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By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores such as cobalt, platinum, gold, silver, etc. Froth flotation is a physico-chemical process for separating finely ground minerals from their associated gangue material. The process involves chemical treatment of a finely divided (ground) ore in a water pulp to create conditions favorable for the attachment of certain of the mineral particles to air bubbles. The air bubbles then carry the selected minerals, called valuable, to the surface of the pulp to form a stabilized froth which is removed and recovered. The unattached gangue material remains submerged in the pulp and is either discarded or reprocessed. To obtain the adherence of the desired mineral particles to the air bubbles, at least two specific steps must occur: a hydrophobic (water hating) surface film must be formed on the particles to be floated along with a hydrophillic (water wetting) film on all other particles; and a controlled bubble surface tension interface must be maintained, allowing for high particlelbubble collision frequency and efficient attachment or sticking of the particle to the bubble once collision has taken place. In most flotation applications, the above two steps are controlled by chemical flotation reagents. The collector is a chemical reagent which produces the hydrophobic film on the valuable mineral particle and is the primary driving force that initiates the flotation process. The frother is a chemical reagent which influences the collision frequency and attachment efficiency of hydrophobic particles and air bubbles. Thousands of research papers and books have been published on the chemical theory behind sulfide mineral collectors, e.g., Fuerstenau (1962), Fuerstenau (1976), Fuerstenau, et. al. (1985), King (1982), Leja (1982) and Moudgil and Somasundaran (1987), This article will only address and summarize some of the more practical aspects of collector usage. The industrial scale practice of froth flotation in sulfide mineral concentration has changed little since the 1950's. For example, of the approximately 80,000 metric tons of sulfide mineral collectors used commercially (1980) in the free world, almost 98% were known and manufactured in some form 25 years ago. A very interesting and informative history of collector development has been given by Crozier (1984). In addition, the industrial scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully at an industrial scale for many years. As a result of the above environment which makes reagent cost/performance analysis difficult for new reagents, there is a strong tendency for the flotation operation to use, and reagent companies to supply, as cheap as possible raw materials and manufactured products that are quite general in application. In the last 20 years or so, there has been increasing technical effort to tailor make reagents for specific applications, but to date such work has had little commercial impact. It is clearly the hope of this author and others that this situation will change in future years as technology improves and pressure for improved flotation performance intensifies. This article is a condensation of collector usage trends quantified as part of a comprehensive industrially oriented applications program on froth flotation organized by Klimpel and coworkers (1979-1987) and as reported in various countries that participated in the program from 1978-1983. No attempt will be made to provide specific.

Jan 1, 1986

By Robert Mermelstein, Robert W. Kilpper

In a long term toxicology study SPF, F-344 rats were exposed 6 hours/day, 5 days/week for up to 24-months to a special test toner at 0, 1 (low), 4 (medium), and 16 (high) mg/m3, or Ti02 at 5 mg/m3, or Si02 at 1 mg/m3 by the inhalation route. The test toner was composed of about 90% random, high molecular weight acrylic copolymer and 10% high purity, medium color, furnace type, carbon black. It was enriched in fine particles such that the respirable aerosol concentrations (ACGIH) were to 0, 0.35, 1.4 and 5.6 mg/m3. There were no effects of exposure on survival, causes of death, body weight, food consumption or clinical chemistry and hematology parameters. There was no evidence for systemic or upper respiratory system toxicity as all of the changes were restricted to the lungs and associated lymph nodes. The lung tumor frequencies in the control (air and Ti02) and toner exposed groups were comparable. There were no changes at the low exposure level, which is the most relevant in regard to potential human exposures. The lung changes at the high exposure level consisted of decreased alveolar clearance, increased toner retention, increased lung weight, chronic inflammation, decreased pulmonary function and a moderate degree of focal fibrosis in all animals. The changes in the middle exposure level were smaller in magnitude than at the high level, and a very slight degree of fibrosis was found in 20% of the animals. In a follow-up investigation, female F-344 rats were exposed to one of three toner concentrations; control, low (10 mg/m3) or high (40 mg/m3) for 13 weeks followed by five, 13 week observation intervals. At the low exposure level, slight clearance retardation and modest inflammatory changes were observed at the end of exposure. They returned to normal values within six months post-exposure. In contrast, at the high exposure level, increased lung weight, decreased alveolar clearance and a disproportionate increase in toner retention, as well as an inflammatory response, were noted, which persisted throughout the post-exposure interval. While no significant histopathological changes were noted at the end of exposure, a very slight degree of fibrosis was observed in all of the animals in the high exposure group, after 65 weeks of observation. All of the above alterations are ascribable to "lung overloading", a generic response to excessive amounts of any dust retained in the lungs for a prolonged time interval. Published accounts in the toxicology literature document that long-term inhalation of large quantities of even innocuous particles such as Ti02, carbon black, petroleum coke, toner, and other polymeric and insoluble materials has resulted in a spectrum of generically similar pulmonary changes. Scaling of the rat-data, using 1 mg dust/g of lung as a threshold for "lung overloading", normal ventilatory parameters and 300-500 day human retention halftimes project "lung overloading" upon long-term exposure to insoluble respirable dust above 0.2-0.33 mg/m3. Other types of information are consistent with an exposure limit in the range of 0.4-1.0 mg/m3 for respirable dust. The absence of definitive toxicology data in support of the current OSHA respirable nuisance dust limit, coupled with the lack of sound epidemiology data, indicate the need to reexamine the validity of the current nuisance dust health standard.

Jan 1, 1991

By A. Iordanidis

Mining activities have recently been taken place in Vourinos mountain, northern Greece, exploiting the so-called ?olivinite?. A representative sample was collected from this mining area. Mineralogical analysis was conducted using X-Ray powder Diffraction (XRD). A thin section was prepared and analyzed under optical microscope. X-Ray Fluorescence (XRF) was employed for the chemical analyses of major, mi-nor and trace elements. Environmental Scanning Electron Microscopy coupled with energy dispersive x-ray analysis system (ESEM-EDX) was also applied for the characterization of the sample. Forsterite (Mg-rich olivine) and enstatite (Mg-rich orthopyroxene) were the main mineral phases determined by XRD, with a 56% and 44% content respectively. Trace amounts of chromite spinels were also determined. Based on this characterization, the industrial rock proved to be a hartzburgite (a type of peridotite in petrological terms). Optical microscopy and ESEM-EDX provided similar observations and confirmed the above-mentioned results. XRF revealed a composition of 43.31 MgO, 44.25 SiO2, 9.98 Fe2O3, 1.40 CaO, 0.67 Al2O3, 0.22 MnO, 0.01 TiO2 and negligible amount of LOI (<0.1%). Trace elements with high concentrations were Cr (2678 ppm) and Ni (2044 ppm). The comprehensive characterization of this mining product suggested the suitability for various uses, such as refractories, foundry and slag conditioner. Additional experiments are being carried out for environmental applications, e.g. wastewater treatment.

Jan 1, 2006

By S. E. Phillipson

Underground mine mapping is a critical component of understanding the controls on ground instability. Many relatively localized examples of ground failure are associated with geologic features, such as faults, joints, compaction zones, or sedimentological transition zones. In contrast, long-running, directional roof falls are associated with regionally directed horizontal stress fields, which are induced by tectonic plate interactions in the earth?s crust. When stress-related features are distinguished from geological features, underground mapping can be an effective method for inferring the direction of horizontal stress at individual mines. Comparisons with data compiled in the World Stress Map confirm that underground mapping can be a valid method for determining the orientation of maximum principal horizontal stress at a variety of scales, ranging from individual mines to an entire mining district or coal basin. Comparisons are presented between the maximum horizontal stress directions inferred by MSHA?s Roof Control Division through underground mapping and the azimuths of maximum horizontal stress in the World Stress Map database for North America?s Pennsylvanian-aged Appalachian, Illinois, Black Warrior, and Arkoma Basin coal fields. The comparison indicates that azimuths are similar for both data sets, even where variations in the direction of horizontal stress are present. It is interpreted that inconsistencies in direction are attributable to near-surface affects such as deeply incised stream valleys, or to fault zones that represent heterogeneities in the crust.

Jan 1, 2009

By Howard G. Coopersmith

Kelsey Lake kimberlites are a cluster of eight diamondiferous pipes that occur in the northern State Line kimber lite district of the Colorado-Wyo¬ming diamond province (Coopersmith 1991, 1993) (Fig. 1). This is located at the southern margin of the Wyoming Archean province. Two relatively large pipes and a few smaller bodies define the cluster (Fig. 2). The Kelsey Lake kimberlite cluster has been developed as the first diamond mine in the United States. The cluster was discovered in 1987. Full-scale production began in the spring of 1996. Mining has produced gem-quality diamonds up to 28.3 carats in size. The 30 State Line kimberlites, intruded about 390 million years ago into Proterozoic crystalline rocks, were presumed to be underlain by Archean basement of the Wyoming Province. Nearly all the kimberlite pipes contain diamond but only the Kelsey Lake pipes are economic. The Kelsey Lake kimberlites contain predominately diatreme facies kimberlite, with minor hypabyssal and crater facies rocks locally occurring. Alteration and weathering of the kimberlites is extensive. Petrologic studies indicated an underlying lithosphere that has a largely Proterozoic signature and shows evidence of deep melt and metasomatic events. Current development has concentrated on the two largest pipes. These have been bulk-sampled after geological study (1987) identified the presence of small diamonds and mineral chemistry indicating favorable diamond conditions. A 10-t/hr (11-stph) rotary pan plant was established in 1992 to process thousands of tons of bulk-sample material. Grades and gem-quality diamonds to 14.2 carats in size were recovered. Early in 1995, it was decided to move from evaluation to commercial mining. Production began in early 1996. A 195-t/hr (215-stph) rotary pan plant was built. A water supply system and tailings disposal have also been incorporated. The project straddles the Colorado/Wyoming border. Existing mining permits were amended to incorporate the increased scale and new plant facility. The permits for this project included the following: Colorado mining permit, Wyoming dozing permit, Corps of Engineers Section 404, Colorado NPDES, Wyoming NPDES, Storm water Management Plan, Air Pollution Emission Notice, various Colorado and Wyoming State Engineer and well permits, Substitute Water Sup¬ply Plan, Larimer County Special Use Review, and US Forest Service road use permits. The temporary destruction of a wetlands area was a sensitive issue. It has been addressed through the Clean Water Act Section 404 permit administered by the US Corps of Engineers. This use will be mitigated by creating new, self-sustaining wetlands in the immediate drainage system. A comprehensive and bonded reclamation plan is incorporated in the permits. Diamond mining and ore processing is environmentally benign. No chemicals are used in processing the diamonds. The plant is a water wash facility. The kimberlite ore is inert, as is the granite wall rock. The exposure of the ore body and the subsequent ore processing and waste rock and tailings containments pose no threat of release, leaching or dispersion of any detrimental elements. There is no acid-forming waste. Even a prolonged acid rain environment will not produce a harmful leachate. Slimes containment and clarification ponds will prevent unacceptable levels of suspended solids in water discharge. The isolated, restricted access and nonvisible nature of the mine and processing has led to straightforward permitting. Geology The State Line district is located in the northern

Jan 1, 1997

By A. K. Mukhopadhyay, B. Sikdar, V. P. Arora

Introduction During the past several years, persistent concerns about an economic slowdown in India have tempered capital-expan¬sion plans. This has encouraged companies to find methods of increasing the capacity of existing production equipment. One of the more cost-effective methods of increasing produc¬tion capacity without major capital investment is to improve equipment maintenance to prevent needless machine down¬time. This can have a significant financial impact because of reduced maintenance costs and increased machine availabil¬ity. The maintenance program employs different techniques, but the analysis of the vibration behavior of machine compo¬nents to assess their condition was found to be the most accurate technique for monitoring various machine disorders. A dragline that handles overburden in open pit mines is one of the more capital-intensive pieces of production equip¬ment. It is imperative that the machine be run at its optimum capacity under the prevailing environmental conditions. Condition monitoring of the various components of the dragline through vibration analysis not only improves the machines availability but also ensures a long operation time free from sudden failure. This paper describes the application of vibration monitor¬ing and analysis techniques to assess the components of a W2000 dragline. The paper reveals that condition monitoring through vibration analysis can be as valuable in failure prevention as it is in failure analysis. Finally, the paper presents an approach to a knowledge-based expert system to the quick and correct diagnosis of mechanical faults in a dragline. Equipment A dragline is an intermittent-discharge-type excavator having a long boom at one end. Unlike other excavators, the bucket is not rigidly held by the frame structure. In favorable geomining conditions, the intensive operation of these exca¬vators is the least-expensive means of overburden removal. The basic working elements of this equipment are the hoist system, the crowd and retract system, the wing mechanism and the propel mechanism. The walking dragline is accepted as the most economical machine for stripping overburden at depths of up to 60 m (200 ft). They are electrically powered and the various motions are effected by individual drives. The Ward-Leonard system is used to provide the optimum flexibility in controlling different dc-drive motors. To mini¬mize the cost of mineral extraction, the largest such machine in India has a 24-m3 (850-cu ft) bucket with a boom length of 96 m (315 ft). Draglines are very costly, and considerable financial penal¬ties may arise from sudden stoppages due to component fail¬ures. Thus, operational reliability is of the greatest importance. Active measures are now being taken to minimize downtime by adopting condition-based maintenance techniques. Though a number of techniques come under this scheme, it has been observed that assessing the health of the dragline by measuring vibration is the best method of determining the dynamic behavior of its various rotating components to avoid any catastrophic failure. Based on criticality, the vibration-behav¬ior of the following components are studied: • motor-generator unit, • exciter unit, • hoist motor-gearbox unit, • hoist drum-gearbox unit, • drag motor-gearbox unit, • drag drum-gearbox unit and • swing motor unit The approach Increases in the vibration levels in different subsystems of the dragline were observed due to the presence of faults such as unbalance, misalignment, looseness and bad bearings. A vibration behavior study included the measurement of vibra¬tion level, trend analysis of time-domain vibration signals and vibration frequency analysis at bearing points of the components. A vibration meter, a vibration analyzer and a data collector are used. Vibration-level measurement. When the dragline was assembled and put into operation for the first time, vibration data were collected to generate baseline data for the equip¬ment (based on International Standard ISO 2372). After¬wards, the vibration levels were recorded regularly using a vibration meter. The measurements were taken at all bearing

Jan 1, 2000

By John C. Draper

[Duquesne Light Company&apos;s three power plants in western Pennsylvania are now receiving high BTU clean coal from a new coal preparation plant. When burning coal from this plant, particulate emissions are greatly reduced and the sulfur content of stack ace descend. The cleaner coal also leads to more effecting overall operating of the power plants because of less bulk handling of fuel and ash. Environmental improvement is an important benefit of the cleaner, more complete combustion.]

Jan 1, 1976

By J. A. Delgadillo, C. A. Pérez-Alonso

"Tumbling mills using balls as grinding media are used extensively in the mining and cement industries to produce fine powders; however, it is well known that the process of size reduction is highly energy-intensive. For this reason, much of today’s comminution research is aimed at understanding grinding mechanisms, by estimating power draft and modeling milling circuits using the discrete element method (DEM). DEM predicts individual particle trajectories, the distribution of energy and contact forces at each collision, lifter wear and power draw. This paper presents a prediction of the discharge particle size distribution using three components, mass balance model (sometimes referred to as a population balance model, PBM), impact energy distribution of the mill obtained from the simulation of the charge motion using DEM and breakage characteristics of particles determined from drop-ball tests. DEM simulation of the charge of a tumbling mill was done using the code MinProSim 1.0 to obtain impact spectra at different operative conditions. Then, the population balance model (PBM), based on impact energy distribution, was used to predict the final particle size distribution in a batch mill and compared with experimental data. It is demonstrated that the DEM-PBM approach allows a reliable solution to predict particle size distribution for tumbling mills. Ideally, the model has to predict size distribution for any operational condition and feed characteristic; however, it has been shown that predictions include some degree of error.IntroductionThe discrete element method (DEM) has recently become a powerful numerical tool that can be used to investigate internal mill charge in ball mills. The early work of Cundall and Strack (1979) used DEM to analyze soil as a discrete medium. Later, the method was improved to simulate the operation of tumbling mills and optimize the design in order to increase the efficiency of energy utilization. Also, it is of great interest to relate ore breakage based on impact energy spectra obtained by DEM simulations and the breakage distribution function used in the PBM to predict particle size distribution in tumbling mills.King and Borgeous (1993) broke single particles in an ultra fast impact load cell to compute the phenomenological parameters in a microscopically more detailed population balance model. Kapur et al. (1997) simulated the size spectra of progeny particles produced by impact grinding as a function of the impact energy. Narayanan (1987), Narayanan and Whiten (1988) and Lira and Kavetsky (1990) estimated the energy-dependent breakage distribution function for scaleup and modeling of industrial ball mills."

Jan 1, 2015

By Neall. George M.

No method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions experienced in U.S. coal mines. However, a relatively new innovation, the roof truss, has proven itself capable of preventing roof failures which both conventional and resin bolting could not. Presently, the economics of roof trusses are not favorable when compared to bolting due primarily to drilling and installation procedures, but this is expected to change with improved truss and equipment designs. To realize the full potential of this new method of roof control, design criteria are needed which can optimize the effect of the truss for varying strata conditions. A review of the literature reveals that some work has been conducted on the roof truss. 1 Gambrell and Haynes (1970) showed that the roof truss is superior to angled roofbolts installed through a header. Sheorey, Verma, and singh2 (1973) used statics to prove that the optimum angle of inclination should be greater than 450? However, to date, no systematic attempt has been made to establish design criteria. In this report, the problem was approached in three ways. First, model studies were performed to determine general trends of truss support as a function of a number of variables. This provided a better understanding of the support mechanism involved. Secondly, beam theory was employed to calculate the actual support capacity of a truss, realizing the many simplifications, assumptions, and shortcomings of this approach to such a complicated problem. Last, field testing will be utilized to test the hypotheses previously developed and provide hard data to further synthesize roof truss design criteria. Field testing has recently commenced and should be complete before mid 1977.

Jan 1, 1977

By Thomas S. Nye

Mineral exploration in the United States relies heavily on theoretical geologic concepts and indirect methods such as geophysics and geochemical prospecting, as there are few exposed areas of mineralization which have not already been explored at least once. In Mexico ore-grade samples can be cut at the surface on vein, replacement, and disseminated types of deposits in many areas. Here the problem is not in finding a mineralized area but in deciding which mineralized area can be explored and developed at the least risk and cost, and have the greatest profit potential, within the framework of Mexican mineral economics and government policies. Mexico&apos;s min¬eral industry is regulated through a system of permits for various phases of mineral exploitation, and by taxes. This regulation and its effects are not unlike those which may be expected in the United States if present legislative and administrative trends continue.

Jan 1, 1972

By T. M. Yegulalp

Geometric programming is a mathematical optimization technique, first formulated by Zener [1], for dealing with certain types of functions which typically arise in engineering design problems. The theory provides algorithms for finding the optimal solution, existence theorems by which one can show that a particular design is possible or not; and characterization theorems that especially make possible to obtain the relationships among the optimal design parameters at optimum. It has been demonstrated [2] that this technique can be applied to several design problems in chemical, electrical and mechanical engineering. Design problems in mining, metallurgical and mineral engineering have many aspects similar to that of other engineering problems. Thus it is only natural to consider geometric programming as a potential tool for solving certain optimization problems in mineral industries.

Jan 1, 1974

By T. P. Meloy

Understanding the illusive nature of the catastrophic failure of brittle solids is difficult. Proposed is a hypothesis of crack cessation. If, during catastrophic failure, cracks in a brittle solid stop before terminating on a preexisting crack, this hypothesis suggests that the slope of the size distribution curve will be lower. Moreover, this hypothesis suggests that when cracks stop, negative valleys and negative comers appear. Furthermore, particle shapes also change with particle size. While the hypothesis is unverified, it suggests that crack cessation could play a role in the comminution process.

Jan 1, 1996

By Richard A. Thomas, Liming Yuan, Lihong Zhou, James H. Rowland

Hydrogen (H2) gas released during battery charging can result in cross-interference for carbon monoxide (CO) sensors used for early fire detection and compromise the integrity of the mine atmospheric monitoring system (AMS). In this study, a series of laboratory-scale and full-scale experiments were conducted to evaluate the responses of different CO sensors to H2 gas. In the laboratory-scale experiments, constant H2 concentrations in the airflow, from 100 to 500 ppm, pass through sensors. While in the full-scale experiments, increasing H2 concentrations generated as a byproduct from charging the batteries at the battery charging station rise to the sensors under different ventilation scenarios. The H2 concentrations at the CO sensor location were measured using H2 sensors and were correlated with the CO sensor response. The effects of ventilation and sensor location on the CO sensors responses were also analyzed. The results of this study can help mining companies to select appropriate CO sensors and improve the deployment of these sensors to ensure the safeguard of underground miners.

By Richard R. Klimpel

INTRODUCTION The process of froth flotation as a means of upgrading the quality of coal by removing water and/or ash and/or pyrite has been receiving increasing attention since the 1960&apos;s by the world-wide coal industry. Historically, coal preparation practice has involved the use of primarily gravity based separation techniques, screening, and/or water washing in removing the larger fragments of inert material from coarse raw feed coal. The preparation, handling, and cleaning of raw coal finer than 500 micrometers was deliberately avoided. Most often the fine coal that was produced would be judiciously incorporated into the bulk of the larger sized cleaned coal product or simply discarded. Since the 1960&apos;s and accelerating in the 19701s, the desirability of performing some type of coal cleaning or upgrading on the finer raw coal feed materials has been increasing. There are several apparent industrial incentives for implementing fine coal ((500 urn) processing alternatives including the gradually increasing value of the normally lost fine coal product; the additional amounts of fine coal being produced due to changes in mining techniques; the need to appropriately dewater and upgrade fine coal streams in order to meet increasingly stringent environmental regulations on water quality. storage of fines, etc.; the increasing need to lower undesired sulfur and ash contents in some coal end-use applications; and the increasing development of speciality coal use processes that require inherently finer and cleaner coal particles such as in gasification, liquefaction, fuel injection, etc. Looking at the over 70 years of mineral oriented fine particle separation successes at the industrial level, the use of froth flotation in fine coal would also seem to be a natural process to utilize at the industrial level. In the world of strictly mineral processing technology and, more importantly in the world of mineral commodity economics, there are clearly a number of advantages to using froth flotation that are well documented and universally accepted. These include: the relatively low capital and space requirements required when compared to other process options available to upgrade minerals; the extreme flexibility inherent to the froth flotation process both with regard to throughput and differences in raw feed characteristics; the general availability of a wide range of equipment sizes and reliable suppliers of equipment; and the general availability of appropriate and relatively cheap chemicals that are the major driving forces behind mineral flotation. Just how has the implementation of froth flotation fared in coal upgrading over the last 20 years? It is towards answering this question that this paper will be oriented. As such, the discussion will be somewhat philosophical. The comments offered are based on the personal experiences of this author who has worked in the basic side of inventing new flotation chemistry and flotation analysis techniques with the major overriding goal of implementing this new technology successfully at the industrial level in both coal and mineral flotation throughout much of the world. This paper is not intended to be a complete reference to all published scientific work on the subject of industrial froth flotation of coal. As a brief background to what will be expanded upon in this paper, the author would like to offer the following summary comments: 1) the industrial practice of fine coal froth flotation is qualitatively similar but quantitatively different than industrial mineral froth flotation when viewed from either research considerations, operating viewpoints or economic evaluation criteria; 2) it is clearly possible to invent and/or develop new technology (such as new chemistry and/or equipment) that can do statistically better than current plant practice in well-controlled tests in either coal or mineral flotation circuits; 3) there is so much inherent variability to the raw feed coal streams and other operating parameters that even if there is a documented technical incentive to implement some new piece of technology, the people, instrumentation and control methodology required are often simply not available or not economic unless the potential payoff is quite large.

Jan 1, 1988

By R. Poulin, R. Pakalnis, D. Peterson

Introduction A comprehensive project has been completed at the University of British Columbia that focused on the development of a commercially competitive composite cable bolt. Laboratory evaluation consisted of pull tests, preliminary shear tests, grout column testing and scanning electron microscope evaluation of the composite cable bolt. Field implementation at five underground mines throughout Canada were completed for technical and overall performance evaluation of the new product. The project began in 1989 under the sponsorship of HDRK Ltd., a consortium of mining companies including Inco, Noranda and Falconbridge Ltd. The objective was to develop a cuttable cable bolt that would serve as support for a continuous miner within a hardrock environment (Mah, 1994). The bolt was required to be cuttable by a continuous roadheader without adverse effects to equipment, personnel and the milling process. From the beginning, the focus was to employ existing cable bolt installation technology (Goris et al., 1994) whereby the cable bolts are grouted within a 4.8-cm-diam (1.08-in.diam) or larger employing a portland cement grout having a water: cement ratio between 0.35:1 to 0.4:1. The prototype, developed from Phase I, was technically comparable to that of conventional steel cable bolts. However, high cost and limited accessibility hindered its commercial acceptance. The second phase of the project was sponsored by the British Columbia Science Council and Pacific Pultrusions Ltd. of British Columbia to locally develop a glass fiber cable bolt. This phase of the project has resulted in the DAPPAM cable bolt (Pakalnis et al., 1994). DAPPAM Cable bolts are the second most used support method in Canadian underground mines after mechanical rock bolts. It is projected that 870,000 m (2.9 million ft) of cable bolts are installed annually in Canada (Poulin et al, 1994). The objective of the research program was to develop a glass fiber cable bolt with similar pullout strengths to the conventional steel cable bolts (267 kN - 60,000 lbf). More than 200 laboratory pull tests were conducted employing standard cable bolt test procedures, as outlined by the US Bureau of Mines (Goris, 1994), to assess optimum rod and surface composition, geometry, number required, pullout and shear characteristics, among others. This resulted in the DAPPAM cable bolt (Fig. 1). It consists of 10 individual 0.6cm-diam (0.25-in.-diam) rods enveloping a 2.2-cm (0.8-in.) outside diameter, high-density polyvinyl grout tube. The composite used in the construction of the rods is comprised of about 65% glass fiber within a 35% polyester resin matrix (by volume). Through pultrusion (a continuous fabrication technique), the fibers and resin are combined into a composite material. The strength of the overall composite is largely governed by the material properties, the alignment [ ] of the fibers relative to the applied load and the density and interaction of the overall components within the composite. A further modification to the composite was the addition of a sand grit to the surface of the individual rod. This increased the overall bond strength between the rod and the cement-based grout, used to solidify the cable bolt to the rock mass. The grout tube has an inside diameter of 1.9 cm (0.75 in.) and is rated at a pressure of 1.72 MPa (250 psi). During installation, the grout is pumped through the annulus of the grout tube thereby filling the hole from toe to collar. Laboratory and field trials concluded that a minimum drill hole diameter of 4.8 cm (1.08 in.) be used to ensure optimum grout coverage. Table 1 compares the design strengths (as tested) for DAPPAM vs. the ultimate strengths for a conventional 1.6cm (0.6-in.) steel cable bolt. The characteristics for the steel cable have been derived directly or indirectly by Goris et al. (1994). A design tensile strength of 289 kN (65,000 lbf) was proposed for the composite developed. However, recent tests have indicated maximum breaking strengths of more than 400 kN (90,000 lbf) (Fig. 2). Figure 2 shows the load vs. displacement pull out curves for steel and DAPPAM cable bolts. The methods of testing were standardized so as to be able to compare the results directly. The shear strength of the fiberglass cable is largely a function of the surface area sheared. Shear strengths of 89 kN (20,045 lbf) were obtained for the configuration shown in Fig. 1 when sheared normal to the axis of the cable. The testing procedure was similar to the shear tests performed on steel by Goris (1993). It is important to reinforce the implication that critical bond strength has on the success of an overall support system.

Jan 1, 1995

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By Felicia F. Peng

In operation of column flotation of fine coal, the smaller bubbles are required to effectively recovery of fine hydrophobic particles. However, smaller size bubbles alone does not ensure an efficient flotation for fine particles. The sufficient available bubble surface area per unit column are essential for effective column flotation. The experiments are conducted in a lab-scale column using two different bubble size distributions and three different particle size distributions for flotation of Pittsburgh No.8 seam coal. The interaction between particle size, aeration rate, feed rate and wash water addition are examined in this investigation to establish the optimal operation conditions for flotation of fine coal. Despite of wide range of operation conditions used in this study, it is found that the experimental data fall in a single characteristic recovery-ash curve for a given bubble size distribution and particle size distribution. The major parameters affect the flotation performance points along -the characteristic recovery-ash curves are aeration rate/frother dosage, feed rate and wash water rate-.

Jan 1, 1994

By Petr Salak, Adrian St. John, Angelos Gakis

"INTRODUCTIONCrossrail will provide London and the South East of England with a world class, high-capacity railway to ease congestion on London’s public transport system and enable better access between business, residential and entertainment districts. At the heart of this £14.8bn scheme lies Farringdon Station, which will form a major interchange between Crossrail, Thameslink and London Underground. The £400m Farringdon Crossrail station was awarded to a joint venture of BAM Ferrovial Kier (BFK) in 2011, and comprises two ticket halls, two platform tunnels and a multitude of connecting tunnels all constructed using open face sprayed concrete linings (SCL). BFK appointed Dr. Sauer & Partners (DSP) as their specialist SCL designer. As the main contractor for both C300/C410, the Western Running Tunnels and Station Caverns, and also for C435 Farringdon Station, BFK were in a unique position to develop strategies to reduce cost, time and risk for the Client. One such innovation involved the continuation of the TBM drives for the running tunnels right through the length of the new Farringdon platform caverns. This created large diameter precast concrete pilot tunnels for the subsequent platform enlargement using SCL. The benefits of such an approach are described in detail by St.John et al 2015. BFK also introduced another alternative strategy to reduce cost, time and risk, by proposing to abandon the TBMs in the ground (“Turn & Bury”) permanently rather than recovering them through the ticket hall shafts. This de-risked the programme, enabling critical ticket hall construction to proceed unhindered. However, one consequence of this revised strategy was an area of increased underground congestion, with two SCL tunnels (RTE2 and ES2) and a bored tunnel (complete with abandoned TBM) in close proximity to a ticket hall excavation (see Figure 1) and a sensitive, listed railway structure. Faced with this complex geometry, and combined with challenging geotechnical conditions (Gakis et al 2015), the team chose to install enhanced monitoring to study the behavior of tunnel RTE2 and provide additional assurance that the tunnels were performing as predicted. Applying advanced 3D numerical modelling and taking into account the residual tensile strength of the SCL helped avoid unnecessary reinforcement installation in primary lining and/or early installation of a secondary lining in RTE2 prior to ES2 construction."

Jan 1, 2016

By Carlon C. Chambers

INTRODUCTION Manufacturer sponsored research over the past three years has proven the suitability of styrene-butadiene/ asphalt membranes in many waste disposal seepage control applications for the mining industry in general and uranium mining and milling in particular. The class of membranes that have been tested are SBR/asphalt blends containing up to 68 weight percent asphalt, the balance being elastomeric polymers, fillers and phase compatibility promoters. These membranes are classified as thermoplastic, elastomeric materials and can be hot spray applied to any moderately dry substrate at 177°C (350°F), and placed in service within one hour of application. The properties of this membrane that make it suitable for use as an engineering_ material in these applications are discussed. These properties include chemical resistance, bacterial resistance, UV light resistance, permeability, stress/strain characteristics and bonding data. The favorable economics of the system are also discussed. The preliminary raw test data on the above properties were reported in a previous paper, (Chambers, 1980) and most will not be repeated here. Instead, this report considers the original data in relation to subsequent data and concentrates on interpretation of the trends for engineering purposes. However, several of the pictures contained in the original paper are reproduced in this report for illustration purposes. BACKGROUND Asphalt alone has a long history of use as a water-stop material. Its limitations of becoming brittle when cold and soft when hot, however, are well recognized. The service temperature for common grades of asphalt is usually in the 4° to 49°C (40° to 120° F) range. Modifications to the base asphalt can generally improve its performance at the high end of the temperature range by raising its softening point, but only the addition of one class of modifiers, elastomeric polymers, improves low temperature flexibility as well and also adds resilience. Both of these properties are of extreme importance if the waterstop material is to maintain its integrity when used on substrates which are unstable, non-cohesive, discontinuous or have significant expansion/constraction coefficients. Although there are many elastomeric polymers which can be used with

Jan 1, 1980

By Haydn H. Murray, Sam M. Pickering

The term kaolin is used here, in a commercial sense, for near- white clay deposits which are dominantly comprised of the mineral kaolinite, or for products refined from such deposits. The term "china clay" has been used in the past as a synonym for kaolin, a reference to its first being mined some 3000 years ago at a hill near Jauchau Fu, Jiangxi Province, China as a porcelain raw material. The name kaolin was derived from the Chinese Kauling, which means high ridge. For most modem industrial applications kaolin must be extensively refined and processed from the crude state to enhance whiteness, purity, and other important commercial characteristics. After such processing, kaolin ranks as one of the most valuable of industrial clays, and it is used to some extent in most manufactured consumer products. In 1991, the United States alone produced kaolin valued at $1,072,366,000. It is applied very widely as a pigment, filler, coater, extender, ceramic raw material, catalyst base, electrical insulator, and pharmaceutical. Some more prominent uses include paper filling and coating; paint, plastic, adhesive, and ink pigment; rubber reinforcing agent; ceramic raw material for porcelain, dinnerware, tile, and enamels; catalyst base for petroleum cracking and auto exhaust emission catalytic control devices; cosmetics base; and digestive coating remedy. Those properties of kaolin clay which make it suited to these applications are related to its physical, optical, mineralogical, and chemical nature.

Jan 1, 1994