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By Sabine Verryn, Johan de Villiers

"Through the application of the Rietveld method, X-ray diffraction analysis has developed into a fully quantitative method for phase analysis. The Rietveld method itself has progressed from an unstable method that required constant operator intervention to a stable and robust method suited even for routine analysis. The method however, requires a good initial calibration and periodic checks on its precision and accuracy.Several examples from minerals processing, pyrometallurgy, and materials science show the strengths and applicability of the method. The limitations and sources of error are also discussed.1. IntroductionDiffraction of X-rays by matter has been used for more than 80 years, initially to elucidate the structure of crystalline matter, and to investigate the bonding between atoms. The first development of powder diffraction cameras and the later development of X-ray diffractometers saw the development of methods to identify phases and to try to quantify them in mixtures. A comprehensive database and search-match techniques have been developed over the ensuing years. Effective and accurate quantification had to wait for the so-called Rietveld' analysis where the total diffraction pattern is used and where the crystal structure data for the constituent phases are needed as input. Overlapping of diffraction peaks of different phases is handled without difficulty and reliable results can be obtained in spite of this occurrence.The development of methods for structure analysis from powder data has also progressed rapidly and more and more complex structures are being solved from first principles, i.e. without knowledge of unit cell or space group data. Because of the ease of use of the Rietveld method, it is finding increased application in metallurgy, especially in the quantification of phases of importance in minerals processing, in pyrometallurgical processes, and in materials characterization. This study will examine the principles of the Rietveld method, its application in metallurgy, and finally also the common pitfalls that need to be avoided.2. Rietve1d AnalysisAs the Rietveld method compares the experimental diffraction pattern with one calculated from known crystal structure data, the following data are needed:•,An XRD Pattern with as Iowa background as possible. This should have good resolution and high diffraction peak intensities. With modem X-ray detectors, this can be achieved in minutes rather than hours.•,Instrumental parameters such as goniometer radii and slit sizes are needed. These are the so-called ""Fundamental parameters"". They are used to calculate the shapes of the resultant diffraction peak profiles that are caused by the instrument. The parameters are used to correctly separate the instrumental effects on the diffraction pattern from the sample effects."

Jan 1, 2014

By R. F. Favreau, P. Favreau

Rock excavation is the first process in the cycle of operation of a mine, and efficient blasting is paramount to the profitability of the mine. The traditional way to design blasts has been by trial and error, based on the experience of the designer. This is probably the only important engineering practice which continues to be done without a precise engineering technique. The efficiency of most other mine processes has been improved by the use of high-tech, while a majority of blast design has not yet profited from the advantages of the computer. However, with the advent of the high-tech blast simulator Blaspa, it has become possible to design the parameters of the blast according to normal engineering practices. Over the last 50 years, the chemical and physical mechanisms which occur during a blast have come to be identified; they are reviewed in section B. Although these mechanisms are complex, they have nevertheless been expressed in terms of mathematical equations which have been assembled together in the computer to create the blast model Blaspa.

Jan 1, 2002

By Wolfgang G. Brunner

The closing of the gap in the European waterway- net is one of the big and challenging traffic projects following the German reunification. The water canals in the North German plain connect the Rhine river, the border between Germany and France, with the Oder, the German border river to Poland. As early as in the 30's of the 19th century the planning and construction of a ship- lift and a canal bridge started, in order to close this last big gap in the system between the Mittelland Canal and the Elbe- Havel Canal near Magdeburg. This main waterway connects the West of Germany, known for its coal mines and steel mills, with the capital Berlin, and from there it goes on into Poland. The building activities started in 1934 but came to a complete stop in 1942 because of the war. The communist government in East Germany was not interested in continuing this huge project, including the ship lock Rothensee, the canal bridge crossing the river Elbe, the former border river between West and East Germany, and the lock Hohenwarthe due to lack of political backing along with low construction capacity.

Jan 1, 2000

By Lawrence Adler

In-situ rock has well-recognized rupture and yielding responses. Attention is called to an intermediate response, granulation, which consists of an evenly distributed, small-scale crumbling accompanied by swelling. Granulation results in a negative slope for the inelastic portion of the stress-strain curve as successive interlocks are destroyed, and the load-carrying capacity of the material is reduced. Volumetric swelling with softening also takes place. The prerequisites for such a response are moderate confinement and a uniformly distributed defect pattern. Since rock masses, as well as rock as an intact material, have such patterns in their jointing and texture, respectively, granulation is characteristic of them. The presence of granulation can be demonstrated for a very wide scale, ranging from laboratory testing to regional geologic phenomena. The response of an engineered structure is considered in detail. Granulation appears to be dependent upon internal moments, the formation of which can be controlled by engineering design. The practical implications of such control can be significant in optimizing underground instability and in improving excavation techniques. Rock is usually considered to behave inelastically under high load, by fracturing as a brittle material, or by yielding as a ductile material, ignoring time dependency. These responses, which are characteristic of metallic behavior, appear idealized in Fig. 1, where linear-elasticity approximates the initial behavior. However, another major deformation mechanism peculiar to rock has long been recognized in geology, the granulation of an interlocked material. Granulation consists of a uniformly distributed fracturing and crumbling of the material, accompanied by swelling and softening. Predictably, the associated inelastic portion of the load-deformation curve has a negative slope,

Jan 1, 1971

By W. D. Robertson

SINCE the comprehensive paper of Moore, Beckin-sale, and Mallinson,' little consideration has been given to the mechanism of mercury stress cracking of copper-base alloys, apart from extensive work on metallurgical and fabrication variables employing mercurous nitrate as an evaluation test. However, with the phase diagrams now available and the recent considerations of interfacial tension at grain boundaries, developed by Smith,2 it appears that the elements of a detailed solution to the problem are available. The facts that require explanation are: 1—Pure copper is not embrittled by mercury, with or without an applied or residual stress. 2—Cu-Zn and other copper-base alloys, in excess of some undefined solute concentration, are embrittled provided an external or residual stress is acting; cracking is not observed in the absence of stress. 3—The path of fracture is invariably intergranu-lar. 4—Copper and copper-base alloys do not dissolve readily in mercury at room temperature.' Consideration of the Cu-Hg phase diagram% hows that, at room temperature, solution of copper in mercury is probably limited by the formation on the surface of a microscopically thin layer of inter-metallic compound, CuHg, having a very limited range of solubility which effectively restricts diffusion of copper and mercury. But, above 150°C (the highest peritectic temperature) copper solid solution is in equilibrium with a liquid Hg-Cu solution and, consequently, copper should dissolve freely in a large volume of mercury with the formation of an equilibrium dihedral angle at grain boundaries in contact with the liquid phase. Depending on the magnitude of this angle, mercury will or will not penetrate the solid phase along grain boundaries and spread over grain faces. Thus the measurement of the dihedral angle of grain boundaries in equilibrium with the liquid phase above 150°C should indicate whether copper is intrinsically resistant to embrittlement by mercury. In view of the fact that mercury is appreciably soluble in copper, it may also be anticipated that diffusion of mercury will take place at higher temperatures in the absence of the intermetallic compound and possibly at a more rapid rate along grain boundaries than through the grains, resulting in a Cu-Hg alloy of unknown properties. In the case of brass, the ternary system Cu-Zn-Hg is involved and, from the resistance of brass to general solution in mercury at room temperature, it is probable that a compound of limited solubility range also exists, similar to the binary Cu-Hg sys- tem. Also, compared with copper, a change in the dihedral angle at grain boundaries in contact with mercury may be expected to result from the addition of zinc to copper. The preceding generalizations appear to be confirmed by the following experiments. Experimental Results Oxygen-free copper wire, annealed at 700°C for 3 hr in copper foil, was immersed in mercury at 170°C, with and without an applied stress. After 24 hr immersion, specimens which were spring loaded to a nominal stress of 10,000 psi had not failed, and there was no apparent embrittlement as indicated by a 180" bend around a diameter approximately equal to the diameter of the wire (0.125 in.) : General solution of the surface had taken place, indicated by a change in diameter of the wire, and the resulting structure of the surface is shown in Fig. 1. It is apparent that angles have developed at grain boundaries and examination at a high magnification indicates that the angle is in excess of 120°, completely excluding penetration of mercury. In view of the measurements made by Smith n the

Jan 1, 1952

By David Schlain, C. B. Kenahan, F. X. McCawley

The electrodeposition of protective coatings of tungsten from a sodium and lithium borate - tungstate - tungstic oxide fused salt electrolyte was investigated . Soluble tungsten anodes were employed . Deposits were made on molybdenum , Carpenter 20 stainless steel , Inconel , and nickel . An atmosphere of argon was maintained around the electrolytic cell , and the electrolyte was stirred by rotating the cathode . Thick deposits of tungsten (up to 26 mils ) were made at a temperature of 900 ° C and cathode current densities of 0.3 to 0.4 ampere per square inch with high current efficiencies . The effects of variations in temperature , current density , electrolyte agitation , and tungstic oxide content of the electrolyte and of the presence of impurities such as moisture , air , iron , and nickel were investigated .

Jan 1, 1964

By S M. Iveson, K P. Galvin, D Kumar

Dry separation of iron ore fines and other heavy minerals has previously been achieved using one of two main approaches. The first is an inclined table with an upward airflow via small holes to help form an autogenous dense medium. The second is based on fluidised beds comprising a dense medium formed by fine particles suspended in an upward flow of air. In either case, vibration can be introduced. This paper describes recent findings using a novel system referred to as the Sink-Hole Fluidiser, which utilises elements of both these two former approaches. This device consists of a conventional gas-fluidised bed of fine medium with a 1 mm aperture mesh located at the surface. This mesh contains relatively large holes that act as sink-holes. The fluidised media flows upwards through the sink-holes and then falls back into the fluidised bed via the fine apertures in the mesh. The particles being separated, which are much larger than the fine media, pass over these sink-holes, where they either sink or float. The sink-holes offer a more precise balance of forces, applied on a single particle basis, resulting in robust density separation.Experiments have been performed using 2.0–4.0 mm tracer particles of density varying from 2.4 to 4.8 relative densities (RD) separated in a 200 mm diameter Sink-Hole Fluidiser with a single 13 mm diameter centrally-located sink-hole. In this preliminary semi-batch work, it was found that the Sink-Hole Fluidiser was able to achieve sharp separations (Ep as low as 0.20). This work underpins the development of a new continuous steady-state Sink-Hole Fluidiser, applicable to the dry separation of iron ore fines.CITATION:Kumar, D, Iveson, S M and Galvin, K P, 2017. A novel sink-hole fluidisation method for dry separation of iron ore fines, in Proceedings Iron Ore 2017, pp 221–226 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By D O’Dea, B Godel, T Honeyands, B Ellis, T Harvey

X-ray computed tomography (XCT) is a non-destructive technique that allows the exploration of the internal structure of solid objects in 3D. Over the past ten years, the combination of XCT with more conventional 2D techniques and quantitative 3D image processing and analysis (commonly referred to as digital rocks) has allowed us to address some of the most challenging problems in ore deposit studies. The range of X-ray energies used in XCT allows the X-rays to be transmitted through complex and dense materials including iron ores (eg sinters, lumps, granules). Recent high-resolution XCT scanners have the ability to generate multiscale (25 down to 0.3 µm voxel size) imaging of a given sample enabling the characterisation of 3D textures across a range of scales. This paper will focus on the application of digital rocks to iron ore sinter particles (using filtered polychromatic X-ray beam). The paper is divided into three parts and includes: i) details on the methodologies including their benefits and limitations and the effect of spatial resolution; (ii) the quantification of pore network topology and pore characteristics (eg open versus closed pores, pore size distribution and pore morphologies; degree of connectivity of the pore network and its tortuosity); and (iii) the evaluation of the technique to characterise various phases within the material and quantify 3D mineral textures.CITATION:Godel, B, Ellis, B, O’Dea, D, Honeyands, T and Harvey, T, 2017. Digital rocks for iron ore sinters – towards a D quantification of sinter textures, in Proceedings Iron Ore 2017, pp 529–534 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By Lu Lin, Yan-ping Bao, Min Wang

In order to forecast the cost of dephosphorization during BOF process, and achieve economical dephosphorization of BOF. The mathematical model of dephosphorization cost for BOF process was established. The model calculations show that, when the silicon content and phosphorus content are increased in hot metal, and the final phosphorus content of BOF is decreased, the dephosphorization cost will be increased. The silicon content of hot metal is increased of 0.1% in the plant, the dephosphorization cost is added by 9 RMB/ton steel. The De-P cost of mode A is much higher than the other two modes for three typical production modes of single slag melting process, when the phosphorus content of hot metal is higher than 0.14%, an increase in the phosphorus content by 0.01%, the De-P cost of mode A is added by about 7 RMB/ton steel. Meanwhile, when the phosphorus content of hot metal is lower than 0.165%, the De-P cost of mode C is higher than mode B’s. For BOF single slag melting process, the dephosphorization cost calculated by model is between ± 2%, proving that the model is reliable for the prediction of dephosphorization cost.

Jan 1, 2015

By F. A. Gates, Bengt Samuelson, James E. Edmunds, S. McCune, Neil Hario, Norman L. Weiss, J. M. Bertram, A. M. Cavaliere, F. M. Jr. Stephens, D. D. Chiang

DESIGN BASIS Size of Project The size of the project generally is expressed in tons of ore milled per day, but company policies differ widely in this respect with a few adhering to this set figure but many expecting the design capacity to be liberal. To some companies that follow the latter practice a 25,000-tpd milling plant that cannot reach 30,000 or 35,000 tpd with¬out major additions is a disappointment. The engineers responsible for planning the operation must know what is intended. Other projects are sized according to the tonnage of product rather than of mine ore; examples are iron ore mills that express size in terms of tons per year of pellets or high-grade concentrates, heavy minerals plants in tons per year of rutile or ilmenite, cement plants in barrels per year, potash in tons per year of muriate, phosphate in tons per year of washer rock and flotation concentrate, and others. However, tons per day milling rate is also the common denominator in most of those instances because the major equipment has to be selected on that criterion. Equally important are operating schedules, particularly the number of operating days per week and shifts per day for ore delivery, crushing, grinding, and processing. Hours per shift also are an impor¬tant factor in design, particularly in the case of ore delivery, and the lost time expected for maintenance, cleanup, inspection, preventive maintenance shutdowns, and many others, has to be taken into ac¬count. Frequently, the ore deposit will vary in hardness and grade to such a degree that the capacity of the mill will vary widely from year to year and cannot be expressed as a fixed figure, but rather as an average over a calculable period, such as ten years, or as a specific daily capacity over the first two years, next four years, etc., or some similar projection. A good example of this kind of situation is Asarco's Mission operation 15 miles southwest of Tucson, Ariz., described by Weiss and Vincent,1 where the major rock types differed widely in hardness, so that a change in the proportion of these types in the mill feed could have imposed difficult design problems if a constant tonnage rate had been expected. Expected Life The life of a mining operation depends upon the size of the deposit and the rate of mining. The latter is a policy decision based on an economic analysis2 which, in turn, is based on many factors. From the point of view of mill planning, the estimated life of the operation determines many criteria and affects many decisions on strength of structures and degree of protection of men and materials, the quality of the materials-handling and process machinery, the type of delivery systems for ore to the mill and concentrates to the smelter or market, environmental considerations, and many others. Costly installations that may be excellent investments for a 20-year project may be waste¬ful for a 5-year operation. An example is the comparison of different methods of crushing or grinding. Take single- vs. multiple-stage grind¬ing-the former may be found more economical for a 5-year operation, and the other, more costly method more economical for a 15- to 20-year life. Mining and Ore Transportation The mining method and the means of transportation of the ore to the first crushing stage have far-reaching effects upon mill design, particularly the crushing sections, and upon site selection (see Chapter 2, Situation of Mill). Consequently, mill design cannot advance until the mine planning has reached at least tentative decisions concerning these important matters, or reach the final stages until much is known about the ore that may be expected month by month and year by year for a reasonable period. Methods of ore transportation used in over 90 underground and open-pit mines are listed briefly in Section 32 (Table 1), and more detail is found later in this section and in Section 10, Storage and Transport. These methods include surface railroad, trucks, conveyors, surface skips, underground skips, and tramming from hoist or tunnel. In surface mining, transportation of the uncrushed ore by rail or truck to the primary crushing plant-conveniently located for lowest overall cost-is most common. Rail transportation provides low perton-mile costs. Truck transportation, though in some cases less eco¬nomical, is more flexible and more adaptable to mountainous terrain; furthermore, recent improvements have reduced the operating costs to keep trucks competitive on low-grade ores. The design of the primary crushing plant will be essentially the same whether delivery is by truck or rail, but the location of the plant is subject to more rigid restrictions in the latter case because of grade limitations, and these in turn may adversely affect the selec¬tion of the mill site as well. The influence of truck haulage in reducing the stripping ratio in open-pit mines is an added reason for predicting a minor role for rail transportation in the future. The use of conveyors to transport ore from the pit to the milling plant involves (in most cases) crushing the ore to a reasonable size (10 to 20 in. maximum with today's crushers) before conveying. The Twin Buttes system3 which went into operation in late 1969, utilizes primary crushers (two) and conveyors to receive rock waste, ore, and low-grade stockpile material from trucks, then transport them to the appropriate destination. Flexibility is dependent upon the porta¬bility of the primary crushers. In the face of this development, inclined skip hoists are no longer important to the mill designers. Underground mines generally pose fewer problems of design; these will be discussed in Chapter 2. Character of Deposit The degree of homogeneity of the ore in various parts of the mine is an important factor. The characteristics that concern the millman (approximately in order of importance) are: hardness, assay, metallurgical response, size distribution,4 and moisture content. Under metallurgical response may be included degree of oxidation, mineral¬ogy, clay content, and reagent consumption, as well as process re¬sponse in general. Short-term variations-hour-to-hour, day-to-day, or even week-to-week-make operation of the plant more difficult, require greater flexibility of process design and, perhaps, also require more instrumentation and manpower. Uniformity of the ore in the characteristics noted previously and

Jan 1, 1985

By Adam Bedell, Brad Crenshaw

Tunnel project contract documents often present groundwater infiltration requirements for components of the underground portions of the project. All too often these documents require quantifiable target inflows for these components, but only refer to them in subjective terms such as “dry,” “manageable,” etc. The Contractor’s definition of these terms will inevitably differ from the Engineer’s definition of these terms. The purpose of this paper is to look at common subjective terms used in documents and provide suggestions for reasonable quantifiable language. Additionally, the paper will provide information and recommendations on relative levels of cost and effort increase needed to provide increasing levels of water cutoff (i.e., achieve “dryness”?).

Dec 1, 2022

By Raymond F. Dalton

Despite the ready availability of numerous zinc sources, the process of solvent extraction- electrowinning (SX-EW) has as yet found little favour in the zinc industry. It is suggested that this is largely due to the previous lack of availability of suitable selective reagents having the particular features required to enable them to fit in to a solvent extraction-electrowinning based process for zinc, particularly with regard to strength and selectivity for zinc over iron. This paper describes work on the development of new selective extractants for the recovery of zinc from both sulfate and chloride aqueous solutions and suggests how they might be integrated into new zinc processes.

Jan 1, 1994

By George J. Saulnier

The U.S. Geological Survey has developed mathematical-modeling techniques to assess potential impacts on the hydrologic system resulting from mining of coal and oil shale in northwestern Colorado. Potential impacts of oil-shale mining on ground-water hydrology have been simulated using a quasi-three- dimensional finite-difference model. The model has been used to simulate hypothetical mine-dewatering operations at two proposed oil-shale mines. Dissolved-solids distribution and solute transport were simulated with a three- dimensional finite-difference model that was used to evaluate the effects of flooding in hypothetical abandoned mines. The potential impacts of coal and oil-shale mining on watersheds were investigated using a digital watershed model. This deterministic physical- process model predicts responses of the hydrologic system to modifications resulting from mining in basins having snow as the major source of runoff and recharge. Potential impacts of coal development on surface-water quality were investigated using a river-quality assessment model.

Jan 1, 1979

By C. E. Brechtel

A testing program was conducted to design a cemented backfill fur hydraulic placement in underhand cut and fill mining of a steeply dipping coal seam. The testing evaluated the use of three different backfill aggregates, locally available at the coal mine, with Portland cement and five different pozzolans, including a Class F fly ash, a Class C fly ash, a cement kiln dust and ground smelter slag. Both viscosity and uniaxial compressive strength of backfill mixtures with different pozzolan and Portland cement contents were used to evaluate their suitability. Based upon the results of preliminary testing, an optimum mixture of alluvial sand with six percent Portland cement and ten percent Class F pozzolan was chosen for further compression testing.

Jan 1, 1985

By Frederick Laist

(Butte Meeting, August, 1913.) WHILE remodeling No. 1 section of the concentrator at the Washoe Reduction Works of the Anaconda Copper Mining Co. during the summer of 1912, for the purpose of ascertaining what additional metal recoveries could be made by improvements in concentration practice which had been developed by the staff of the Boston & Montana Reduction Department, at Great Falls, it was deemed advisable to carry on some experiments on the treatment of the regular mill tailings by roasting and leaching, to see whether the losses of metal values might not be still further decreased. In the course of these experiments about 5,000 tons of tailings were roasted, and of the resulting calcine about 200 tons were leached. The results were so satisfactory that it was decided to continue the work on a larger scale, particularly as regards the leaching end of the process, and construction work on an 80-ton roasting and leaching plant was accordingly commenced during the latter part of February, 1913, and is being pushed as rapidly as possible. 80-Ton Leaching Plant. This plant will consist of one 20-ft. roasting furnace of the MacDougall type, specially adapted to the work of roasting for leaching; two leaching tanks, .32 ft. in diameter by 12 ft, deep, together with the necessary solution tanks, air lifts, precipitating launders, etc. It has been decided to precipitate the copper from solution on scrap iron, so as not to hamper the roasting and leaching departments by any weakness that might develop in the precipitating department. It is, however, the intention to experiment with a number of different methods of precipitation ; for example, by electrolysis and by hydrogen sulphide.

Jan 7, 1913

By Frederick Laist

(Butte Meeting, August; 1913.) WHILE remodeling No. 1 section of the concentrator at the Washoe Reduction Works of the Anaconda Copper Mining Co. during the summer of 1912, for the purpose of ascertaining what additional metal recoveries could be made by improvements in concentration practice which had been developed by the staff of the Boston & Montana Reduction Department, at Great Falls, it was deemed advisable to carry on some experiments on the -treatment of the regular mill tailings by roasting and leaching, to see whether the losses of metal values might not be still further decreased. In the course of these experiments about 5,000 tons of tailings were roasted, and of the resulting calcine about 200 tons were leached. The results were so satisfactory that it was, decided to continue the work on a larger scale, particularly as regards the leaching end of the process, and construction work on an 80-ton roasting and leaching plant was accordingly commenced during the latter part of February, 1913, and is being pushed as rapidly as possible. 80-Ton Leaching Plant. This plant will consist of one 20-ft. roasting furnace of the MacDougall type, specially adapted to the work of roasting for leaching; two leaching tanks, 32 ft. in diameter by 12 ft. deep, together with the necessary solution tanks, air lifts, precipitating launders, etc. It has been decided to precipitate the copper from solution on scrap iron, so as not to hamper the roasting and leaching departments by any weakness that might develop in the precipitating department. It is, however, the intention to experiment with a number of different methods of precipitation; for example, by electrolysis and by hydrogen sulphide.

Jan 7, 1913

By Charles Kirchhoff

For twenty years it has been my work to watch and record progress in both the technical and the commercial branches of mining engineering in the wide sense in which it is represented by our Institute. While you who have been actively engaged in our profession have made history, a few of us have attempted to be the historians and have been kept keenly on the qui vive to follow the developments of so broad a field. One so employed may be pardoned if he ask you to pause for the unusual purpose of looking backward, that you may measure what you and others have done, and that the younger men may gather inspiration from the work accomplished by their elders and their predecessors. It is not so long since that the commercial management dominated our industries. Bred in the school of science, which seeks its reward in establishing the truth, our engineers were sometimes indifferent to trade conditions and requirements. Much of that feeling has disappeared, and the engineer recognizes that it is his mission to apply the achievements of science to a useful commercial end. On the other hand, business men now understand that the utilization of our natural resources and the production of articles for consumption is impossible without the intervention of the engineer. They must co-operate, unless, indeed, the two functions are combined in one man —a rare but powerful combination. Now the final measure of success of both is the cost sheet on the one hand and the expansion of markets on the other, so closely interwoven, so mutually interdependent. There is one cardinal feature, however, which characterizes the engineer's work—namely, the permanence of his achievements. Every improvement, particularly when it has been recorded in such lasting form as a paper in the Transactions of our societies—

Jan 1, 1900

By Edward F. Divers

Face ventilation is generally the most critical area in the coal mine ventilation system. Various studies have repeatedly shown that good ventilation is by far the most effective and least costly method of controlling dust and gas problems created during face operations. The Bureau of Mines has recently developed and assisted in the development of several new concepts in face ventilation systems, and in methods to accurately evaluate these systems underground. A new type of scrubber and ventilation system design built into the auger support frame of a continu- ous miner was developed. The objective was to improve face ventilation and reduce respirable dust concentrations. With this arrangement, one scrubber is mounted into each side of the boom. Laboratory and underground tests of various scrubber discharge arrangements showed good to excellent dust and methane control. Combined face ventilation systems, using both blowing and exhaust curtain, potentially offer dust and gas control and deeper cutting. Bureau tests at a mine using a combined face ventilation system and remote- control-operated continuous miners showed excellent dust and methane control. The advancement of line curtain systems beyond permanent supports without exposing workmen to roof fall hazards can be accomplished only with specially designed extendable devices. Numerous concepts for line curtain extension have been developed and tested. For the past several years, the Bureau has been testing various types of extensible face ventilation systems. These extensible systems can generally be classified as floor-, roof-, machine-, and rib-mounted systems. In the past 6 years, approximately 40 systems have been evaluated for feasibility. From these, several have been selected as practical or innovative and are presented herein. A technique developed by the Bureau to accurately and safely evaluate the effectiveness of face ventilation is also presented.

Jan 1, 1982

By M Begbie, R J. Norris, D J. MacKenzie

The Otago Schist is crossed by numerous deformation zones that formed in the latter stages of metamorphism and uplift of the metamorphic belt. These deformation zones are long (tens of km) linear features at the regional scale, and are generally narrow (km scale or less). Many of these structural zones separate schist domains with different rock types and structural/metamorphic histories. The Hyde-Macraes Shear Zone is one such deformation zone, and this zone hosts the active Macraes gold mine. The Rise & Shine Shear Zone, 80 km NW of Macraes, is also enriched in gold and has been mined historically. Mineralisation in both these shear zones occurred during the transition from ductile to brittle deformation as the schist belt was uplifted from greenschist facies metamorphic conditions. Recrystallisation of quartz and micas in the early stages was overprinted by cataclasis, with sulfide and gold precipitation throughout. Many of the other late metamorphic deformation zones formed under similar greenschist to sub-greenschist facies conditions, with similar overprinting of recrystallisation textures and folds by more brittle textures such as breccias. Under these conditions, originally regionally pervasive deformation at the highest metamorphic grades was evolving towards more focussed high strain zones at lower temperatures and pressures. Shear zones initially focussed fluid flow along microshears and mineral grain boundaries, resulting in 1 to 100 m scale alteration and replacement in the host schist, with only local development of quartz veins. Fluid flow was not controlled by open fractures, and the rate of fluid movement was probably slow and pervasive. After uplift into the brittle regime, the schist belt underwent tectonic extension. Extension was accompanied by development of swarms of steeply dipping veins, with localised mineralised normal faults. Hydrothermal fluid flow throughout this uplift history was a normal consequence of late stage devolatilisation of the metamorphic belt. Hence, all late metamorphic deformation zones should be considered prospective for gold accumulation.

Jan 1, 2005

By D. Randolph Berry

Volume II contains appendices which explain and document the recommendations of Volume I. The following topics are covered: Appendix A Discussion of Recommended Guidelines Appendix B Worldwide Use of Oxygen Self-Rescuers Appendix C Hypothetical Use of Oxygen Self-Rescuers in Post-1972 Coal Mine Explosions and Fires Appendix D Vibration Testing of SCSRs Appendix E Testing of In-Service Chemical Oxygen SCSRs Appendix F Discussions with Representatives of Underground Coal Mining Appendix G Recommendations for SCSR Storage Containers Appendix H Example SCSR Storage Plans Appendix I Testing of In-Service Compressed Oxygen SCSRs Appendix J 30 CFR 75.1714, 30 Regulations Pertaining to Self-Rescuers

Jan 1, 1981