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By S. C. Rhoads

The Bureau of Mines is investigating low-chromium alloy substitutes for type 304 stainless steel (type 304 SS). The purpose of the research is to reduce the chromium content in austenitic stainless steels used for corrosion-resistant applications. Nine Fe-9Cr-12Ni alloys containing additions of 0 to 5 pct Mo, 0 to 2 pct V, and 0 to 2 pct Cu were selected for evaluation. All of the experimental alloys containing 5 pct Mo and an alloy containing 2Mo-2V-2Cu possessed structure and tensile strengths similar to those of type 304 SSe Alloys containing SMo-2V and SMo-2V-2Cu had ac-ceptable welding properties as measured by the Varestraint test. Alloys containing at least 2 pct Mo had cQrrosion resistance comparable to that of type 304 SS in solutions of nitric acid, phosphoric acid, acetic acid, and citric acid. In chlorides, all the experimental alloys were susceptible to localized corrosion and stress corrosion cracking similar to that of type 304 SSe The corrosion properties were established by weight-loss tests, electrochemical polarization tests, and stress-corrosion cracking tests. An Fe-9Cr-12Ni-SMo-2V alloy and an Fe-9Cr-12Ni-SMo-2V-2Cu alloy had a combination of welding properties, tensile properties, and corrosion resistances that were most comparable to those of 304 SS.

Jan 1, 1984

By T Blyth, D Holden, N Archibald, S Huffadi

The Morning Star mine is owned by Hill 50 Gold NL. Gold is hosted in quartz-carbonate ¦stibnite ¦molybdenite veins that occur in a series of deformed, altered and metamorphosed mafic volcanics with subordinate sediments and felsic intrusive rocks. The irregular gold distribution and its relationship to structure and alteration intensity makes the definition of stoping limits difficult. Understanding the structural evolution of the orebodies and the contemporaneous and later structures that affect the orebodies has been very useful in this context. This paper reports on a detailed mapping exercise (as part of a more extensive and on-going investigation) and the integration of this mapping with other geological and grade control data in a 3D modelling environment to provide an expanded basis for stope design. Three types of mineralised zones can be discriminated based on the morphology of the veins within. They are:those associated with strongly folded, thick veins (approximately 1 m thick eg 05 orebody); those comprising planar, boudinaged and some folded veins which consolidate to form tabular, steeply dipping bodies (eg 010 orebody); and zones containing the above vein types and thin, highly deformed veins within zones of intensely foliated country rock that result in localised, irregularly distributed high-grade gold zones (eg 02 orebody). The mineralised zones associated with thick folded veins are typically predictable over many tens of vertical metres. The second and third types of ore zone is generally less predictable. Detailed mapping and structural kinematic studies indicate all ore zone types are affected by progressive deformation involving strong strain partitioning. It is interpreted that high strain zones, recognisable by intense foliation development and highly deformed veins, are the latest stage of a progressive deformation which approximates bulk flattening. Structural controls on mineralisation can be defined at two different scales.Regional scale folding has deformed a sequence of mafic volcanics with interlayered sediments and felsic intrusives. Initial geometry of felsic bodies relative to orientation of the primary lithological layering and the bulk shortening direction is interpreted to be critical for initiation and development of strongly dilatant zones within this package of rocks, and Local scale dilation is evidenced by development of several vein sets whose range in orientations and opposed vergence relationships suggest a pre-folding origin comprising both extensional and hybrid shear-extensional types. Continued deformation resulted in folding of these veins with strong strain partitioning. Major vein packages are closely spatially associated with the felsic/mafic contact and an interpreted early structure which appears to have partitioned early dilation. This structure is now obscured by strong deformation but is currently represented by a zone of planar veins.

Jan 1, 2000

By C. E. Jordan

The Bureau of Mines is conducting research into the pneumatic recovery of coarse mica and has used this method to produce mica concentrates that contain more than 90 percent mica. This research is being carried out pursuant to the Bureau's objective to develop technology that will help maintain an adequate supply of minerals and metals to meet national economic and strategic needs. Researchers used a Bureau-designed system of crushers, screens, and zig-zag air classifiers to concentrate coarse liberated mica particles from mica-bearing materials. This pneumatic system was used to concentrate four mica ores from Arizona, North Carolina, and South Dakota and three waste tailings from Alabama, Georgia, and South Dakota. Using these samples, it was demonstrated that plus 65-mesh size mica can be effectively recovered by the pneumatic method. Not only were the concentrates high in mica content; it was also demonstrated that this method can be used to recover up to 78 percent of the mica that was originally contained in the samples. Because it is a dry concentration method, the pneumatic beneficiation technique may be advantageous in areas where water resources are limited.

Jan 1, 1980

By Emil B. Milosavljevic

Problems associated with distillation and other classical methods for analyzing operationally defined cyanide (CATC, WAD, Total and Free Cyanide) will be discussed. These methods often achieve incomplete, species dependent cyanide recoveries and suffer from serious interferences. Depending on the relative concentrations of interferents and other matrix constituents, underestimation (ultimately Type II errors probability of not detecting a constituent when it is actually present) or overestimation (ultimately Type I errors -probability of deciding a constituent is present when it is actually absent) of the cyanide levels present in a ,given sample may occur, causing in the first case false security and in the second case, excessive cyanide treatment costs. Mahy, if not all, enumerated problems can be solved by using novel flow injection (FI) gas-diffusion amperometric methods. A modification of the recently approved FI ligand exchange (FIILE) method for determination of available cyanide (USEPA Method OIA-'l677) that obviates sulfide interference will be discussed. Eight laboratory round-robin data obtained for the FI UV -digestion (FIIUV) method that determines total cyanide (Proposed Method OIA-1678) will also be presented. In addition, some methods for cyanide speciation will be presented. Finally, research needed for developing reliable methods for determination of operationally defined cyanide in solid samples will be summarized.

Jan 1, 1998

By H. S. Price, F. N. Kissell, R. C. McCulloch, J. C. Edwards

"The methane em1ss10n from deep coalbeds which are actively mined is a serious problem facing the mining industry today. Pinpointing the source of methane and predicting the emission of this gas from these coalbeds is necessary if preventive techniques are to be effectively used to control methane emission. In general, because most coalbeds have water associated with them, the system to be studied may be either single-phase (methane gas) or two-phase (methane and water). Because coal mine geometry is complex and the fluids have variable properties as functions of pressure and methane concentration, the differential equations describing the flow of methane and water are not amenable to analytic solution, but must be solved numerically using high-speed computers. This paper presents the development of such a complex mathematical model along with some preliminary applications of the model to methane emission from one- and two-dimensional coalbeds with and without water present."

Jan 1, 1973

By D J. Brake

Most mine ventilation engineers are involved in ventilation planning and design in some capacity. Ventilation modelling software used by a competent experienced ventilation engineer is extremely useful in developing good ventilation designs by allowing assessment of a wide range of potential options. Unfortunately, ventilation modelling undertaken by persons who do not have sufficient experience or knowledge usually results in a façade that covers up a fundamentally unsatisfactory design. Ventilation models are produced for a wide variety of purposes including: fault-finding of a problem area in a mine and options analysis for resolution of such problems; a complete review or optimisation exercise of an entire mine’s ventilation system; or a much longer life-of-mine type of study. Models can be intended as the basis for studies with diverse purposes such as primary shaft or airway sizing, primary fan specification, examining re-entry times after blasting, resolving leakage or recirculation, or investigating the impact of fires or underground climate or cooling requirements. The concepts of ‘materiality’ and ‘fitness for purpose’ are essential to developing or using a ventilation model and serious mistakes have been made in wrongly using a ventilation model for purposes for which it was never intended, often because it is simply the most recent model on the mine site. Mistakes at this level often translate into faulty ventilation strategies and inefficient or ineffective use of scarce capital for ventilation projects. Validating a ventilation model is a time-consuming and expensive process and not every model must (or even should) be fully validated to meet the objectives at that time. This paper discusses the application of quality assurance in ventilation planning with particular respect to the ‘basis of design’ (BOD) as well as the standards for validating a ventilation model. It also provides a recommended way of dealing with non-conformances in measured versus modelled values of critical parameters in the model.CITATION:Brake, D J, 2015. Quality assurance standards for mine ventilation models and ventilation planning, in Proceedings The Australian Mine Ventilation Conference, pp 221–228 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

By R. L. Grayson, S. S. Peng

"Abstract. Sampling of respirable dust and analyses of samples have been partially completed for a longwall panel. Physical and compositional properties which have been characterized include:l. Respirable mass concentration, 2. Mineral composition,3. Particle size distribution by mineral,4. Particle shape distribution by mineral,5. Particle angularity distribution by mineral, and6. Respirable size distribution.Analytical techniques used in the characterization of samples are described, and results are reported with emphasis placed on relationships between the various characteristics and worker locations. Details of existing mining conditions, equipment arrangements, dust control methods and operating system parameters are presented.INTRODUCTIONThrough the work of many researchers worldwide, the health of coal miners in modern societies is now protected from dangerous exposure to respirable coal mine dust. A vast amount of research culminated in the exposure-response model of Jacobsen, et al. (1971) whi3h has been used to establish the 2mg/m exposure limit in the United States. According to the model, exposure to respirable dust for 35 years at this level or less would prevent the occurrence of a clinically significant case of coal workers' pneumoconiosis (CWP). The exposure limit, of course, must be adjusted whenever greater than five percent free silica exists in the respirable dust.The Jacobsen model was built around a mass concentration exposure, which showed the highest correlation with incidence and progression of CWP. As Morgan; et al. (1973) noted, howev.er, the quantity of respirable dust inhaled is not the only significant factor. It is noted that although mass of respirable dust correlates highest with the incidence of CWP, it is not a causal factor. Indeed, the insidious processes causing the disease occur on the cell-particle scale, requiring microscopic investigation."

Mar 1, 1989

By J. P. T. Kapusta

"The copper smelting industry faces increasingly stringent environmental regulations and must continue driving innovation to reduce energy usage in its production chain. With the coming of age of ‘bottom-blowing’ copper technologies in China over the last decade, much renewed attention has been brought to tuyere bath smelting for clean copper-making. Since bottom-blowing sonic injection was originally developed and implemented in steelmaking in the 1970s and lead smelting in the 1990s, the time seems appropriate to revisit the sonic injection concepts and dispel some myths. After reviewing the trends in the copper industry and providing a brief historical context of research and development related to sonic injection, the author presents a comparison of the main submerged tuyere bath smelting technologies, namely the Noranda Reactor (NR), the Teniente Converter (TC), and the Chinese Bottom-Blowing Smelting Furnace (SKS/BBS) for copper, and the Queneau-Schuhmann-Lurgi Reactor (QSL) for lead. The core of this article, however, consists of a brief description of the fundamental theories for sonic injection (jetting regime) and ‘sonic flow mapping’ to illustrate the relationship between the minimum tuyere back-pressure required to achieve sonic flow and the tuyere internal diameter. The article also offers a comparison between side- and bottom-blowing in jetting regime before concluding with the author’s vision of the new trend in submerged tuyere bath smelting. IntroductionIn the last decade, developments in the copper smelting industry have been driven by two key concepts: (1) increasing cleaner smelting capacity to meet the world’s growing copper demand, as illustrated in Figure 1, and (2) increasing process off-gas capture to comply with more stringent environmental regulations and abide by a corporate social responsibility ethic of a clean work environment (for smelter employees) and clean air (for families living around smelters). The copper smelting industry has also faced some major challenges in the last few decades, including (1) low treatment and refining charges (TC/RC or processing fees) impacting custom smelters, (2) lower concentrate quality and grades with higher levels of impurities (e.g. As and Hg), (3) higher processing and transportation costs due to the rising cost of energy, (4) restrictions in process intensification (tonnage oxygen usage) due to furnace integrity limits, and (5) scarcity of educated and trained personnel due to fewer metallurgists graduating from pyrometallurgical school programmes combined with a high retirement rate of skilled workers."

Nov 1, 2018

By A L. Zawadski, B G. Putland

"This paper discusses a method to build a calibrated comminution circuit model, using historic operating data and blend information that is useful for forecasting and optimising plant performance, as well as evaluating expansion opportunities. A preferred method for building a calibrated power-based comminution model is discussed and justified based on two case studies, offering low-cost and effective strategies to utilise existing information for the maximum possible benefit.The first example is the Lefa Gold Mine located in Guinea, West Africa operating a semi-autogenous grinding (SAG) and ball milling circuit with recycle crushing (SABC). Daily blend information and operating data was utilised to build and calibrate a power-based model of the comminution circuit. The least squares regression method was used to solve the SAG mill specific energy requirement and the operating work index (Wio) for each individual ore type. The calibrated model was used to evaluate expansion strategies to sustain throughput at higher proportions of fresh ore based on the proposed life-of-mine schedule.The second example is the Ok Tedi Copper-Gold Mine in Papua New Guinea operating a SAG and ball milling circuit (SAB). The circuit limitation would often shift between the SAG mill and the ball mill due to variable ore characteristics and feed blend. Typical SAG mill relief strategies, such as recycle crushing or secondary crushing, required significant capital expenditure and the potential throughput benefit would vary depending on ore type, making economic justification difficult. Daily blend information and operating data was used to build and calibrate a power-based model of the comminution circuit to identify ore types that might benefit from higher blast intensities to increase fines content, thus relieving the SAG mill limitation and increasing overall throughput without a major capital investment.Citation:Zawadski, A L and Putland, B G, 2016. Calibrating Comminution Models Using Operating Data to Forecast and Optimise Plant Performance, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 17–26 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By G. T. Fisher

The U.S. Bureau of Mines conducted research on calcium ferrite slags to determine the distribution of Pb and minor elements (Ag, Cd, Zn, and Co) between Pb bullion and slag. These fundamental data are required to evaluate nonferrous smelting with calcium ferrite slags as a substitute for conventional iron silicate slags. Determinations were made in 25-cm3-capacity MgO crucibles at 1,250° C for 2 h in a CO-CO2 atmosphere with an O2 potential that would retard the formation of magnetite. Calcium ferrite slag composition variables studied were from 16 to 32 wt pct CaO in 2-pct increments and Si02 ranging from 1 to 5 wt pct in 1-pct increments with equal additions of MgO. For all compositions studied, the retention of Pb in the calcium ferrite slag was less than 1 wt pct, which compares favorably to about 6 wt pct Pb in conventional iron silicate slags used in smelting of Pb concentrates. Minor element distribution was insensitive to changes in slag composition in the range studied. However, slag composition had significant impact on the melting and fluidity of the slag.

Jan 1, 2010

By D. H. Rubisov, V. G. Papangelakis

Sulphuric acid pressure leaching is the process of choice to recover nickel and cobalt from laterites with variable iron and magnesium contents. The latter influences the acidity levels required for effective leaching of nickel and cobalt, and during leaching, iron is rejected in-situ as hematite. Classical solid/liquid separation is performed via a CCD circuit, the performance of which depends upon the size and shape of precipitated particles. In this paper, precipitates obtained by acid-leaching of various lateritic feeds at 270°C are characterized using SEM/EDX and laser particle size analysis. It is shown that as acidity increases, the shape of the precipitates gradually changes from platelets to spheres, thus improving the S/L separation properties of the residue. The use of hydrogen ion concentration “at temperature,” calculated from a speciation analysis, allows the explanation of this change of morphology independently of the type of the ore feed. The possibility of precipitating magnesium sulphate “at temperature” is also discussed. In order to take advantage of this precipitation for scale reduction (as has been proposed in the past) without compromising the residue quality, [H+] “at temperature” must be controlled within an optimum range.

Jan 1, 2000

By L. P. Kleinmann

To control the formation of acid drainage at its source, the Bureau of Mines has investigated the role of iron-oxidizing bacteria in the oxidation of pyrite. In laboratory tests, it was determined that Thio-bacillus ferrooxidans could be controlled in coal refuse using anionic surfactants, with an associated decrease in acid production of about 75 pet. Based on these results, large-scale tests were conducted at an 8-acre active coal refuse area in northern West Virginia and an 11-acre inactive refuse pile in southern West Virginia. Acid production was decreased 60 to 95 pet for 4 to 5 months after application of 55 gal/ acre of 30-pct sodium lauryl sulfate solution. To date, effluent surfactant concentrations have been extremely low. Either a hydroseeder or a road-watering truck can be used to apply the surfactant. Since the surfactant solution must reach the pyritic material to be effective, sites that have been reclaimed are generally not amenable to these procedures. A simple laboratory test is described that provides an estimate of adsorption potential; this can be used to calculate a safe application rate.

Jan 1, 1983

By Mahesh C. Jha

Cyanidation has established itself as the main process for extracting gold from ores in this century. However, many ores exhibit refractoriness to the conventional cyanidation process. The main cause of the refractoriness is generally the presence of carbon, tellurium, and sulfides, particularly those of antimony, arsenic, and iron. The problem is usually solved by either modifying the cyanidation conditions or providing an oxidation pretreatment. This paper presents results from studies that help in understanding the causes of refractoriness of various gold ores. This is followed by a discussion of many processes that have been developed to solve the problem such as intensive cyanidation, pressure cyanidation, carbon-in-leach cyanidation, and pretreatment processes such as roasting, chemical oxidation, autoclave oxidation, and bacterial oxidation.

Jan 1, 1986

By N. Agron-Olshina

The application of the QEM*SEM image acquisition and analysis system to the measurement of the mineral matter in coal and the materials in fly ash is described. A new sample mounting technique was developed using iodinated epoxy to provide a backscattered electron contrast between the coal, the mounting medium and the mineral matter. Several comprehensive and detailed species identification programs for silicate minerals and materials were developed to analyse the 1000 count x-ray spectra used for the measurements. Case studies are presented to demonstrate the capabilities of the new technique to distinguish different coals and fly ashes by their mineral and material composition rather than by conventional chemical analysis. Applications of the new technique are aimed at demonstrating that composition and size distributions obtained from OEM*SEM measurements can be used: 1. to characterize coal with respect to combustion problems in power stations, 2. to characterize coal for fine coal cleaning, 3. to characterize fly ashes in relation to problems such as erosion in combustion systems and the ease of precipitability in dust collection facilities, 4. to refine the classification systems used in marketing coal for pulverized fuel firing and fly ashes as a component of cement and building materials.

Jan 1, 1992

By Roorda H. J, R. R. Saddington, Curlook W

THE copper-nickel sulfide ore deposits of the International Nickel Company of Canada, Limited contain the minerals pyrrhotite, pentlandite, and chalcopyrite. In the extraction of the nickel, copper, and iron, much of the processing involves oxidation of the metals as well as sulfur. Roasting plays a predominant role at several stages in the operations which prepare the constituent metals for refining. Copper, nickel and iron concentrates of various grades, ranging from low-grade nickeliferous pyrrhotite to high-grade nickel sulfide, must be partially or completely oxidized. Figure 1 shows the main points at which roasting is conducted at INCO. Roasting experience at INCO dates back to the time when lump ore was partially roasted in heaps. It has progressed through single hearth roasters, sinter machines, multiple-hearth roasters, flash roasters, and more recently has centered on the development and utilization of fluid bed roasters. The primitive, slow, and inefficient method of heap

Jan 1, 1961

Pillars are used in various forms in underground mining. For several decades pillar design has been an issue in both soft and hard rock underground mining. The major use of pillars is in room-and-pillar type mining. Pillars are critical structures in the production levels of block cave mines. With most mines leaning towards the use of block caving for ore extraction today, pillar design has become even more important. Empirical pillar design is popular. The use of numerical modeling is becoming more common in the design of underground mine pillars. A critical review of both the empirical and numerical modeling approaches to the design of underground mine pillars reveal some fundamental flaws. In general, pillar design approaches often assume the pillars are vertical, and are subjected to pure compression. Pillar failure criteria are then based on this assumption. Most papers on numerical modeling for pillar stability assessment use vertical pillars. This paper shows that where pillars are not subjected to pure compression the empirical pillar design approaches are inadequate, and that when using numerical models for the assessment of pillar performance, more care is needed in the determination of the average pillar stresses. The practice of relating field scale pillar strength to laboratory scale intact rock strength determined from conventional laboratory compression tests in empirical design equations is misleading when pillars are inclined or subjected to compressive and shear stresses. New data is added to the hard rock empirical pillar stability chart, and it is shown that it is misleading to conclude that failure of pillars with width to height ratios greater than 2.3 should not be anticipated. New empirical pillar design charts are presented for pillars at various degrees of inclination for different ko-ratios.

Aug 1, 2013

By B Kilgariff

Flotation has been in use at Broken Hill for nearly 100 years, and the process has gone through many different optimisation studies over that time. This paper reviews the current flotation practice at the Perilya Broken Hill Mine, and details one of the recent optimisation studies conducted using mathematical modelling and simulation tools. The optimisation study consisted of surveys of the lead and zinc circuits, mapping of the floatability components of the ore and developing models of the lead and zinc circuits. The models were then used to conduct simulations to simplify the circuit. The study allowed plant personnel to identify optimum operating conditions for their current flotation cells, as well as possible improvements in circuit configuration and cell capacity.

Jan 1, 2005

By R. H. Dickson

This paper, which describes the milling practice of the Verde Central Mines, Inc., is one of a series on milling methods and costs being prepared by the United States Bureau of Mines. GENERAL DESCRIPTION The Verde Central concentrator is located in Walnut Gulch, about 1 mile west of Jerome, Ariz., and on the Jerome-Prescott highway. It is about 1,000 feet distant from the main shaft of the Verde Central mine and treats ore from this mine only. The mill is connected with the main shaft of the mine by a 24-inch gave railway, over which the ore is hauled in 5-cubic yard capacity, rocker, side-dump cars, by a 3-ton Brookville gasoline locomotive.

Jan 1, 1931

By J. M. Stuve

Low-temperature heat capacities and the enthalpy of formation were deter-mined calorimetrically by the Bureau of Mines for CuO?CuS04. The derived standard entropy (S°298.15) was 40.36±0.04 cal/deg-mole, and the enthalpy of formation (?H°f) was -219.l±0.3 kcal/g-mole. A large thermal anomaly was observed in the heat capacity of CuO?CuS04 at 20±0.1 K. Gibbs energy of formation data are tabulated up to 1,200 K. Construction details of a precision low-temperature adiabatic calorimeter capable of determining heat capacities of solids in the temperature range 5 to 320 K are given. A calibrated miniature germanium resistance thermometer was incorporated into the calorimeter to extend the range of accurate measurements below 15 K. High-resistance dc heater windings were used throughout the cryostat to reduce heat conduction of inner connecting wires.

Jan 1, 1975

By D. E. G. Maschmeyer

The operation of the pilot plant for the S-C Copper Process during 1976 has provided the basic data for the design of a commercial-scale plant. This paper introduces the commercial aspects of the S-C Copper Process through the presentation of a process flow diagram, a schematic flow diagram and basic design parameters for 75,000 tonnes per year (tpy) copper rod production from chalcopyrite concentrate. Basic operating requirements are outlined and a preliminary estimate of operating and capital costs for a Western Canadian location are summarized. The results have been compared with a modern smelter/electrorefinery to highlight the advantages and disadvantages of these two alternatives and to show why the new S-C Copper Process technology should be of interest to the copper industry.

Jan 1, 1978