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By Harry E. LeGrand

The purpose of this paper is to outline briefly some basic principles of ground-water hydrology that have application in mine hydrology and to suggest that an understanding of these principles can help to forestall problems that are internal as regards mining operations and external as regards damage or harm to surrounding areas. The development of a mine below the water table pleads for accurate values to answers of hydrologic problems. Such internal problems involve both an inflow of water without adequate drainage facilities and a deficiency of water for the required mining or milling use. Methods which give acceptable quantitative values for the solution of many ground-water problems depend on rather simple geologic structures, characterized chiefly by homogeneous rocks of considerable areal extent. Unfortunately, mining operations are likely to be centered where rocks are not homogeneous. Consequently, methods that attempt to show values of the storage and transmitting capacity of rocks in mining areas must be applied with caution. In spite of complex geology and apparent uniqueness of individual mines and quarries, adherence to some concepts of ground-water occurrence and movement should be helpful to mine operators. Mining operations that extend below the water table disturb the natural hydrologic regimen. Prior to development, the water table slopes away from the site of a mine in at least one direction, but as soon as development begins and water is removed from the mine, the mine becomes the apex of a depressed cone in the water table, causing a convergence in the flow of water toward the mine, The behavior of the cone of depression, as it tends to deepen and to increase in areal extent when the mine is deepened, is related to the rate of withdrawal of water necessary to keep the mine dry. The significance and nature of the cone of depression in the vicinity of pumped wells has been studied by ground-water hydrologists for years, but the cone caused by pumping water from mines has generally been less systematically appraised. The geologic and hydrologic settings of mines range greatly so that a simple standard procedure for dealing with problems

Jan 1, 1972

By Adrian Daniel

"Dating back to 1934, Con Mine is located directly south to the city of Yellowknife in the NWT, Canada. With mill operations ceasing in 2004, a challenging surface remediation situation resulted which required the treatment of close to 40,000 m3 of Arsenic impacted material.Through much of the period of 1941 to 1970, Con Mine treated their refractory arsenopyrite ore through the use of an Edwards Heath roaster. With mounting environmental pressure and through Canadian governmental studies during 1950 to 1970, attention was directed to the Arsenic Trioxide material being collected by two Doyle scrubbers on the Con’s roaster stack. In 1970, the roaster was decommissioned leaving the resulting Arsenic material in two containment areas on surface.As a result of Con Mine entering into remediation in 2004, the Con’s autoclave, commissioned in 1992, was dedicated to the immobilization of high impacted Arsenic material through the combination of excess Iron under high temperature and oxygen overpressure. Examined, herein, is the history and chemistry of the Con Mine autoclave as well as the two stages of chemistry for the Water Treatment Plant used to remove arsenic, heavy metals and WAD cyanide from tailings water. Simplified capping plans of the tailing containment areas and a high arsenic impacted area are also presented.INTRODUCTION TO CON MINECon Mine (Con) is situated just south of the City of Yellowknife in the North West Territories (NWT) Canada and is bounded locally by Yellowknife Bay (North Extension of Great Slave Lake) to the east and Kam Lake to the west (Figure 6). The discovery of gold at Con was made in 1934 and mineral rights were acquired by Consolidated Mining and Smelting Limited (Cominco) and Falconbridge soon thereafter. Nerco Minerals Limited (Nerco) took control in 1986 and in 1993, was subsequently sold to, Miramar Mining Corporation which in turn was purchased by friendly takeover involving Newmont Mining in November 2007 as a result of acquiring the Miramar’s Hope Bay project. Con’s underground operations ceased in September 2003 and have since entered into the remediation of the site. The current surface lease holdings are about 340 Hectares (Table 1) which sits on top of the Archean-aged Yellowknife Greenstone Belt."

Jan 1, 2009

By K A. Bunker, A D. Campbell, E Ah Mu, A R. Penney

The Ernest Henry mine is a transverse sublevel cave (SLC), which is relatively unique due to the 45 degree inclination of the orebody. To economically recover ore, each sublevel is horizontally offset from the previous level by around 25 m, depending on the orebody geometry. Mine development, slotting and initial production within this zone occurs in a high stress abutment that increases in magnitude at depth. These areas are seismically active and close to the hanging wall where cave propagation is yet to initiate. High stress conditions and the close proximity to the cave seismogenic zone results in challenging geotechnical conditions in the step out zone of each level and requires management of hazards including stress driven rockfalls, rock bursts and pillar failure.This paper outlines how these hazards were forecast using numerical modelling and geotechnical monitoring. Risk management techniques included dynamic ground support design, development of a new mine layout and extraction sequence as well as delayed re-entry into seismically active areas. Ongoing monitoring, instrumentation measurements and field observations were used to track the effectiveness of the controls.Future production levels are forecast to experience increasing stress and seismicity, which will increase demand on excavations and installed ground support. Ensuring the techniques and controls for monitoring, measuring and mitigating stress and seismic hazards are enforced is essential for continued safe production.CITATION:Campbell, A D, Bunker, K A, Ah Mu, E and Penney, A R, 2016. Managing stress and seismicity in an inclined sublevel cave, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 319–328 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By H. Hao

This paper outlines the development of a 2-D axisymmetric mathematical model to predict the evolution of temperature during the Direct Chill (DC) casting of AZ31 magnesium billets. Model predictions of temperature have been compared to measurements made on an industrial DC casting machine at Timminco Ltd, in Haley, Ontario. These measurements have been used to fine-tune the thermal boundary conditions employed in the mathematical model, The model, which is based on the commercial finite-element code ABAQUS, has been developed as part of a strategic project aimed at improving the DC casting process for light metals through a reduction in casting related defects and increased productivity.

Jan 1, 2004

By R. Venugopal, T. C. Rao, N. Suresh

Experiments were carried out using the cyclone to treat coal in the size range -2.0+0 mm. The diameters of the cyclone outlets (vortex finder and spigot) were varied to study their effects on cyclone performance. The cyclone products were deslimed over a screen with an aperture of 0.075 mm. The performance of the separator was studied for washing coarse coal (-2.0+0.075 mm) and fines (-0.075 mm). Performance curves were plotted for the separation of -2.0+0.075 mm coal. The concept of a generalised distribution curve was adopted to predict cyclone performance. A mathematical representation was derived from the generalised distribution curve. The specific gravity of separation for individual size fractions showed a decreasing trend with decreasing particle size down to 0.85 mm, beyond which a reverse trend was noted. It was also found that the fine size fraction (-0.075 mm) does not separate on the basis of size or specific gravity but follows the water distribution

Apr 1, 1995

By K Komori

Naoshima Smelter and Refinery has treated shredder residue and incinerator fly ash in Japan since 2003. As a highlight, a new incinerating and melting furnace, as well as fly ash washing treatment plant were newly constructed for pre-treatment of shredder residue and fly ash respectively, and the pre-treated materials are ultimately treated in the existing Mitsubishi continuous copper smelting and converting process line. Since wastes such as shredder residue tend to vary in size and in shape, and contain some chlorine, the installation of a new pre-treatment line was required. The combination of the new pre-treatment line with the existing Mitsubishi Continuous Copper Process line has created a highly efficient and environmentally friendly system for recovering and recycling of valuable metals that previously used to be sent to a landfill. This paper reviews the approach that Naoshima took to establish its current recycling business strategy.

Jan 1, 2007

By A. J. Richins, M. G. Nelson, R. Dunne

From the inception of flotation in the late 1800s, coarse particle recovery has been a challenge to practitioners. By the early 1900s, flotation processes included oil flotation, skin flotation, agglomeration flotation, and froth flotation, and for each there was a proliferation of flotation equipment. Ensuring good coarse particle recovery included desliming of the feed and proper conditioning were required. Froth flotation was at first used mainly to treat finer materials from grinding circuits. The maximum size for coarse particles was from 3.18 to 4.78 mm. In froth flotation, large particles with high specific gravity tend not to be recovered. Coal and minerals of low specific gravity and flaky minerals such as graphite, mica, and talc can be floated in larger sizes. In the early 1960s, the phosphate industry began using column cells to recover coarse particle by froth flotation, but early columns suffered from poor bubble generation and sanding at the flat base of the column. Recently, improved coarse particle recovery has been the object of new cell designs—the Eriez HydroFloat™, the Jord NovaCell™, the FLS coarseAir™ cell, and the Cidra P29 integrated system. The paper will follow the development of coarse particle flotation with a focus on the development, laboratory assessment, and operation of the new flotation equipment and technology designed to recover coarse particles.

Feb 22, 2026

By E. J. Matson

Under the investigation for critical and strategic minerals the Kaiser mine, formerly the Baxter fluorspar property, in northeastern Mineral County, Nev., was explored. The work comprised excavating 32 trenches averaging 90 feet in length, 8 feet in width, and 1 1/2 feet in depth. The trenches were excavated across the strike of the ore body to enable the engineer to spot the diamond-drill holes; 7 such holes were drilled, aggregating 3,276 linear feet. A total of 20 core samples and 107 sludge samples was taken. The work under this program indicated that the fluorspar southwest of the deposit extended laterally and also at depth at least 250 feet below the 400-foot level. INTRODUCTION The Kaiser fluorspar deposit is in the northeastern portion of Mineral County, Nev. Mining was begun about 1928, and the property has been a consistent producer of both Acid- and Metallurgical-grade fluorspar. The mineralized zone has been developed along a strike length of about 2,000 feet and to a depth of over 700 feet along the dip.

Jan 1, 1957

By Jiang Feng, Zhang Libo, Peng Jinhui, Guo Zhanyong, Ju Shaohua, Lei Ting

The clean utilization of the residue containing chloride, CuCl residue, produced from zinc hydrometallurgy is very important for the recycle of valuable metals. And drying is the main process in recycling. In this study, response surface methodology was used to optimize operating conditions of the drying of CuCl residue in a convection oven and desirability function used as the methodology for the optimization. Optimization factors were air temperature (80–100°C), material thickness (10–25mm) and process time (80–120min) while investigated responses was final moisture content. The optimum conditions were found to be the temperature of 95.48°C with the material thickness of 13.47mm for the process time of 89.48 min. Under optimum conditions, the predicted moisture content of CuCl residue was 4.11% while the experimental date was 4.13%. The experimental value after process optimization was found to agree satisfactory with the predicted value. After drying, XRD was used to analyze the main phases of CuCl residue. The CuCl residue is made up of CuCl,Cu2O, ZnS and other compounds.

Jan 1, 2015

By Walter Ensinger

Working safely means working cost-effectively. Every "dangerous occurrence" and - even more so - every industrial accident causes disruptions in the planned course of construction. Besides moral and legal obligations, economic aspects are therefore also factors motivating increased efforts towards optimal prevention: improvements in work safety and protection of health are wise investments! Ever since its earliest days, the European Federation of Foundation Contractors (EFFC) has made this subject one of its key objectives and therefore set up a Safety + Environmental Working Group, in which experts from the member organisations exchange national experience and develop proposals for "best practices". The paper presents current developments and results from the work of the S+E WG with particular reference to the project "Ideas for Safety Improvement".

Jan 1, 2006

By Carlos Kovacs-Figueroa

Because of air pollution control regulations in the early 701s it became necessary to clean effluent gases from the Morenci Reduction Works Smelter. Feasibility studies indicated that this could be accomplished most efficiently by using the smelter gaseous effluent to feed a sulfuric acid plant. Complying with air pollution control regulations thus created the need to dispose of relatively large amounts of concentrated sulfuric acid. Typical Morenci ores contain substantial amounts of oxidized copper minerals which are not recoverable by flotation. In addition, there is low sulfide, high oxide material which can be leached, but is not suitable for flotation concentration. Early in 1971 sufficent data had been accumulated to show that a leaching plant for the Morenci tailings could use a large portion of the sulfuric acid produced by the Smelter. It was determined that for all practical purposes leaching of the available copper was complete in two hours or less. Capital expenditures could not be justified for more than two stages of counter current decantation for recovery of the leached copper, and the plant was designed with these criteria in mind.

Jan 1, 1976

By William Kent

In a paper by Mr. Frank Firmstone, published in vol. iv, of the Transactions of the Institute, on " Comparison of Results from Open-topped and Closed-topped Furnaces," the regularity of the average grade of pig iron procluced by the furnace under different conditions is expressed by means of a curve or diagram. The object of this note is to show that the graphic method, or that of representing variations by plotted curves or diagrams, is capable of a much more extended application to blast-furnace records, making a valuable auxiliary to the " furnace hook," and that by it the practical metallurgist may obtain at a glance information concerning the variations of the furnace during long periods, which could be found in the furnace book only after a tedious search. Plate XII is a graphic record of an anthracite blast furnace for one month. The table on page 554 gives the figures obtained from the furnace book, together with a record of temperature of the atmosphere, barometric pressure, and moisture in the air, which was obtained from a meteorological observatory not far from the furnace. In the records of a furnace, two of the variable qualities are final results, viz., the quantity of pig iron produced, and its grade; all the other variables are antecedent causes which produce the final results. Of these variable callses which contribute to the quantity of product and its grade, some are partially under the control of the furnace manager, others depend upon the state of the atmosphere, and still others are accidental. Of the first class are the character of the materials charged, and their relative proportions, their mechanical conditions, as wetor dry, large or small pieces, etc., the revolutions of the engine, the heat and pressure of blast, and the distribution of the charge. Of the second class are the number of pounds of air blown into the furnace for each stroke of the engine, the amount of moisture in the air, and the condition of the chimney draughts. These vary largely with the season of the year. Of the third class is the leaking of a tuyere or a temporary stoppage. To keep a record of this kind for a whole year, it is only necessary to procure a sheet of common profile or cross section paper, about 38 inches long and 8 or 10 inches wide, with cross lines ruled about 0.1 of an inch apart. The vertical lines, or those in direction of the

Jan 1, 1879

By Pramod C. Thakur

Introduction Methane is contained under pressure within the fractures and adsorbed on the surface of the coal seams and adjacent strata. It is released into the mine atmosphere during mining of the seam creating explosive gas-air mixtures. Large volumes of air, sometimes as high as 20 tons of air for each ton of coal mined, are circulated constantly to dilute and carry away methane from coal mines. Removal of methane from solid coal, wherever high methane emissions occur, has been discussed by Spindler and Poundstone (1960) and Thakur and Davis (1977). Methane control in longwall gobs is much more complicated. As longwall mining progresses, the overburden continually subsides, and the immediate roof caves behind the roof supports. Even the underlying strata heave and open up communication channels feeding gas to the longwall gobs. The pressure of gas in the underlying and overlying strata is usually higher than the ambient air pressure in the mine. Thus the mine workings become natural pressure sinks, into which gas flows from the entire disturbed zone. Winter (1975) defines this zone as the "gas emission space". Depending on the number of gas-bearing zones in the gas emission space and their gas contents, the total methane emission from longwall gobs could vary from a few to more than one hundred cubic meters of gas per ton of mined coal. Hence the ventilation of longwall faces demands a large quantity of air. This ventilation need is further enhanced by high air losses specially on caving faces (with no stowing of the gobs). The old system of longwall ventilation, the 'U' pattern, where all the air was brought down one gate road and exited through the other gate road, loses large quantities of air through gobs giving a high methane concentration at the return end of the face. The 'Y' ventilation pattern where intake air is brought down both head and tail gate roads relieves the situation to some extent but, on most longwalls, some kind of methane control is still needed. With the onset of longwall mining in Europe several decades ago, the initial work on methane control on longwall has naturally been done there. The basic principle of methane control has always been, and still remains, some means of "bypassing methane from the gas emission space without letting it mix with the mine air". In most cases the bypassing mechanism is either a strategically located borehole or a roadway. A successful methane control program depends on the following basic premises: a. Determination of the geometry of the gas emission space and location of gas bearing horizons therein. b. Estimation of the rate of methane influx into the longwall gob. c. A scheme to bypass the gas in the most economic and efficient manner and thus prevent it from entering the mine atmosphere. In successful methane drainage programs a high proportion, usually between 50 and 70 percent, of the total gas emissions in a working district is removed before it can enter the mine airways. Advantages of methane drainage can be summarized as: 1. Generally reduced methane levels in the mine leading to increased safety and productivity. 2. Reduced air requirements and corresponding savings in ventilation horsepower. 3. Faster rate of advance for development headings and economy in the size of headings. 4. Possible use of mine gas as an additional source of fuel. The purpose of this paper is to briefly review the geometry of gas emission space and methane influx therein and discuss more successful European and U.S. longwall gob gas control techniques.

Jan 1, 1981

By William Adamson

Precise and data driven methods of predicting and modelling blast induced vibration intensity are created and applied under field conditions on a routine basis today. These models are fed with controlled field measurements of vibration pulses generated by controlled charges under known geometric conditions. With such models blasting and geomechanical engineers can estimate vibration intensity at established critical points in mine bench geometry (surface mining) as well as within semi-permanent crown and rib pillars (in underground mining).Damage to bench walls and destabilization of large-scale geomechanical structures can be partially attributed to high intensity blast induced stresses on the rock mass and structures that define important features such as safety catch berms and haul roads. In the underground environment, weakening of critical pillars in block caving or Long Hole Open Stoping(LHOS) operations may lead to ore sterilisation. Blast vibration intensity can be linked with physical alteration in critical rock mass volumes within themine. Often the potential for vibration damage is estimated by the calculation of a critical value of vibration intensity, derived from a consideration of laboratory measured rock matrix properties. Efforts to confirm the accuracy of these estimates of critical vibration level include post blast excavation photography as well as the visual examination of rock mass, pre-and post-blast, using video camera observation of the walls of purpose drilled observation holes. These methods clearly indicate a change in fracture state due to the blast induced vibration, however further modelling of geomechanical stability will benefit if seismic measurement data can be interpreted in terms of geomechanically relevant strength parameters. This paper describes some of the results of previously unpublished research studies conducted in underground operations in Chile and Australia, where cross-hole seismic measurement profiles were interpreted in terms of dynamic modulus. This work dates from a period between 1987 and 1990 and the Author suggests that 30 years on, applying modern day technology, instrumentation and modelling methods, more meaningful, practical and applicable results may be obtained.

By J Czerwinski, M Kasper, M Wyser, N Heeb

The VERT project 1993 - 1998 laid the technical foundation for curtailing the emission of fine particles from diesel engines at Swiss underground workplaces. The targeted maximum was 100 ¦g/m3 EC. Extensive laboratory and field investigations demonstrated that this specification is only attainable with modern diesel particulate filter (DPF) systems that almost eliminate particles in the entire lung penetrating size range from 20 nm to 2.5 ¦m. In January 2000, Switzerland mandated compulsory particle filters in underground workplaces, particularly in tunnelling, irrespective of engine age, size and type. After the implementation, the particle concentration in the respiratory air was below 40 ¦g/m3 EC, compared to earlier values of up to 750 ¦g/m3 EC. Deployed DPFs have proven filtration rates of 99 per cent in the entire size range of alveoli intruding particles. The DPFs are as durable as the engine, suffer no ageing and require minimal maintenance. The failure rate is below three per cent annually. Of the 6000 retrofitted construction machines in Switzerland, about 400 are deployed underground. Prerequisites for this quality are a sophisticated certification procedure, electronic monitoring of DPF operation and meticulous periodic exhaust gas inspection of the engines.

Jan 1, 2005

By G. Mucsi

This paper deals with the laboratory investigation of cathode ray tube (CRT) glass based glass foam, a relatively new product so called ?Geofil bubbles?. The main process engineering material properties of raw materials, such as particle size distribution, moisture content, density and specific surface area are shown. Then the preparation of raw cathode ray tube glass waste is presented including the following steps: crushing in impact crusher, grinding in laboratory scale ball mill, mixing, pelleting and sintering. We found that it was possible to produce foam glass (with proper mechanical stability and particle density) from the investigated waste CRT glass. The material characteristics of the final product strongly depend on the sintering conditions. Optimum conditions were determined particle size fraction was found to be 4-6 mm, temperature 800 0C, residence time 7.5 min. Keywords: waste CRT glass, grinding, pelleting, sintering, glass foam

Sep 1, 2012

The prospect of utilisation of phosphate ores which contain high P2O5 values associated with Al and Fe has produced a great deal of interest over many years. Processing options have generally been based on the utilisation of by-products which were required to offset the processing costs.   This paper presents a range of processing options which have been investigated at laboratory or technical scale on phosphate ores which predominantly consist of crandallite, CaAl3(PO4)2(OH)5.3H2O and millisite, Ca(NaK)Al6(PO4)4(OH)9.3H2O. Christmas Island (Indian Ocean) and Mt Weld (Western Australia) phosphate ores are good examples of such feedstocks. It is expected that these processes may also be applicable to such phosphate ores from other locations.

Jan 1, 1990

By W Ludlow, C Harrington

This paper presents an overview of the design and construction challenges to implement capping systems on Tailings Storage Facility’s (TSF) based on recent projects in Australia. Key aspects discussed include the cover design and barrier system to encapsulate the tailings, designing for a service life of 1,000 years, water management considerations. construction on tailings, vegetation selection and agronomy of the available capping soils plus the Health, Safety and Enivornmental challenges faced in working on tailings. Lessons learned from experience in capping TSFs are also presented. CITATION:Ludlow, W and Harrington, C, 2018. Experience with Capping of Tailings Storage Facilities for Decommissioning Projects in Australia, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 157–165 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By Robert Bittner

The following is a report on the development of two unique and separate cofferdam systems used in the construction of the marine foundations for the I-205 Columbia River Bridge at Portland, Oregon for the Oregon Department of Transportation. The first of these two systems utilized interlocking H-pile 100 ft long to form a cofferdam 145 ft by 58 ft in plan and 100 ft deep. The second system consisted of two immense bell shaped steel forms. The larger of the two forms was 83 ft by 66 ft in plan, 62 ft high and weighed 450 tons. This form was externally stiffened for a designed hydrostatic head of 54 ft. Both forms were designed, except for the top 14 ft, to be set and stripped in one solidly welded unit. The supporting geological formation for the bridge foundations is the Troutdale, a cemented sand and gravel. This formation is at the ground surface on the Washington Shore and slopes downward to a depth of 115 ft below normal high water at about mid-channel. In this area the Troutdale formation is covered with a layer of fine to medium sand 90 to 100 ft thick. It was the variable depth of the Troutdale which determined the two types of foundations and the corresponding to the two cofferdam systems.

Jan 1, 2013

By J. Dubbert, T. H. Jones

"This paper describes a rib-control methodology that uses standing support to provide most of the lateral resistance to the rib. The method is analyzed, and its boundary conditions for rib control are described. In this method, an animal-fencing mesh product covers 80 percent of the rib surface and is sparsely bolted to the rib. Can supports or other similar stiff supports are installed primarily as roof support for the entry. As the rib spalls and dilates, the mesh catches the material and expands, pressing against the standing support. Mining proceeds, and vertical roof loads increase the normal force on the Can supports, causing friction at the top and bottom of the support to go up, thereby improving lateral resistance to rib swelling.At one mine in the western United States, this method has proved excellent at preventing the buildup of rib spall. It substantially reduces the number of bolts used for screen installation, improves installation efficiency, reduces manual labor and provides more rib coverage than many screening methods. Because standing support is typically part of the roof-control plan in longwall tailgates, no additional standing-support costs are incurred. IntroductionRib safety in underground coal mines is an oft-studied topic because ribs are high-hazard areas. An overview of early rib safety research can be found in Jones and Mohammed (2014). There is a significant difference between rib support and rib control. Support improves safety through the installation of strengthening elements, such as bolts and dowels, designed to reduce rib spalling. Although spall reduction is preferred, controlling how the rib fails is enough to ensure safety in some situations. Rib control uses external elements to provide confinement, thereby holding sloughed material in place, or to prevent falling rib material from striking mine personnel. This paper presents a study of a control method used on the pillar side of a longwall coal mine tailgate in the western United States."

Jan 1, 2015