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Council for 1995/96 President D.A. J. Ross-Watt President Elect Dr N.A. Barcza Vice Presidents Senior R.P. Mohring Immediate Past President J.A. Cruise Junior J.R. Dixon Honorary Treasurer J.A. Cruise

Jan 1, 1996

By D. S. Choi, H. D. Dahl, D. L. McCain

The paper examines current room and pillar mining practice in the U.S.A. from the standpoint of rock mechanics. Consideration is given to the deformation characteristics of the roof formation and the coal seam as determined in the laboratory as well as in the field. The results can be utilized by the practicing mining engineer to improve current mine design practice.

Jan 1, 1982

By Herbert Everest

THE idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentage expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2: Development, including all necessary yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, drawslate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1 ½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the Beadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high.

Jan 1, 1916

By W. J. Fullerton

This paper is one of a series prepared for the United States Bureau of Mines discussing the methods used in crushing and grinding materials for cement manufacture, and deal particularly with the methods used at the Howes Cave plant of the North American Cement Corporation. These papers are designed to disseminate technical information regarding the methods used. The cost tabulations represent local operating expenditures only and not total production costs. It is recognized that publication of total production costs might in many instances cause embarrassment to individual producers as well as to the industry as a whole. On the other hand, operating costs are essential to the technical discussion and study of methods employed. The attention of the reader is specifically called to this differentiation in order that no misunderstanding of the scope of the cost tabulations shall ensue.

Jan 1, 1932

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The authors claim that their proposed jack-knife method, which is a rekriging of each block by eliminating, in turn, one sample from m the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" .of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper con¬sist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data value-based measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text.

Jan 1, 1997

By Robert E. Wallhagen

This report describes the results of a four year investigation into improved face ventilation systems for underground coal mines. The objective was to experimentally determine configurations for a diffuser fan system and a water spray (spray fan) system which will provide a maximum reduction in face methane concentrations without causing an increase in methane concentrations at the operator's position. The experimental evaluation was performed in full scale mock-up mine entries, simulating continuous mining operations in both a six foot and four foot coal seam complete with normal ventilation, water sprays, and methane emission. A limited study was also performed to assess the effect of scrubber discharge location on methane concentrations in the face area. The results indicated that diffuser fan and spray fan systems that were designed to augment the natural airflow pattern in the entry could provide a factor of 3 to 5 reduction in average methane concentration at the face, compared to a mining machine with just conventional water sprays. The properly designed diffuser and spray fan systems also significantly reduced the total methane present in the entry. A properly designed dust scrubber discharge can reduce face methane levels by a factor of two.

Jan 1, 1977

By S. M. Moore

Each year more than 400 coal miners are injured (fatally and nonfatally) by rock falling from between or around roof supports. Many of these injuries can be prevented by the installation of roof screen. However, many coal mines are reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential savings in workers' compensation (W C) premiums that can be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by Illinois and Kentucky were investigated in this study. Using data obtained from the Mine Safety and Health Administration, national and state WC bodies and individual insurance companies (e.g., average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), baseline mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state's actual WC rate-setting formulas, a sensitivity analysis was conducted to determine the total savings in WC costs after a three-year period. Annual savings in WC premiums ranged from $41,000 to $326,000 for the 67-person mine, and $96,000 to $843,000 for the 150-person mine. An economic analysis of the cost of a roof screening program at a 67-person mine was also conducted. The annual cost of a roof screening program here was estimated at $240,000. At this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By Wu Chao, Wang Pinglong, Meng Tingrang

This paper discusses the research results of sulphide ore properties with respects on self-heating and the assessment techniques on mine spontaneous combustion for several mines of sulphide ores in China. A group of experimental devices for testing the chemical behaviour of sulphide minerals, the geological structures, the rate of oxygen absorption, the initial self-heating temperature, and the ignition point of ore samples, etc. are also described in detail. These methods have been proved to be effective by several sulphide mines which all have the history of spontaneous combustion of sulphide ores. INTRODUCTION A fire disaster from the spontaneous combustion of sulphide ores in a mine can produce large quantities of toxic gases and heat to pollute the working environment, lead to large economic losses and loss of human life, and even compell the mine to be closed. Some mines of copper, lead, zinc, tin and iron in China are of high content of sulphur. It is of great importance to evaluate the potential of spontaneous combustion of sulphide ores before beginning to design and mine the ore deposits. At present, the mechanism of spontaneous combustion accepted commonly is that sulphide ores undergo slow oxidation on exposure to moisture air at ambient temperature, with the evolution of heat, gases and moisture. The heat generated, if not dissipated, gives rise to an increase in the temperature of the sulphide ores. As the temperature of the sulphide ores rises, the rate of oxidation increases. If this is allowed to proceed unchecked it can eventually result in the ignition of the sulphide ores. For pyrite (FeS2), it reacts with oxygen and water according to the following equations.

Jan 1, 1993

By Oleg Terichow

Electrophoretic mobility measurements were made on samples of Florida phosphate slimes to determine the optimum conditions of coagulation and precipitation in aluminum sulfate solutions. The addition of silica sand and other selected materials as mutual coagulants has demonstrated the practicability of coagulating two dissimilar particle systems by elimination of the repulsive double layers. By proper selection of the solution and its concentration, the reduction of the double layer to a zero point of charge causes coagulation and rapid precipitation of both particle systems. Confirmation was found in filtration tests and sedimentation tests.

Jan 1, 1973

By J. L. Ellenberger

The National Institute for Occupational Safety and Health (NIOSH) has continued the research role of the former US Bureau of Mines (USBM) to engineer techniques that will reduce the hazards in the mining workplace associated with coal bumps. Recent research focused on a longwall coal mine in Utah with overburden greater than 750 m (2,500 ft) containing several massive sandstone units. The primary field instrumentation at the site was three-dimensional, full waveform, autonomous microseismic arrays placed underground and on the surface in order to surround the active multi-panel longwall district. The purpose of these arrays was to help investigate the strata mechanics associated with the redistribution of stress and the associated gob formation of the longwall. Specifically, the seismic arrays were used to determine the timing and location of the failure in the strata surrounding the active mining. Overall 13,000 seismic events were detected and located with on-site processing during the five months the panel was being mined, including a magnitude (ML) 4.2 event. Of these, a smaller subset of 5,000 well-located events was selected during post-processing to form a consistent data set for analysis in this paper. From this data set, it was observed that the seismic events generally occurred in advance of the longwall face, both above and below the panel, consistent with failure of the strata in the forward stress abutment zone. Also, the occurrence of the ML 4.2 seismic event within 150-180 m (500-600 ft) of a deep cover longwall face with no associated bump caused a re-evaluation of the nature of the connection between seismic activity and coal bumps.

Jan 1, 2002

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Jan 1, 1938

By Paul W. Kingston

In New Brunswick, during the 1970's, rocks of the Lower Carboniferous Windsor Group in the Moncton Basin of the Fundy Epieugeosyncline have been the target of intense exploration and drilling programs. Sylvinite ore was discovered in three structural basins of the Moncton Basin near Sussex through a drilling program sponsored jointly by the Federal Department of Regional Economic Expansion and the New Brunswick Department of Natural Resources. In 1973, exploration and development rights to the western end of the Moncton Basin were awarded to the Potash Company of America. One high-grade sylvinite unit, varying between 15 and 70feet in true thickness, with a strike length in excess of 5 miles, is located near the stratigraphic top of the assymetric, semi-diapiric, locally overturned anticline. Late in 1975, exploration and development rights to the Marchbank Syncline were awarded to International Minerals and Chemical Corporation (Canada) Limited. To date, a stratigraphic evaporite sequence, forming the semi-diapiric core of the syncline, has been shown to contain one major high-grade sylvinite unit averaging 70 feet in true thickness over a strike length in excess of 2 miles. Both properties appear to have potential considerably in excess of that revealed by the exploratory drilling, and both, because of numerous mineralogical, structural and stratigraphic similarities, are thought to be depositionally contemporaneous. Both potash deposits lie within 50 miles of the Saint John all-weather deepwater port, thus affording easy access to eastern and mid-western U.S. markets. Further opportunities for potash exploration in New Brunswick are considered excellent, as there are large untested areas with favourable geology and known potassium water anomalies and negative gravity anomalies occurring in the Moncton and Cumberland basins, and in the Central Basin of New Brunswick.

Jan 1, 1979

By K. V. Raj, T. Bhowmick

"Control of fugitive dust in high latitude open-pit mines is challenging due to unique atmospheric phenomenon resulting in complicated flow regimes, and atmospheric inversion due to lack of adequate solar insolation during prolonged winter seasons. In this study, two idealized (One conical and one trapezoidal) three-dimensional open-pit mine geometries are simulated for differing seasonal conditions using the CRADLE-CFD package. The airflow is modeled using Standard Kappa-Epsilon Turbulence Model solved by RANS equations. The concept of particle tracking is used to predict the flow pattern of dust particles. For various climatic conditions, and two different pit geometries, fugitive dust particles varying in size (PM0.1 to PM10) and concentration are generated at various locations of a mine and dispersed by the airflow pattern inside the open-pit mine. The amount and location of dust particles inside the pit is reported at various time intervals. The development of atmospheric inversion during the winter season significantly affected the amount of dust retention inside the open-pit domain. While during the summer, airborne dust particles were transported outside the open-pit domain in a short time. INTRODUCTIONComputational fluid dynamics (CFD) modeling of open-pit mines is challenging. In a recent paper, Bhowmick et. al. [1] presented different constraints and challenges in developing a 3-Dimensional CFD model. In order to obtain realistic results from a simulation, however, appropriate values of various inputs are required. Simulation results of a 3D non-hydrostatic model by Shi et al. [2] showed that both the mechanical and thermal forcing are important mechanisms, controlling the development of the atmosphere inside an open-pit. Alvarez et al [3] used a CFD model for predicting dust dispersions in a medium sized limestone quarry, and presented several iso-surfaces of varying dust concentrations using ANSYS CFX. A similar study by Torno et al [4] introduced a more sophisticated parameterization of the dust sources. Flores et al. [5] used an OpenFOAM CFD simulation package to simulate and predict pollutant dispersion in an idealized and an actual open-pit mine in Chile with an intense insolation condition. The model was simulated for three different conditions: (i) air advection (10 m/s) driven mechanical turbulence; (ii) surface heat flux (240 W/m2) driven thermal buoyancy; and (iii) both air advection and surface heat flux driven conditions. The results of the simulations showed that the buoyant currents contribute to the removal of a large percentage of the dust particles. The buoyant currents modify the purely mechanical-turbulence induced flow pattern and reduce the particle residence time, commonly observed in a purely mechanical turbulent flow (non-buoyant case)"

Jan 1, 2015

By Wolfgang Koch, Bernd Friedrich, Ann-Kathrin Maurell-Lopez

Due to the high standards for silver in electronic uses, it is essential to control the behavior of the main impurities such as copper or palladium during electrolytic refining. In regard to palladium, it is indispensable to understand and manage its co-deposition behavior. The influence of electrolyte composition, current density, and acid concentration on Cu and Pd co-deposition during silver electrorefining is systematically investigated.

Mar 1, 2017

By D Maconochie, S Pradhan, M Reed

The Waterview Connection will provide a high-quality strategic alternative to State Highway One and complete Auckland’s Western Ring Route connecting State Highways 20 and 16. The project includes twin three-lane motorway tunnels 2.4 km long and 13.1 m internal diameter (ID). The tunnel construction started in November 2013 using the tenth largest earth pressure balance (EPB) tunnel boring machine (TBM) in the world. The tunnels will pass beneath urban and greenfield areas as well as a major local arterial road. The project is being delivered by an alliance of constructors, designers and the New Zealand Transport Agency. This paper describes some of the key innovative and cost saving design solutions that have been achieved through collaboration in the alliance environment with the construction team, owner’s verifier, the segmental lining precast suballiance and the M&E suballiance for the underground works. The key design elements of the tunnel comprise a three-lane roadway, invert culvert and backfill, precast segmental lining, sixteen mined cross passages including a low point sump cross passage and steel segments for opening to the cross passages. This paper reviews interdisciplinary design integration and the economies achieved in the permanent works tunnel lining design, including reduction in lining thickness and optimising reinforcement, cross passage separation of 150 m, efficient design of the invert culvert and backfill design, main drainage pipe and low point sump. The paper also describes the attributes of the TBM and the invert culvert construction methodology.CITATION:Pradhan, S, Maconochie, D and Reed, M, 2014. Tunnel design in an alliance for Waterview Connection, New Zealand, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 425–430 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By B. A. Zaharov, J. M. Shneerson, L. V. Chugaev, M. T. Zhunusov, A. M. Markelov, S. V. Drozdov

The main direction of research was the application of autoclave technology for the flotation concentrates or biooxidation feedstocks. The industry modernization also expects the possibility of increasing the yield of the flotation concentrate, to allow an increase in gold extraction in flotation circuit. This paper provides brief description of results obtained for concentrates with the yield of 4 and 8%, as well as corresponding biooxidation feedstocks. Technological efficiency of autoclave oxidation is shown for both concentrates and biooxidation feedstocks. The process has been evaluated on a laboratory and a continuous pilot plant scale. Gold extractions from POX discharge are ranging from 94-97% for each product, with only 5-7 kg/t cyanide consumed.

Jan 1, 2014

By V. Sheikhzadeh, M. Saghaeian, N. Esmaeili, R. Hejazi, A. D. Ahmadabadi

The main challenges in most low grade iron ore deposits are finding the optimum flowsheet for beneficiation to achieve a final concentrate from both technical and economic viewpoints. Dry preconcentration techniques are being used to assist in reducing the amount of ROM feed to the plant. In this paper the benefits of utilising a dry cobbing magnetic separation technique for upgrading the product of HPGR in a closed circuit with dry screen 4 and 1.4 mm by performing a pilot scale test work was investigated. The experimental test work was done on two sample types of Yogi BIF magnetite iron ore located in Western Australia. Prior to magnetic separation studies, detailed characterisation of the feed materials including physical, chemical, mineralogical and magnetic properties were carried out. Furthermore, the effects of process parameters: drum rotation speed (1.5, 2.5 and 3.5 m/s), magnetic field intensity (1100 and 3800 gauss) and feed size (minus 4 and 1.4 mm) were investigated. The results revealed that sample type1 with 33.90% Fe could be upgraded to 41.51% Fe by using a stage of 1100 gauss dry magnetic separator with 69.12% weight recovery and 84.36% Fe recovery at a size of 4 mm. Also for sample type2, it is possible to increase Fe grade from 29.84% to above 36.68% with 75.58% weight recovery and 90.32% Fe recovery by applying one stage of 1100 gauss dry magnetic separator to a 4 mm size sample. Keywords: Banded iron formation (BIF), pre-concentration, dry cobbing technique, magnetic separator

Jan 1, 2020

By Clarence R. Claghorn

Few facts relating to the Bernice anthracite coal-basin have been published in our Transactions, and little is known among engineers in general of the character and composition of the coal found in that basin.

Jan 1, 1889

To carry out its responsibilities the SECV needs to manage the visual environment effectively. Every land oriented decision effects what we see and the value in pursuing visual excellence is supported by informed opinion, widespread community interest and evidence of cost- effectiveness.By knowing how we perceive the landscape and then changing or conserving it we can assist in fostering a positive mental image of a countyside which is welcomed by the local and regional community.It is useful to examine key design elements and the value we place on visual impressions in relation to three significant SECV rehabilitation projects, at Loy Yang, Newport and Morwell Open Cut.

Jan 1, 1992

T. T. READ, BUREAU OF MINES: Mr. McAuliffe has asked a very practical question. Expressed in another way, the question is: If we have decided to use rock dust to prevent coal dust from exploding, have we got to take rock, that is, solid material,, break it up and finally grind it into dust (and fine grinding is very expensive), or can we take material that is already in a pretty fine state of subdivision and thereby save the expense of pulverizing it? Most of the preceding discussion has been from the view-point of practical experience, and I should like to suggest that we have not had enough practical experience in rock dusting to be able to answer this question, from the viewpoint of experience, with the degree of definiteness that Mr. McAuliffe desires. With your permission I will discuss the matter briefly from the theoretical standpoint, because after all the theoretical considerations, insofar as we understand them correctly, will apply to every' case, whereas in coming to a decision on this matter it will be necessary to take into consideration local conditions, which will necessarily vary in every case. Considering the matter in a rather elementary way, the first point is: What makes coal dust explode? The air of this room is full of dust at the present moment, even though we do not notice it. If you were to close up the room for a week, enough dust would settle on the floor so you could write your name on it. Of course, the dust which is in the air of this room is not coal dust but even if it were coal dust and you were to bring an open light in' here we all know that the dust would not explode. The reason for that is that the dust particles are not close enough together for the flame to propagate itself from one to the next. You will probably be able to see in the demonstration in the gallery on Saturday that the coal dust explosion is practically a very rapid burning of the dust. When you light one piece of dust it burns and gives out heat and if there is another piece of dust close enough to it so that the heat from the first piece is able to bring a second piece up to the burning temperature, then the flame will propagate itself through the mixture. But, if the particles are so far apart in the air that the' heat from one particle is dissipated before it reaches an adjoining one, and is therefore not able to bring the second particle up to the burning temperature, then the flame will not propagate. To get a coal dust explosion, it is not sufficient to have coal dust suspended in the air. The coal dust particles must be close enough together in the air so that the heat from one burning particle is able to heat the adjoining particles up to the point at which they will burn. There has been some discussion of the size of dust particles that will cause an explosion to propagate, and it has been pointed out that the minus 200-mesh dust is the most dangerous, and that particles coarser than 20-mesh will not propagate an explosion. There are two reasons for that. A large particle has very much smaller surface in proportion to its weight than a small particle has. You know the old story about the Indian who when asked how he could go around in cold weather without any clothes on said that an Indian was all face. A 200-mesh particle is almost all surface and there= fore is able to burn almost instantaneously, which means that it gives out its heat almost instantaneously, whereas a large particle with less surface in proportion to its weight burns more slowly, gives out its heat more slowly, and the heat has more time to be dissipated. The second reason why coarse dust does not easily propagate a flame is because it is so hard to make it

Jan 1, 1924