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By A. Atrafi, L. Barroilhet

The top size of particles in the flotation of potash ores, which can be as large as 4 mm, exceeds by far the top size in other flotation plants. It is then understandable that while long chain amines are used in the flotation of fine fractions (below 0.8 mm) of potash ores the coarse particle flotation also requires an extender oil. While new and very promising fluidized-bed flotation machines (e.g. HydroFloat) constitute a significant development in the flotation of coarse particles, in the area of extender oils, as reviewed in 2007 by Strathdee, Gotts and McEachern, there have been no new developments to replace the industry standard asphaltene-based carcinogenic extenders. This situation first of all results from the experimental problems that arise when water-soluble minerals like sylvite are to be tested in saturated brine. Both oils and longchain amines are water/brine insoluble. The oil with the amine dissolved in it can however be emulsified and then used as an extender oil. Such oil droplets must wet KCl particles and the emulsion must be stable to avoid handling problems. In this paper we test the use of the phase inversion technique as a method of studying wettability of KCl particles by the oil in saturated brine, and the effect of amine dissolved in this oil on the wettability. The results show that this relatively simple technique can be used to characterize the wettability of fine KCl particles by the tested oil and to determine the effect of added amine on it, and so it can be useful in selecting proper extender oils.

Jan 1, 2016

Producing robust agglomerates with desirable structural properties is the key to success of the heap leaching of complex, low-grade nickel laterite ores. Mechanically strong but porous agglomerates with narrow size distribution (e.g., 5 ? 40 mm) are required to build permeable and perennially stable heaps of 4 ? 10 m height. The characterization of granules? microstructure is essential for greater understanding and correlation with the concomitant mechanical properties and robustness. In this study, we used non-destructive, computed X-Ray tomography analyses and scanning electron microscopy (SEM) in tandem with conventional compressive strength,and integrity/stability tests to draw links between nickel (Ni) laterite agglomerate strength/stability, internal microstructure (e.g.,porosity, inter-particle bridges) and binder type. It is shown that binder composition (e.g., tap water vs. 30% w/w H2SO4) and post-agglomeration drying both significantly impact on granules? physicochemical and microstructural properties. Acidic binder, in contrast with tap water, leads to significant incongruent leaching of different species from particles during ore-binder mixing and agglomeration. Upon wet agglomerate curing, solidified/recrystallized leached species form strong intra-granular solid bridges which, lead to greater compressive strength of dry agglomerates. Structural stability/integrity of cured agglomerates in acidic solution, however, decreases dramatically due to greater porosity and acid solubility of intra-granular solid bridges. These findings are useful in shedding more light on agglomerate structure characteristics and their optimization for enhanced column and heap leaching with concentrated sulphuric acid.

Sep 1, 2012

By H. J. Seim

An ion-exchange-instrumental analysis technique is described for analyzing impurities at 1 to 1,000 parts per million level in rare-earth metals. A cation-exchange flame photometric method is used for the alkali and alkaline earth groups of metals, and an anion-exchange procedure is used in conjunction with polarographic and X-ray fluorescence techniques for selected transition elements. The standard deviation is about 6 percent, and the average error, about 5 percent.

Jan 1, 1962

By J N. Farrell

Metallurgists frequently use a small number of large samples to characterise expected mill performance, with the a priori assumption that these few samples are representative of the entire deposit. Geometallurgical sampling is becoming more prevalent as a way to use many smaller samples to indicate the expected recovery as part of the Mineral Resource or Ore Reserve estimation.The evaluation of magnetite deposits involves specialised approaches to sample preparation, sample analysis and resource estimation. The sampling protocol for magnetite deposits will typically include the use of Davis Tube analysis. The Davis Tube is a mini-pilot plant that is used for determining the mass magnetic recovery from a small amount of sample ground to a specific size. The magnetic concentrate may then be analysed and together with the mass magnetic recovery, may be used to simulate the expected mill recovery. The magnetite sampling protocol is specific to each deposit, and is typically designed around a specific metallurgical flow sheet.During resource modelling, the mass magnetic recoveries and concentrate grades may be estimated directly or assigned to a block model using regressions with other chemical constituents. For estimation, the concentrate grades are weighted to account for the different mass magnetic recoveries of each sample and the impact of these differences on the concentrate estimates.

Sep 5, 2011

By H. W. C. Prommel

While Utah remains an unimportant factor in oil production, exploratory drilling was carried on in six widely separated areas, extending from the southwest to the northeast corner of the state. One new gas field, the Last Chance or Starvation Creek anticline in Emery County, and two new gas horizons in the Clay Basin field of Daggett County, were thereby discovered during 1934. Exploratory Drilling South Central Utah Emery County—Last Chance Anticline.—The test of the Ramsey Petroleum Corporation in sec. 18, T.26S., R.7E., on the apex of this anticline (vert. closure 700 ft.; area within closing contour line, as mapped by U.S. Geological Survey, 26,000 acres) encountered gas in the Moenkopi formation of lower Triassic age at depths ranging from 2530 to 2863 ft. The total thickness of sandstone, sandy shale and -limestone beds in which gas was encountered is approximately 50 ft. The well was tested at a depth of 2755 ft., gaging 21,000,000 cu. ft. of gas, R.P. 410 lb. The well is being plugged back to the lowest gas horizon at 2771 ft. from a total depth of 3168 ft. Gas from this well probably will be used for the drilling of a second well structurally lower, to test the oil possibilities of the Moenkopi formation as well as of the upper and lower Pennsylvania formations, of which the latter so far remains untested in South Central Utah. The well was not drilled deep enough to determine more definitely the western margin of the petroliferous Paradox Basin of lower Pennsyl-vanian age of southeastern Utah. Woodside Anticline.—This field, lying immediately northeast of the San Rafael swell, contains, according to maps of the U.S. Geological Survey, 8380 acres within its lowest closing contour line 700 ft. below the apex. A well was drilled by Utah Oil Refining Co. at the apex in sec. 12, T.l9S., R.l3E., during 1924. Gas was encountered in the Moenkopi and Kaibab formations of Triassic and Carboniferous (Permian) age respectively. The flow is estimated at 21,000,000 cu. ft., R.P. 900 lb.,

Jan 1, 1935

By Peter Fullagar, Adel Vatandoost, Steve Walters, Toni Kojovic

"Traditional comminution tests are a vital component in every mine plant design. Comminution behavior governs both the cost of equipment and its operating efficiency. Most comminution tests are conducted off site on a limited number of samples, owing to the high costs and long lead times involved. Therefore a strong incentive exists to find alternative means for efficiently characterizing comminution behavior for a truly representative suite of samples. The potential for inferring comminution behavior from drill core is investigated in this paper. Our approach involves development of small scale comminution tests on the one hand, and measurement of a wide range of core attributes on the other. Our focus is the relationship between petrophysical properties and comminution behavior.Petrophysical properties of archival diamond drill core from the Ernest Henry Mine (Queensland, Australia) have been measured using a Geotek multi-sensor petrophysical core logger. We have investigated the relationship between the petrophysical properties and the corresponding small scale A*b and estimated Bond mill work index test results for 2m composites of the core. Core samples were petrophysically classified by means of cluster analysis. Comminution behaviour was then modeled for each cluster via multiple regression. The results suggest that prediction of the comminution behavior on the basis of petrophysics is viable at Ernest Henry.INTRODUCTIONMineral processing is the most important downstream operation for mining engineers and metallurgists since most mined materials are subjected to some type of comminution or beneficiation operation (SME, 1992). Comminution or size reduction is typically the first stage of processing. Size reduction steps (crushing and grinding) in any plant consume large amounts of energy, normally more than 50% of total energy (Weiss, 1985). Therefore, it is desirable to obtain laboratory results to design the size reduction equipment appropriately. One of the main factors that control the selection of hard-rock crushing equipment and the design of a crushing circuit is the mechanical character of the ore(s). Work index, density, abrasivity, compressive strength and clay content are essential parameters for ore characterisation (SME, 1992)."

Jan 1, 2009

By Thomas C. Mowatt

Preliminary Bureau of Mines studies of mafic igneous rocks collected from the Rabbit Creek area, northwestern Alaska, indicated concentrations of platinum (412 to 1,406 ppb) and palladium (343 to 892 ppb) in five of eight samples analyzed. The samples represent boulders and cobbles of mafic/ultramafic rocks from this locality. They are not from outcrop. Petrologically, (on the hand specimen scale) the samples are crudely to well-layered and are troctolitic ± gabbroie in character, with primary igneous mafic minerals strongly altered. Small patches of interstitial sulfide minerals, including chalcopyrite, bornite, pyrrhotite, pyrite, and covellite are associated with these altered zones and elsewhere through the specimens. Copper values ranged from 0.1 to 0.3 pct in samples analyzed semiquantitatively; nickel, cobalt, and chromium contents are low. Chemical, mine alogic, and petrographic relationships suggest that the rocks of similar character outcropping in northwestern Alaska could also be considered as prospective for- concentrations or platinum, palladium, and copper.

Jan 1, 2011

By Schofield D, Denby B, Stout F, Moli A

This paper documents the development of a user-friendly computational program, which has been constructed to assist the mine ventilation engineer in the optimal design and operation of methane drainage ranges. In the past, many computer tools for ventilation engineers led to the development of `research systems'. These computer systems, although developed with the needs of the industry in mind were due to a combination of inexperienced programming methods and the availability of suitable user interface software not easily accessible to the practising engineer. Recent developments in computer hardware and software (particularly windows based graphical user interfaces) mean that engineers no longer need to dedicate large amounts of valuable time to become familiar with often complex items of software. The authors of this paper intend to demonstrate a practical system which successfully integrates complex mathematical models, an easy to use interface and knowledge based techniques. The graphical interface guides the user to draw on the screen a mimic diagram representation of the methane drainage network. The user is prompted, during the construction of the drawing, to input the information and data required by the calculation and knowledge based procedures. The computer program then calculates all of the flows, purities and pressures throughout the network. Given reliable measured data from an existing range a comparative analysis can be performed between the predicted and measured network flow data. A knowledge-based system may utilise this analysis to either update the operational status of the network or suggest either a reasoned maintenance or remedial strategy to improve the performance of the drainage system. The development of the OMEN project has been carried out at the request of the Midlands Group, British Coal Corporation, Edwinstowe. The system is about to be field tested at a number of mines. It is hoped that when this paper is presented that the system will be used at both Group and mine level.

Jan 1, 1995

By Ruixiang Gu

The de-confinement mechanisms of unstable failures as they occur in deep coal mining conditions are investigated using a distinct element numerical modeling code. The Mohr-Coulomb and continuously yielding joint constitutive models were used to study the compressive failures in coal and slip behavior of coal-rock interfaces, respectively. The effect of strength variations within interfaces on the failure mechanism of mining faces is also studied. The stability of interface slip failures was found to have a significant role in the failure mode of coal sidewalls. When the interfaces experience stable slip failures, the sidewalls are more likely to fail in a stable manner. With occurrence of unstable slips at interfaces, unstable compressive failures of sidewalls are triggered. The numerical modeling results support the proposed de-confinement mechanisms and emphasize the importance of including the coal-rock interface behavior in the analyses of unstable compressive failures. The significance of the proposed mechanisms in underground mining conditions is that rock discontinuities with large cohesion drop can become potential factors, inducing unstable compressive failures at ribs and mining faces. Therefore, the de-confinement mechanisms should be taken into consideration in mine design.

Jan 1, 2014

In 2001, 14,623 mining operations reported employment statistics to the Mine Safety and Health Administration (MSHA). This was an increase of 210 mining operations compared to 2000. Almost half (49%) were sand and gravel mines, followed by stone (29%), coal (15%), nonmetal mines (5%), and metal mines (2%).

May 1, 2003

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By J. E. Scott, L. M. Heaman, R. A. Creaser, G. S. Salazar

"Schaft Creek is a calc-alkaline porphyry Cu-Mo-(Au) deposit located in northwestern British Columbia. The deposit formed in a volcanic arc setting, likely on a back-arc rifted continental fragment. It is hosted by Late Triassic basaltic to andesitic volcanic rocks of the Stuhini Group (Stikine Terrane), and is associated with porphyritic granodiorite dikes emanating from the nearby Hickman batholith. The age of the Hickman batholith is approximately constrained here by a composite U-Pb zircon date of 222.1 ± 9.6 Ma, which is in broad agreement with a well-constrained age for mineralization at Schaft Creek of 222.0 ± 0.8 Ma (Re-Os molybdenite). The deposit is in most respects typical of calc-alkaline porphyry systems, but displays silica-poor sericite-chlorite alteration in mafic country rocks in place of classic phyllic altera¬tion. The deposit comprises three distinct, but related, zones: the northern Paramount zone, the Main zone, and the West Breccia zone. Two phases of mineralization are observed. The first phase occurs as hydro-thermal veins and breccias, and minor disseminations. It consists of bornite, chalcopyrite, molybdenite, and pyrite with potassic and sericite-chlorite alteration. The second phase is minor and consists of veins of molybdenite ± specularite, as well as Cu-Pb-Zn sulfide veins without any significant corresponding alteration. Extensive structural modification has affected the deposit both during and after its formation.Schaft Creek est un gîte porphyrique de Cu-Mo-(Au) calco-alkalin situé dans le nord-ouest de la Colombie Britannique. Ce gîte s’est formé dans environnement d’arc volcanique, probablement situé sur un fragment continental bordé par un bassin d’arrière-arc. Le gîte est encaissé par les roches volcaniques basaltiques à andésitiques du Groupe de Stuhini (Terrane de Stikine) d’âge Triasique tardif, et est associé à des dykes de granodiorite porphyrique provenant du batholithe de Hickman avoisinant. L’âge du batholithe de Hickman est estimé par un âge U-Pb composite sur zircon de 222.1 ± 9.6 Ma, lequel est grossièrement en accord avec l’âge de la minéralisation de Schaft Creek, estimé a 222.0 ± 0.8 Ma (Re-Os sur molybdénite).Ce gisement, bien qu’à plusieurs égards typique des systèmes porphyriques calco-alcalins, présente une altération en séricite-chlorite pauvre en silice dans l’encaissant mafique a la place de l’altération phyllique classique. Ce gîte comprend trois zones distinctes bien que connexes : la zone Paramount au Nord, la zone Principale et la zone West Breccia. Deux phases de minéralisation sont observées. La première phase se présente sous forme de veines et de brèches hydrothermales ainsi qu’en disséminations mineures et consiste en bornite, en chalcopyrite, en molybdénite et en pyrite avec une altération potassique et en séricite-chlorite. La seconde phase est mineure et consiste en veines de molybdénite ± spécularite ainsi qu’en veines de sulfures de Cu-Pb-Zn sans association à une altération particulière. Le gisement a été profondément modifié par la déformation a la fois durant et après sa formation."

Jan 1, 2008

The Research Centres Program initiated by the Federal Government and supported by the mining industry has had a major impact on geoscience research and education during the past five years. Eight major university based-Geoscience Research Centres are currently supported by the Government, including four Key Centres of Teaching and Research and a further four Cooperative Research Centres (CRCs). The Research Centres Program has led to a concentration of high quality research and teaching in areas of demand and strategic importance creating strong university - industry linkages and a source of well trained and motivated graduates. Over the last decade Australia's reputation as a leader in mineral exploration and ore deposit research has grown substantially, in part because of the development of the Research Centres Program, and to increased collaboration and interaction between research groups and exploration geoscientists across the industry. During a period when ore deposit research activity has been decreasing in other countries (eg USA, Canada, Japan and most European countries), equivalent research activities in Australia have been increasing, within a well co-ordinated national framework. The future challenge for universities and the mining industry will be to maintain the momentum of the successful research centres in order for the centres to become self-funding over a six- to nine-year period.

Jan 1, 1996

By H. L. Riley

Cleaning and handling of fine coal are generally conceded to be more difficult problems than preparation of the coarser sizes. Increased mechanization in the coal-mining industry has, in general, resulted in an increased percentage of the fine sizes in the total coal mined. If the coal fines are cleaned by wet methods, the minus-1/16-inch square-mesh particles cause most of the sludge and slurry problems encountered in preparing wet coal. Equipment that separates coal from refuse on the basis of specific gravity is not efficient on sizes that, because of their fineness, do not follow the same laws of hindered settling as the larger particles. Probably the best solution of fine-coal cleaning problems is to use other methods, such as flotation, that utilize other physical characteristics than those determining hindered-settling ratios. Regarding the type of flotation known as the Elmore vacuum-flotation process there has been a scarcity of data and information available in this country. An investigation has therefore been made to obtain additional date and information on this process under laboratory conditions. During this investigation an effort has been made to evaluate the relative efficiencies of vacuum and mechanical laboratory froth flotation cells. The investigation included float-and-sink, screen-sizing, and flotation tests on 2 different samples of 14-mesh to 0 raw coal and an a sample of slurry from the settling-tank overflow of a typical washer. One raw-coal sample was obtained from the American bed and represented a type of coal that is relatively easy to clean. The other raw-coal sample was from the Mary lee bed and represented a type considered difficult to clean.

Jan 1, 1954

By Ellis S. Hertzog

The relatively large amount of phosphorous occurring in many iron ores limits their usefulness aid sometimes renders them valueless under present conditions of nip-iron manufacture. A practical method for removing the phosphorus would be valuable to the iron and steel industry. Gravity concentration has been of little value as far as the removal of phosphorous is concerned, because the compounds of this element are, as a rule, disseminated very finely. Removal of the phosphorous by leaching is not a new idea and has been the subject of some investigation,3,4/ out for data :lave been published. W.R. Crane published analyse of various ore bodies in the Birmingham ham district. The hard or unleached ores a .re much higher in phosphorous than the Soft, leached ores along the outcrops. This fact indicates that as mild a leaching agent as water removes phosphorus effectively if enough time is allowed for it to react. As the forms in which phosphorous occurs in the iron ores are important in the consideration of any leaching operation, much time was expended in devising a method for determining them. According to Professor McVey, University of Alabama, the phosphorous-bearing minerals in the red iron ores of Alabama have not been positively identified by any microscopic methods.

Jan 1, 1935

By T. W. West

INTRODUCTION This paper presents an assessment of conventional flotation technology. In addition it presents some ideas which can be used to change the perception of what may be done with conventional flotation equipment to maximize its efficiency. Perspective of Conventional Flotation Technology 1. Flotation in many coal preparation plants is defined as something between a necessary evil and a non-viable concept because: -Too much art is required on a continuous basis to achieve uniformly acceptable results. -It is expensive, the capital cost in dollars per ton of circuit feed for flotation and dewatering is high. -It is a difficult process to use when SO2 reduction governs system design and equipment selection. -It creates a product (fine coal froth) that is difficult to dewater in the presence of BTU/lb as-received pressures. -It is a difficult process to control from a management standpoint, there are no dependable set points to rely on. 2. Many froth cell circuits are 0.5m x 0 (nominal 28 M x 0) feed which is sourced from screen/cyclone underflows. The original intent was to obtain maximum yield at acceptable product quality and moisture using disc filters. (See Figure #1) 3. An alternate to conventional flotation and its perceived to be limits of efficiency, capital costs, etc. is to move towards such developing flotation technologies as column flotation, reverse flotation, and microbubbles. A lot of work is being done to move these concepts thru pilot stages into full scale, commercial plant operation. The majority of the coal producers and coal preparation plant designers view these concepts as being in the development/pilot plant stages. OBJECTIVES Based on the preceeding comments of the perspectives of the coal producers and plant designers, the options are limited. They can make do with what they have available to them in the way of industry standard equipment and practices or they can move forward with developing technology. I feel there is a third option: what can we really do today with what is available to us, without reinventing the wheel? APPROACH 1. When processing coal against ash and SO2 restrictions, the fine clean coal stream can be the best stream on the clean coal conveyor due to the release by size and the ability to create a high BTU/low sulfur (dry basis) product. 2. Targets: -Consistent raw coal size consist in the plant feed . -Do not put pyrites in the froth cells. -Adjust criteria defining good flotation away from yield towards maximum BTU recovery at minimum ash. -Limit the froth cell feed size consist to 0.15 mm x 0 (nominal 100M x 0). -Establish and maintain a correct, consistent percent solids in the froth cell feed. -Select flotation reagents with care. -Introduce reagents in the proper sequence.

Jan 1, 1988

By John Birkinbine

In the spring of 1878 the Pine Grove Furnace, located in Cumberland County, Pa., was blown in after lying idle for several years. The fnrnace was constructed in 1770, and for over a century it has been in almost continual operation. The plant consisted in 1877, when the writer was first called there, of a stone stack 32 feet in height, inclosing a shaft and boshes, the latter being 8 feet in diameter. The blast was supplid by two wooden blowing-tubs discharging into a third, having a floating piston sustaining a box weighted to give the desired pressure, the power being furnished by a waterwheel. A small 18-pipe hot-blast stove heated the blast. The furnace was remodelled during the winter of'77-78, and a Weimer blowing-engine with a blowing-tub 5 feet in diameter and 2 feet stroke, and the necessary boilers, were substituted for the wooden tubs and water-wheel. The stack was raised and enlarged, and provided with a bell and hopper having a central drop. Water dam and tymp were added, and general repairs were made. The hot oven, however, was not increased. The furnace had always been operated with charcoal as fuel, and its reconstruction was made with a view to continue the use of charcoal, but provision was made for ample blowing capacity should other fuels at any time be employed. Although liberal arrangements had been made in cutting wood, the stock of charcoal supplied in 1878 was insufficient for the increased requirements of the remodelled plant, and notwithstanding the purchase of some 80,000 bushels from an adjacent idle furnace, there war: riot enough to keep the furnace in constant blast until a new supply could be obtained, particularly as the season was unusually backward. Instead of following the established precedent of many charcoal furnace managers, i. e., blow out every February, it was detemined to continue in blast, using coke as fuel. Accordingly, when the charcoal stock was exhansted, on March 22d, 1879, Connellsville coke was substituted, the change of fuel being made at once; that is, the coke charges followed immediately upon the last charcoal charges. After working a few days, the strike in the Connellsville coke region cut of the supply, and anthracite coal was obtained, a mixture of the two fuels being employed. The strike continuing,

Jan 1, 1880

By I Ross, C deWolfe

Block caving has been the favoured mining method for large underground hard rock mines for the last few decades. The orebody must be of a size large enough to cave and generally amenable to a caving method. Typically, these mines ranged from 15 000 to 50 000 tonnes per day (50 kt/d) and have been some of the largest and most productive underground mines in the world.Now mining companies, in the quest for improved shareholder return, are looking for higher production rates from new underground mines. Whether a ‘Greenfields’ development, or a replacement for ageing large open pit mines, minimising capital and maximising production has been a focus of project teams. These new block cave mines are what have become known as ‘super caves’. These caves range from 80 kt/d to over 150 kt/d and are being planned or constructed by several major mining companies including Codelco, Freeport-McMoRan, Newcrest and Rio Tinto.Issues facing these super caves include high initial capital costs, very long development periods and significant technical challenges. These issues tend to make it difficult to present an attractive business case for such a project.This paper will discuss some of the potential benefits as well as the technical challenges and risks of operating a mine with the block or panel caving method at these very high tonnage rates. Considerations including geotechnical conditions, development rates, panel widths and ore column heights are discussed. The risks associated with these technical challenges and potential benefits for successful approaches will also be discussed.CITATION:deWolfe, C and Ross, I, 2016. Super caves – benefits, considerations and risks, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 51–58 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By Alexandre Alex) J. Bredikhin

"DISCLAIMER: the statements presented herein are those of the author and do not necessarily represent the views of US DoD or its Components.ABSTRACTProgress on the $750M lock project on the Monongahela River was derailed by discovery of large anomalies in the 6.5-7-ft dia. drilled shafts of the new river wall foundation. The anomalies were located 45 to 50’ below the top of the wall at the critical shaft-to-wall interface, and ranged in thickness from several inches to 2-plus feet. A Specialty team, comprised of representatives from the contractor, the Government, and structural design consultants, was formed to evaluate the extent of the problem and the impact on the new lock wall. With no standard method in place, the Government took the lead to develop acceptance criteria for the repairs while the contractor took the lead to identify the repair means and methods. The Government developed a detailed Quality Control plan to capture and document all of the repair data and to approve each shaft repair in a timely manner. This paper provides an overview of the anomalies encountered and the need for repairs; discusses the details of the repair plan; and provides details on the unique quality control measures employed to verify the repairs. By developing and adhering to a detailed quality control plan, the structural integrity of all of the shafts was successfully restored.INTRODUCTIONThe new Charleroi Locks Project is currently under construction by the U.S. Army Corps of Engineers (USACE) in Pittsburgh, Pennsylvania. In the region, the Corps of Engineers is modernizing the series of locks and dams on the Monongahela River. One facet of this modernization is replacement of the existing Locks #4 (aka Charleroi Locks) located upstream of Pittsburgh, Pennsylvania with new and larger capacity locks. The relative location of the project is shown on Fig. 1. The original construction of twin locks (56’x720’ & 56’x360’) and the fixed-crest dam was completed in 1932, with lock walls supported on timber pile foundations. The fixed crest dam was replaced with a gated dam in the 1960’s. The existing lock is undersized and in poor condition and thus needs to be replaced. The project requires the construction of the new river chamber adjacent to the existing land chamber which must remain operational to navigation throughout the construction. In order to accomplish this, the lock walls need to be constructed in the wet. The river wall was constructed using two construction methods referred to on the project as “cofferbox” and “non-cofferbox” construction."

Jan 1, 2017

By Michael L. Nave

Mining belt conveyor elements are becoming larger and more expensive. Conveyor reliability requirements are rising. Conveyor condition monitoring can aid maintenance, reliability, and risk management in conveyor system operation. This paper examines the application of condition monitoring in coal mining belt conveyor systems. Types and form factors of condition sensors are described. Example costs are included. Suggestions for OEM developments are examined. Application problems are discussed. Two simple cost case studies are presented.

Jan 1, 2000