opportunity to improve selective extraction of mineral constituents of value. This report describes research involving sulfidation partitioning of byproduct chalcopyrite concentrates from Missouri lead ore processing, and a Utah chalcopyrite concentrate from Cu ore processing. Small chalcopyrite samples were sulfidized with elemental S in glass capsules sealed after evacuating and backfilling with He, by heating for 1 to 24 h at specific temperatures in the range of 250? to 500? C, extending from below to above the boiling point of S (-4450 C).

Jan 1, 1987

By F. Sanchez, A. Mejias, C. Frias, G. Diaz

"Crude zinc oxides are the major source of secondary zinc and are produced from fuming processes like Waelz kilns. The impurities of crude zinc oxides make difficult their processing in primary zinc refineries due to mainly high chloride tenor. Washing with sodium carbonate is a costly operation and chloride removal efficiency is not good enough, so further purification steps are required. The crude zinc oxides are easily treated using ZINCEX™ technology based on leaching, solvent extraction, and electrowinning. The ZINCEX™ process is commercially available and is designed to be a perfect barrier for metallic impurities and halogens, allowing selective zinc extraction in mixed chloride and sulfate leaching solutions. Other valuable components as lead and silver can be efficiently recovered from the leaching residue using PLINT technology in chloride media. Silver is recovered as silver cement and lead is precipitated as pure lead oxide or carbonate, which can be smelted at low temperature.IntroductionZinc is a fully recyclable metal and at present approximately 70% of the zinc produced worldwide comes from primary zinc while 30% originates from secondary materials. Nowadays, the zinc recycling rate is continuously increasing mainly through galvanized steel scrap processing, and crude zinc oxides are the major source of secondary zinc [1]. Secondary zinc oxides are obtained from fuming processes like W aelz Kiln, Rotary Hearth Furnace or Primus Furnace. Waelz is the dominant technology accounting for 80% of total crude zinc oxides production worldwide. In the Waelz process, steel dust is mixed with solid fuel reductant (coke breeze) and pelletized. It is then fed to the kiln and heated to a reaction temperature of l ,200°C. In such conditions, zinc, lead, and other non-ferrous metals are vaporized into the flue gas together with chlorides and alkaline metals. Mixed oxides are drawn from the kiln with the process gases and separated to produce typical crude zinc oxides [2]."

Jan 1, 2011

By F. Russo-Bello

TauTona Mine (the renamed Western Deep Levels East) is using a longwall mining method with stability pillars to extract the Carbon Leader Reef at a depth ranging from 2097 metres to 3456 metres below surface. Seismic activity and the complexity and hazards associated with negotiating geological features at great depth pose the most severe hazard to deep level gold mining in South Africa. It impacts negatively on safety and the profitability of the mining. The critical success factors in longwalling at depth are: ?Sound mining strategies and face shape control ?Ensuring follow behind development infrastructure is kept up to date ?Support system design as a strategy to effectively limit the serious effects of damaging seismicity. This paper describes the following: ?Implementation of the required mining strategies ?The evolution of a new stope support system and problems associated with its development and implementation. How this support system was adapted for the successful negotiation of major geological features with known seismic potential and problems encountered ?Current initiatives to optimize follow behind development infrastructure ?Current initiatives in designing a new gully support system and problems experienced with its development and implementation. Finally, this paper will deal with the way forward, briefly covering areas like the above-mentioned gully support system, the proposed phased introduction of technology, and the use of alternative mining layouts.

Jan 1, 2000

By R. Samuels, R. Klimpel

"IntroductionLaboratory and industrial grinding tests have shown that the process of size reduction of metal containing ores can be significantly influenced by chemicals added to the slurry being ground.,The terms grinding aid or grinding additive refer to a substance which, when mixed into the mill contents, causes an increase in the rate of size reduction. The increased rate can be used to grind a higher feed rate to the desired product size or it can be used to produce a finer product size at a fixed rate. Whether the use of a grinding aid is justified in any given situation depends on the cost of the substance versus the improvement of output or product quality obtained with its use.A wide variety of grinding additive mechanisms have been suggested in the general mining literature (1,2,3,4,5,6,7). The prevention of particle agglomeration, slurry cushioning of grinding media impact, grinding media coating, changing the strength of macro porous rocks, and chemical adsorption on flaw surfaces to change bonding forces and surface energy have all been suggested. More recently, some grinding additive work has been carried out (8,9,10,11) which suggests that the use of efficient chemicals to control the fluidity of dense slurries is a viable economic approach to achieve improved grinding rates. These increased grinding rates are achieved with no increase in power to the grinding mills and apparently with no increase in media consumption costs. Unlike much of the previous work with grinding additives, this recent work is accompanied by a rather complete engineering characterization and prediction of performance capability. The chemicals involved were chosen so as to efficiently control the viscosities of slurries under conditions of high shear yet not adversely affecting flotation, flocculation, and other downstream surface chemistry operations."

Jan 1, 1979

By G. E. Walden

The Bureau of Mines is currently engaged in a continuing investigation of the thermodynamic properties of certain inorganic compounds. As a part of this investigation, the heat contents of bismuth and cerous chlorides were measured throughout the temperature ranges of 370° to 630° K. and 370° to 1,200° K., respectively. Equations representing heat contents as a function of temperature were derived from the data by the method of least squares. These equations were used to calculate heat contents at selected temperature intervals ranging from the ice point to 650° K. for bismuth chloride and to 1,200° K. for cerous chloride; also, the heats of fusion of both compounds were calculated at their respective melting points. EXPERIMENTAL Materials The cerous chloride used in this investigation was supplied by the National Carbon Company of Cleveland, Ohio. A chemical analysis indicated Ce 56.57 percent and Cl 43.39 percent. No attempt was made to purify the compound or identify impurities since the analysis showed only about 0.04 percent total impurities. Reagent grade bismuth chloride was used without additional purification.

Jan 1, 1961

By M MacMillan

THE following is a brief description of the method of hoisting ore automatically at the No. 14 shaft of Mcintyre Porcupine Mines, Limited, Schumacher, Ontario. All ore and waste below the 3875-ft. level is hoisted up No. 14 shaft from the 7000-ft. level to the 3000-ft. level. Seventy to seventy-five per cent of the mining is done in this area. The shaft consists of four compartments: a man way, a ventilation compartment and two skip compartments. The hoist, located on the 3500-ft. level (see Figure l), is an 80-in.-diameter, 4-rope friction winder, utilizing two 6-ton front-dump skips in balance. The hoist is powered by a 1,100-h.p. A.C. motor, the maxi-mum speed being 2,400 ft. per minute. The operation can be either manual, push-button automatic from the pocket, or fully automatic. Men can be raised or lowered in the skips, using a collapsible aluminum platform placed inside the skip box.

Jan 1, 1963

By L. Voisin, K. Itagaki

In the reductive smelting of the copper concentrate like a blast furnace operation using coke as fuel and reductant or in processing the copper concentrate which contains coal in the bulk, a sulfur-deficient matte phase often coexists with an iron base alloy in the bottom of the smelting furnace. This alloy is generally called furnace residue or speiss when considerable amount of arsenic or antimony is contained in the alloy and in the extreme case of reducing condition the alloy is saturated with carbon. As a fundamental study to know the behavior of some precious metals such as silver, gold and platinum between the sulfur-deficient matte and the furnace residue or the speiss which are in equilibrium under reducing condition, the phase relations and distribution of precious elements in the Cu-Fe-S-As and Cu-Fe-S-Sb systems saturated with carbon were determined at 1473 K by a quenching method.

Jan 1, 2007

By Dr. Steven G. Jansto

The market demand for reduced fuel consumption and CO2 emissions in the automotive, energy and construction sectors have further increased the demand for new and advanced higher quality microalloyed grades produced via TMCP. The technological integration of the process and physical metallurgical advancements of value-added niobium (Nb) microalloyed thermo-mechanical controlled process (TMCP) steels continue to be developed and is an important element for these more demanding end user requirements and cost-effective steel production. The reheat furnace process step has a profound effect on the TMCP performance, final hot rolled steel quality and mechanical property consistency during the production of hot rolled steels. The uniformity of heating applied across the entire width, thickness and length of the slab or billet is essential in the successful achievement of consistent customer properties regardless of the chemistry. The resultant ferrite grain size in the final hot rolled product is significantly governed by the initial prior austenite grain size. However, the transition from laboratory melted and TMCP hot rolled heats to the production scale is often quite challenging. Aberrations in the reheating process step at the industrial furnace are often the root cause of mechanical property quality issues. These combustion process metallurgy parameters connect directly to the resultant TMCP product quality and is defined as Combustion Metallurgy

Oct 1, 2019

By Don Dodds

Iron Mountain Mine Rehabilitation Project Brief History The Iron Mountain Mine is located in Northern California approximately 12 miles northwest of Redding. The ore body is a large, very dense pyritic intrusion, discovered in 1860, and initially minded for silver. The silver played out quickly and 34 years later the ore body was mined for copper. Copper mining was scheduled to cease in 1939, but WWII prolonged the closure. After WWII, the pyrite ore was mined to produce sulfuric acid. In 1949 when large amounts of cheaper sulfuric acid were being generated by the oil industry the underground mine shut down. Little was done with the site conditions until 1970 when the U.S. Bureau of Mines began monitoring the pH of the effluent. The initial pH readings were around 2.0. This was lower than anyone expected. Laboratory tests showed that when the ore from the mine was submerged in water, the pH of the water would stabilize above 4.0. However, the pH got steadily lower. In 1989 with the ph around 0.7, the U.S. Environmental Protection Agency (EPA) required the owner to build a capture and treatment facility. During this initial period, the majority of the effort was spent on the treatment plant. The relationship between the mine owner and the EPA was basically an adversarial relationship from the very beginning. In 2001, North Pacific Research was hired to upgrade the capture and control facility. The general layout of the Iron Mountain Mine is shown in Figure 1.

THE following brief paper is submitted with a view to encouraging the exploitation of hitherto neglected 10w~ grade ore-bodies which are suitable for extraction by open-cut methods.The Corinthian North mine, situated 10 miles from Southern Cross, in the Yilgarn district of Western Australia, furnishes an example of low working costs, although operations are carried out on quite a limited scale.The body of quartz which is now being worked strikes N.W. and S.E., has a very slight underlie to the west, and lies between bodies of quartz-hrematite s,chist on the east and amphibolite on the west.So far as the property has been explored, it would appear that the underlying sulphide rises to within about 70 ft. of the surface in the northern end of the workings, and pitches away to the south at a grade...

Jan 1, 1915

Power consumption of stirred media mills has been studied by an approach developed from work on stirred reactors. Relationships between torque required to rotate both disc and pin impellers immersed in dense particulate media as a function of speed, impeller and tank dimensions, and other relevant variables are evaluated. Data collected during numerous different experimental conditions are analyzed. The results can be summarized and presented in the form of characteristic plots, general equations and dimensionless group correlations involving power and modified Reynolds numbers. Special attention is directed towards the rheological properties of the media, which is considered as non-Newtonian liquids and hence, the equation of power law liquids is employed. Procedure for estimating parameters of effective viscosity and scale-up guidelines with respect to power consumption are put forward.

Jan 1, 1994

By O. R. Eckstrand, M. Asif, P. L. Schwann, A. J. Naldrett, R. J. F. Scoates

Paper presented at the IAGOD international symposium on mineralisation related to mafic and ultramafic rocks, held in Orleans, France, 1-3 September 1993. PGE-bearing sulphide mineralisation is contained in the lower 600 m of the Upper Central Layered Zone of the intrusion. There are four regimes with distinctive crystallisation characteristics. PGE content is low, both in bulk rock and in terms of metal in 100% sulphides, in the sulphide-rich cumulates but increases sharply in the sulphide-poor cumulates at the base of regime 1. Increases in regime 2 have no such clear relationship with incoming cumulus sulphides. PGE contents are low throughout regime 3, but regime 4 has sulphides with high PGE content, decreasing rapidly upwards, at the base of many of its cyclic units. The observations are interpreted as indicating that sulphide saturation occurred in the middle of a cyclic unit in regime 1, after which fractional segregation of sulphide caused a systematic decrease in PGE content. The resulting contrast between 40 000 ppb PGE in sulphides below the mineralised zone and less than 1 000 ppb above is of practical importance in exploration for such zones. In regime 2 the magma chamber was open to fresh influxes and PGE content was not so sensitive to sulphide saturation. In many regime 4 units influxes of new magma caused immediate sulphide saturation and PGE depletion, so that PGE-rich cumulates do not underlie PGE-bearing sulphide zones. Theoretical analysis indicates that regime 1-type behaviour is likely to characterise intrusions below the level at which plagioclase has entered as a cumulus phase and regime 4-type behaviour above

Apr 1, 1994

By William R. Yernberg

Most citizens take it for granted, but minerals produced in the mines of the western United States, including Colorado, are crucial to nation’s standard of living and quality of life, said Colorado’s Gov. Bill Owens in his welcoming remarks at the 106th National Western Mining Conference. The conference, which was organized by the Colorado Mining Association (CMA), was held Feb. 4-6, 2004, in Denver. Attendance was425, up substantially from last year’s meeting. The meeting included the CMA annual membership meeting as well as the CMA chairman’s and president’s reports. The meeting also included awards presentations in which companies were recognized for their efforts in pollution prevention, health and safety, and reclamation. The conference also included a tradeshow.

Jan 1, 2004

By K. C. Corin, Z. G. Song, J. G. Wiese

In the present study the effect of using different grinding media, viz, ceramic balls, forged steel balls, 15% Cr steel balls, stainless steel and mixture of stainless steel and 15% Cr steel balls (50:50) on the flotation performance of a Platinum Group Mineral (PGM) UG2 ore was investigated. Test work was conducted using a laboratory scale horizontal tumbling mill and a batch flotation cell. Power draw to achieve a specific P80 was determined for each media type. Changes in flotation behaviour were evaluated in terms of particle size distribution, water recovery, solid recovery and valuable mineral recovery. pH, Eh and DO levels were also measured. Grinding products of five different media produced similar solids and water recovery, which is indicative of the froth characteristics, but the grades and recoveries of PGE and chromite were different. The product obtained when using ceramic media showed the best flotation performance, producing the highest PGE recovery and the lowest Cr2O3 recovery. The products obtained when using stainless steel balls (SS) produced the lowest PGE recovery and highest Cr2O3 grade and recovery. The paper proposes possible reasons to explain these observations which may have significant importance for the operation of a PGM concentrator.

Jan 1, 2015

By Ignacy A. Lipiec

The advantage of microorganisms used in bioleach systems to conventional chemical reaction systems is that these apply the available energy more precisely where it is needed, either directly onto the mineral of interest or by efficiently catalyzing the production of a reactant. The literature cover the range from empirical models used in describing macro heat balance in heaps to detailed studies relating the micro heat effects of bacteria adhesion energy to the surface of minerals. In this paper, the heat effects are reviewed in a general way describing bioleaching in two macro systems, heap leaching and slurry bioreactors. The micro scale heat effects is the reactions and conditions at the microorganism scale, including the mechanism of leaching and bacterial behavior.

Jan 1, 1998

By P J. Fairfield

"This paper reflects on the author’s experiences in undertaking mining studies and reviewing mining projects and technical reports, particularly over the past five years, and the impact of recent changes in reporting standards and listing requirements pertaining to mining-related evaluation and valuation exercises. It also discusses contemporary changes in the expectations of the investment community and the implications for the role of the technical specialist in providing advice to such stakeholders.Not for the first time in history, the resources industry has experienced a significant downturn after a period of strong, sustained growth in line with the traditional view of the cyclical nature of commodity markets. The resource industry is currently faced with capital austerity in the wake of projects ‘failing’ (not delivering to expectations). In response, the industry has experienced changes in investment behaviour (slowing of investment) and an increasing regulatory burden as stakeholders seek a greater understanding of mining project risks.This paper discusses this response and uses case studies to highlight the key learnings (perhaps re-learnt) by the industry. It also discusses the challenges faced by technical specialists in communicating the results of technical and valuation studies to the investment community.As investors seek to take advantage of the current stage in the investment cycle, the key challenge of the technical specialist is to effectively communicate the uncertainty in technical assumptions, risks and opportunities that drive project value.CITATION:Fairfield, P J, 2016. Managing expectations – from buyer beware to buyer be aware, in Proceedings Project Evaluation 2016, pp 66–74 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By Adam Bezuijen, Tiago Gerheim Souza Dias

"The construction of mechanized tunnels in soft ground has evolved significantly over the last 20 years, especially in the control of the face pressure and the closure of the soil-lining void to reduce the induced settlements. On the other hand, several mechanisms of the TBM excavation cycle are still not taken into account for routine design calculations, such as the increment of water pressures in front of the tunnel face, the flow of excavation fluids around the shield, the dynamic equilibrium between the grout pressures and the excavation convergence, among others. This paper discusses specifically the issue of face pressures and how several mechanisms, which are routinely not considered, can be easily verified and incorporated into the state-of-practice of design. INTRODUCTION The stability of underground excavations is traditionally assessed through analytical solutions based on the lower bound theorem of plasticity (Atkinson and Potts 1977; Mühlhaus 1985) or the limit equilibrium method (Anagnostou and Kovári 1994; Messerli et al. 2010). These methodologies depend on the tunnel geometry, certain ground parameters and the support pressure. When conventional construction methods are used, the internal support pressure is an abstract representation of the combined effects of partial excavation, ground reinforcement, and the different support elements. The use of closed-face tunnel boring machines (TBM) changed this perspective, as the magnitude of pressures acting on the excavation boundary could be controlled in a more direct way. In front of the TBM the ground is excavated as the cutterhead rotates and the cutting tools scrape the ground from the tunnel face while additives are injected to condition the material. Water, polymers, bentonite, and foam can be used as additives under different conditions (Thewes et al. 2012). The loosened ground with additives, herein referred to as the mixture, flows through the openings at the cutterhead and into the excavation chamber. This mixture is kept pressurized to support the face."

Jan 1, 2016

By T. M. Drown, P. W. Shimer

The precipitation of phosphoric with titanic acid, by boiling an iron solution which had been reduced to the ferrous condition by sulpharetted hydrogen or sulphurous acid, was first noticed by E. H. Bogardus in 1874.* Since that time, as far as we are aware, not much has been published on the relation of these two acids to each other, and to silicic acid, in the ordinary course of analysis of iron ores. The following investigation may perhaps aid in clewing up some of the obscure points in the analysis of titaniferous ores. THE DETERMINATION OF PHOSPHORUS. From two to five grams of the finely powdered ore are weighed into a beaker and treated with about 50 c.c. of hydrochloric acid (sp. gr. 1.12), evaporated to dryness, and heated ih an air-bath for

Jan 1, 1882

By Albert W. Deurbrouck

This report describes the preparation characteristics of the significant, presently producing coalbeds in Mercer County, W. Va. Approximately 650,000 tons of coal was produced in Mercer County, W. Va., in 1960, of which over 630,000 tons was cleaned in four preparation plants. In 1960 the Pocahontas Nos. 3, 6, and 11 coalbeds supplied 99.5 percent of the total coal mined in the county. Because the number of working mines In this county was small, only six samples were collected; four of which were of metallurgical quality as mined. The remaining two samples could be upgraded by washing at suitable specific gravities. Many of the samples taken were of metallurgical quality as mined; however, the float-and-sink analysis on these coals has been included to point out their amenability to the production of premium quality, special purpose coals. In general the coals were low-volatile bituminous in rank with low ash and sulfur contents.

Jan 1, 1963