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By Verya Nasri, Heather M. Ivory, Kevin Clarke, Eric Cole

Development of the New York City Water Supply System The New York City water supply system serves a population of almost nine million, including residents in neighboring counties. Peak day demands on the system are currently running at about 70 m3/sec (about 1,600 mgd), which is significantly lower than in the 1990s. This trend of reducing demands over the past 25 years, which has been experienced in other major urban centers in the US, is mainly the result of conservation measures and increased cost consciousness on the part of water users. In the long term, demands are expected to once again follow an increasing trend with population growth, so that by the end of the planning horizon of these studies, the year2045, the peak day demand is expected to increase to about 85 m3/sec (2,000 mgd).The system is served largely by surface water sources, regulated by a system of 19reservoirs and three controlled lakes, which are connected to the City by a series of aqueducts, the longest of which is about 87 km (54 miles). The total storage in the sys-tem amounts to about 560 billion gallons, providing a system firm yield of about1300 mgd. The Delaware system provides about 50% of the City’s water, the Catskill about 40%, with the remaining 10% provided by the Croton. The system as a whole, which is shown on Figure 1, has been built up over the past 150 years, and is truly a marvel of engineering. The Catskill and Delaware aqueducts both discharge into Kensico Reservoir, which is a daily balancing reservoir located in Westchester County about 24 km(15 miles) north of the New York City line. The two aqueducts extend south for a further 14 miles to combine again at Hillview Reservoir in the Bronx, which provides hourly balancing storage for New York City system. The Croton aqueduct runs nearby as shown on Figure 2, but does not connect into the two reservoirs. From Hillview Reservoir the water is supplied to the city via three main distribution tunnels, known as City Tunnels 1, 2, and 3. City Tunnel No.1 was developed in the 1920s as part of the Catskill system. City Tunnel No. 2 followed in the 1960s as part of the Delaware system. City Tunnel No. 3 (CT3) is partly under construction and partly in operation, and is being implemented to permit the other distribution tunnels to be taken out of operation for the first time for inspection and maintenance. Presently CT3 is supplied with water from Hillview Reservoir via a large underground valve chamber, the Van

Jan 1, 2007

Research into the mechanisation of ore winning from the very narrow reefs characteristic of South African gold mines has been conducted from as early as 1964 through the Chamber of Mines Research Organisation (COMRO) but with limited success. However, the Council for Scientific and Industrial Research (CSIR) has recently re-evaluated the potential for the mechanisation for the narrow tabular gold mines. It has been determined that there are approximately 22 000 tonnes of gold left in-situ in channel widths of less than 500 mm in thickness. The extraction of ore from such reefs requires a complete paradigm shift in that the current mining cycle (drilling, blasting, cleaning and supporting) is not feasible since such a stoping width would not allow the workers entry into the stope. Consequently, a non-entry form of mining is required. With this target in mind CSIR has invested in a project called the Nederburg Miner which seeks to identify new technologies for drilling, rock breaking, cleaning, and materials handling, the application of which will enable non-entry mining of very narrow ore bodies. The intention is to develop a fully automated system that will not require human involvement at the stope face. The outcomes of research into this proposed system will not only extend the life of South African gold mines but will also enhance South Africa's reputation for excellence in mining research. This paper presents the new technologies under consideration and potential methods for their application into an integrated non-entry mining system.

Jan 1, 2008

By Henk E. Brassinga

In Rotterdam a number of underground parking facilities have been built and are under development in the inner city, and in the suburbs. These parking facilities are located at already densely built areas. The design of these constructions has to take into account the impact of the building activities and that of the final construction on the surroundings. In this paper some examples of the different building methods of newly realized underground parking facilities are described, and problems that occurred during the construction are pointed out. In the soil conditions as present in Rotterdam the cost effectiveness of underground facilities can be improved by deepening the parking facility, and at the same time reducing the surface area so that the total parking volume remains constant. In this way the total volume of the diaphragm walls and the costs of the project is reduced.

Jan 1, 2006

The title of “the world’s biggest volume-produced hydraulic breaker,” that was held for more than a decade by the Atlas HB 7,000, changed hands with the launch of Atlas Copco’s HB 10,000 at bauma 2007. A service weight of 10 t (11 st) and up to 50 percent higher productivity represent major steps toward a new market segment: the use of hydraulic breakers as cost-efficient production machines for primary breaking and as a true alternative to drilling and blasting. The HB 10,000 has all the technical features developed by Atlas Copco for its hydraulic breakers and includes: StartSelect — the breaker’s start-up and shut-down behavior can be adapted to the conditions. AutoControl that ensures an optimized balance of impact energy and impact frequency to maximize percussive performance. Auto-Control also reduces blank firing force and controls energy recovery. PowerAdapt — if oil pressure is too high, the breaker shuts down automatically. This prevents downtime and is particularly useful when breakers are used on different carriers. DustProtector II — a combination of multiple strippers prevents abrasive rock dust and foreign bodies from penetrating the lower breaker section. Conti- Lube II — the automatic lubrication unit guarantees optimal lubrication of the breaker at all times.

Jan 1, 2007

By W Herrmann, R R. Large

Volcanic-hosted massive sulfide (VHMS) deposits in ancient, submarine volcanic belts of Australia include mound, lens and sheet-style Zn-rich polymetallic deposits, massive and disseminated pyritic Cu-Au deposits and disseminated, stratabound Au-only deposits. Variation in morphologies, metal ratios, volcanic environments and alteration features indicates there exists a spectrum of deposits from the classic VHMS model, to hybrids between VHMS-porphyry Cu and VHMS-epithermal end members. Zn-rich, polymetallic deposits form eitheron, or just below, the seafloor, whereas the Cu-Au and Au-only deposits form sub-seafloor by replacement of specific volcanic lithologies. Alteration zoned outwards from quartz -> chlorite -> sericite ¦ carbonate is typical of VHMS deposits across the spectrum. Quartz and carbonate alteration is dominant in the Au-only systems,chlorite alteration is commonly developed close to Cu-rich ores, and sericite and carbonate alteration zones are well developed in the stratiform Zn-rich ores. The most useful geochemical vectors for Zn-rich ores are the alteration index [AI=100(MgO + K2O)/(MgO + K2O + CaO + Na2O)], the chlorite-carbonate-pyrite index [CCPI=100(MgO + FeO)/(MgO + FeO + K2O + Na2O)], Mn content of carbonate, Mg/(Mg+Fe) ratio of chlorite,whole-rock Tl, Sb and Ba/Sr ratio, S34S of pyrite, and whole-rock S18O. The most useful vectors for pyritic Cu-Au ores are AI, CCPI, S/Na2O ratio, Na content of white mica, and Fe content of carbonate.

Jan 1, 2002

By O. Naiker

Xstrata has modified and improved on the solid state reduction of chromite (SRC) process that was pioneered by Showa Denko in Japan and previously used at Xstrata?s Lydenburg plant. The history of pre-reduction is a long one, with the SRC process being the first and only commercially successful one prior to Premus. Krupp developed the chromite direct reduction and the rotary hearth for the pre-reduction of chromite. However, these processes have not been successfully commercialised. The are a number of significant differences between Xstrata?s patented Premus and the original Consolidated Metallurgical Industries (CMI) process. These include, the use of lower cost anthracite fines as the reductant source together with the use of unactivated bentonite as the binding agent. In addition the use of oxygen as an energy source is used in Premus but not in the original CMI process. There was a fundamental change in the operational philosophy of the process in that while the original CMI process aimed to maximize metallization of the pellets, the Premus process seeks to maximize the energy output from the kiln while still achieving the required efficiencies and hence increasing furnace output. Premus is designed to reduce electrical energy consumption, provide high recoveries of metallic oxides, and utilizes low cost reductant and energy sources such as anthracite and oxygen respectively. Due to the fact that Premus is one of the most technologically advanced and competitive processes used today in the production of ferrochrome it was the process selected for Xstrata?s new ferrochrome facility, Project Lion, that is due to be commissioned in the third quarter of 2006 in the Steelpoort Valley in Mpumulanga, South Africa.

Jan 1, 2007

Simulation and Analysis of Media Behaviour in the Tower Mill by Discrete Element Method for Estimating the Power Draw and Spatial Distribution of Grinding Process The trends to the Tower Mill (TM) for grinding the materials are increased recently due to its unique merits. Among the various considerations of this machine, the main important items are increasing the efficiency of grinding as more as possible while keep the power consumption, manufacturing and maintenance costs as less as possible. Because this grinding process includes a huge numbers of media particles (ball or other shapes) that are interacted with each other and body of the machine, the DEM (Discrete Element Method) way is a very useful and essential step to getting confident evaluations. Because the DEM simulation is a very young scientific field and proper commercial software are very rare and still under modification, at first step, a FORTRAN code is developed for DEM simulation of two laboratory Tower Mill models (KM5 and KD5). The simulation and experimental results matched each other properly. Then one another DEM software was implemented to perform simulation of bigger models (huge numbers of balls), meanwhile could handle the complex 3D geometry data. The simulation and experimental results were compared and analysed in detail for not only laboratory but also for industrial models (KW5, KWD5, KWD50, KW300F, KW700F and KW1500) without any manipulation or compromising of the parameters of simulation. Some numeric data of industrial models are hidden here due to confidential reasons. The relative velocity of balls (ball-ball and ball-TM body), dissipate energy, interacted forces and stresses during the process are the key outputs of DEM simulation for design this machine. Therefore the main important processes in a Tower Mill machine are recognised and consequently some horizons of development of this machine are distinguished.

Sep 13, 2010

THE solid nonmetallic inclusions present to some extent in all commercial steels have been variously designated. In early references they were usually called slag inclusions, and this terminology is still retained to some extent. On the premise that all nonmetallic matter produced in a metallurgical process is slag, this can be justified. We have come to think of slag, however, in a more restricted sense, and calling these particles slag inclusions is further objectionable because it fosters the erroneous concept that they are actually incorporated portions of furnace slag. Rarely is this true. The term sonims (coined from solid nonmetallic impurities) was proposed many years ago but never came into common use. In general, the term non-metallic inclusions seems to be satisfactory and adequate. NATURE AND CONS1lTUTION OF NONMETALLIC INCLUSIONS From the standpoint of origin, inclusions fall naturally into two classes: (1) those occurring as a result of reactions taking place in the body of molten or solidifying metal and (2) those resulting from mechanical incorporation of refractories or other materials with which molten steel comes in contact. The former are by far the more numerous and have been called indigenous, natural, or native inclusions. The latter are by their very nature sporadic in occurrence and have been called exogenous, accidental, adventitious, or foreign inclusions. They are often referred to as the "slag-stringer" type, though here also the material is not likely to be furnace slag. Substances Found in Inclusions. Nonmetallic inclusions are composed principally of oxides and sulfides but carbides and nitrides sometimes play an important role. In the case of indigenous inclusions the reactions by which these particles are precipitated from molten or solidifying steel may be caused by: (1) oxidation such as may occur during refining or during tapping and teeming, (2) deoxidation by added alloys, (3) change

Jan 1, 1964

By W. G. Pariseau

Conventional barrier pillars in underground coal mines are intended to protect mains and sub-main entries for relatively long periods of time. Inter-panel barrier pillars that may be used in especially deep coal mines, say, below 800 m (2,400 ft), are intended to protect longwall panel faces and adjacent panel entries for the life of nearby panels. Barrier pillar failures have serious consequences, especially when failure is sudden and catastrophic. Design of barrier pillars is thus of considerable importance to safety and efficiency of operations. Locally, failure occurs at the limit to elasticity when stress tends to exceed strength. Strength of rock is stress dependent, so an analysis of stress is central to rational design of barrier pillars. There are several approaches to stress analysis for coal mine pillar design that range from the simplified but popular tributary area method including variations that use empirical definitions of the tributary area (volume) to more sophisticated numerical methods such as the boundary element method and the finite element method. Despite ease of use, empirical methods are limited and always suspect when applied to conditions beyond their scope of definition, while boundary element methods do not generally include pillars in the solution domain. The finite element method is not so restricted and thus is a natural choice for conventional and inter-panel barrier pillar safety analyses. Caving of roof strata certainly affects the distribution of stress about longwall faces and nearby barrier pillars. Conventional wisdom that considers roof strata to respond as cantilever beams indicates the longer the span the greater are loads on longwall faces and face supports. Failure of roof beams, caving, shortens spans and relieves face and support loads. The same view is applicable to barrier pillars in that long roof spans outside of barrier pillars leads to higher pillar loads than short spans. Hence, caving should be beneficial to barrier pillar safety. However, the benefit may be small depending on the extent of caving and only in the form of a slight reduction in overburden weight transmitted to the barrier pillar, especially in case of inter-panel barrier pillars. FINITE ELEMENT MODEL Finite element models for stress analysis in rock mechanics are formulated in consideration of physical laws, kinematics, and material laws. These models meet the essential requirements for equilibrium of stress, satisfy strain displacements relations, and in a general sense, relate stress to strain through material laws, often referred to as constitutive relations. The last may be complex and replete with nonlinearities. A de facto engineering material law is the elastic model in the form of Hooke?s law for linear, homogeneous materials that may be anisotropic. A limit to elasticity is imposed by strength in the form of a failure criterion or yield function that accommodates triaxial stress states. Examples used in rock mechanics are the well-known Mohr-Coulomb (MC)criterion, the popular Hoek-Brown (HB) criterion, and the Drucker-Prager (DP) criterion. The first is linear in stress, while the second is nonlinear. Both do not include any effect of the intermediate principal stress. The last is also linear in stress but includes the intermediate principal stress. All show an increase in the major principal stress at failure with confining pressure. In the UT2/3 finite element programs, the failure criterion Y is a nonlinear, anisotropic form and is also the plastic potential [1]: [YJI=+ 11NYJI=+- (1) ] where N is an exponent, usually 2, and J2 and I1 are nonlinear anisotropic forms of the second invariant of deviatoric stress and the first invariant of principal stress. The form (1) reduces to DP when N=1and isotropy prevails. When Y = 1, the limit to elasticity is reached and plastic deformation is possible. The material constants that enter (1) are all computed from unconfined compressive, tensile, and, in case of anisotropy, shear strengths. These latter strengths may be estimated from conventional laboratory test data. Of course, joints in field-scale rock masses induce different values. How different is often a matter of conjecture. For comparative studies, laboratory rock properties seem adequate. Inelastic, that is, plastic behavior occurs with strain beyond the elastic limit. Strain hardening, ideal plasticity, and strain softening are possibilities. Ideally brittle behavior with complete loss of strength is an extreme form of strain softening [2]. Caving may also be considered as strain softening with a less drastic reduction of material properties. As intact roof strata fail, fall, and form a rubble pile below, a new material is formed with greatly reduced elastic moduli and strengths, say, 1/100th of the original intact moduli and strengths. From another viewpoint, caving is simply a natural cut and fill process. Input for stress analysis by the finite element method includes: 1. Preexcavation stress state 2. Geology (stratigraphic column) 3. Structure (major faults, joint sets) 4. Elastic moduli 5. Strengths, and specific gravities of each rock type 6. Excavation sequence Output from an analysis includes distributions of: 1. Stress 2. Strain 3. Displacement 4. Local factor of safety (element fs) The last is defined as the ratio of strength to stress and in view of the failure criterion (1), is the ratio [fsJstrengthJstress=] where the numerator is the maximum possible value and the denominator is the actual value of the square root of the second invariant of deviatoric stress. This invariant is a measure of shear stress intensity but also reduces to uniaxial tensile and compressive strengths as the case may be. An fs greater than 1 indicates an elastic state of element stress; a value of 1 indicates the limit to

Jan 1, 2009

By K. Honkivaara, S. Junnikkala, J. Aho, K. Föhr, P. Lampinen

"Bubble area flux has been claimed to be a major tool to dimension flotation equipment. The paper describes laboratory scale (batch and 301 continuous cell using plant feed) and small-scale (1 ,8m3) industrial experiments with coarse-grained massive sulphide ore.Samples were taken with different froth thickness and air addition rates while the bubble size was measured with an UTC capillary device. Problems of averaging the bubble size measurements and the bubble area flux measurements are also discussed.Material balances were calculated using the commercial program BILCO. The flotation rate constants were calculated for pulp and for froth assuming that we can extrapolate the flotation rate constant of the pulp phase from tests using very thin froth thicknesses.The results indicate that the bubble area flux is not alone an adequate dimensioning tool. With fine sizes there was a nearly linear relationship between the bubble area flux and rate constant for constant impeller revolutions. For coarse sizes the form of the relationship is non-linear and seems to go through a maximum value. Thus the final flotation outcome is not a simple function of bubble area flux but depends on the particle size distribution and other more traditional flotation cell dimensioning variables. The effect of the froth residence time in the total recovery of coarse particles is large, but will not be treated in this paper."

Jan 1, 2000

By J. Allaire, P. N. Muteb

The Meadowbank mine is operated by Agnico-Eagle Mines Ltd. and is located in the Kivalliq District of Nunavut in northern Canada. The process plant operates presently at a throughput of 11,500 tpd with gold grade varying between 2.9 and 3.6 g/t. The ore supplied to the process plant come from two different open pits and it expected to come from a new open pit early 2014. The ore is mid-hard materials so the feed size ore have a bigger impact on the throughput. The average gold recovery is 94.0% with about 15-30% of gold recover from the gravity circuit. The Meadowbank mine commissioning started in February 2010 utilizing a flowsheet consisting of one stage of crushing followed by primary and secondary grinding. The commercial production began in March 2010. During the commissioning several difficulties were observed to reach the designed throughput of 8,500 tpd. It was evaluated to insert a second stage of crushing. The insertion of the secondary crushing have two objectives, first to reach easily the process plant design throughput as well as to evaluated the mill capacity increasing. The process plant team evaluated various expansion alternatives eventually deciding to install a secondary crusher, increased the SAG and Ball mills motors capacities and maintain the rest of the process unchanged. Early 2011, with the improvement in the gold markets, decision was made to install the secondary crusher between the primary crusher and the primary grinding. The process plant continued to operate on the initial design until early June 2011. This paper presents the SAG mill throughput increasing from the designed tonnage of 8,500 tpd to 2012 targeted tonnage of 11,500 tpd, secondary crusher commissioning and the difficulties to operate in Canada Artic (Nunavut).

Jan 1, 2013

By S M. C Verryn, K Deysel

In heavy mineral sands ilmenite, rutile and zircon are present in economic concentrations. These minerals are important sources of Fe, Ti and Zr for various industrial uses. The beach sands of northern KwaZulu-Natal, along the Zululand Indian Ocean coastline in South Africa, have been mined for 30 years. The various processes involved in mining and processing these minerals have been controlled by time-consuming and labour-intensive analysis methods which, in some cases, can be replaced by X-ray powder diffraction (XRD)-Rietveld analysis. Heavy minerals are concentrated in the plant to produce a heavy mineral concentrate (HMC). The heavy mineral extraction plant is controlled and informed of the proportions of minerals in the HMC according to their magnetic susceptibility. The different fractions are called magnetic, paramagnetic and non-magnetic fractions and quality control is based on their magnetic fraction weights combined with X-ray fluorescence (XRF) elemental analysis. This current quality control process is very time consuming and labour intensive. We have aimed to show that we can, by XRD, identify and quantify all the minerals in the HMC and also in the magnetic separations. These results are then compared with results obtained by Mineral Liberation Analyser (MLA) as well as with the results from the physical separation. It was possible to detect all minerals present in the material by XRD and use the Rietveld method for a full bulk mineral quantification. A maximum of 15 minerals can be included in the Rietveld refinement due to computational restraints. Major minerals detected are rutile, ilmenite and zircon and minor minerals are quartz, epidote, anatase, amphibole, monazite, magnetite, titanite and augite. The XRD and MLA results correlate relatively well and detailed results will be presented for HMC and magnetic fraction samples. Small differences were observed for some minerals due to the different ways in which minerals are detected or due to preparation methods within each method. Rutile seems to be more abundant in the XRD analysis and the detection of Anatase is possible by XRD, but not MLA. Leucoxene (consisting of mainly microcrystalline anatase, rutile and titanite is not detectable by XRD since it is not a mineral, but ilmenite alteration phases are observed instead. There is, however, a correlation between anatase, titanite and rutile detected by XRD and leucoxene analysed by MLA.

Jan 1, 2008

By Larry Schneider

Are you one of the thousands of hunters, fisherman, or hikers who purchased handheld GPS systems over the past several years and then been disappointed in their accuracy and performance. Or maybe you are one of the hundreds of blasters who have attempted to use GPS as a method for locating the exact position of their blast and measure distances from the blast to nearby structures. If so, you may have found that the accuracy of the handheld GPS systems left much to be desired. It may be time to get the units back out of storage and give them another try. As of midnight, May 1, 2000, the federal government has turned off the “error” built into the system known as “selective availability.” Here is a brief explanation of what that means in a question and answer format.

Jan 1, 2001

By Creelman C. A

With the technique and data obtained in using the blast hole diamond drill for "Removal of Horizontal Level Pillar by means of Blast Hole diamond drilling at New Occidental Gold Mines N.L.," as described in Proceedings No. 133 of 1944, and a series of experiments using the blast hole diamond drill for stoping at New Occidental and New Cobar Mines over the past two years, a benching method of stoping has been developed using 2 inch coring diamond drill bits, resulting in efficient and economical ore extraction in these large ore bodies.During the experimental stages the standard H-inch blast hole was used, but the deflection of the holes, when using this solid bit, either decreased or increased the burden of the hole when it became necessary to drill holes deeper than 45 feet.It was found that after blasting vertical holes deeper than 45 feet, the ore did not break satisfactorily over the full length of the hole; on one occasion as much as 30 feet of bench was left at the bottom of the face, which had to be subsequently re-bored andblasted.As this defect was also noticeable to some extent even when the holes did not deflect, it was then obvious that the explosive had become less sensitive in the lower section of the hole. This was probably brought about by pressure being built up ahead of the detonation wave in H-inch hole, resulting in desensitization of the explosive.Another series of experiments were carried out, using 2-inch solid bits, which enabled a much greater burden on the holes, and the use of primacord for the full length of the hole.This gave satisfactory results regarding breaking and fragmentation, but was uneconomic owing to the excessive consumption of diamonds, and the slow rate of drilling.Further experiments were then carried out using 2-inch coring bits, (Fig. 1) which allowed holes of 90-feet to be bored with little deflection, and proved to be entirely satisfactory as to cost and speed of drilling.

Jan 1, 1945

By Francis Chachula, Niel Erasmus

"In 2002, Titanium Corporation Inc. focused its attention on the development of a process for the recovery of valuable heavy minerals, including zircon and titanium bearing minerals from oil sands froth treatment tailings. The development of a process flowsheet to recover the titanium and zircon minerals from the Athabasca oil sand tailings has evolved through many laboratory, bench and pilot scale iterations.Initial development work focused on so-called ‘beach sand’ material, which is tailings material deposited over a period of time close to the tailings pipe discharge. The derived process flowsheet primarily consisted of gravity concentration stages, which is typical of all mineral sands operations. Titanium Corporation’s business plan calls for the utilization of tailings directly from the tailings pipeline – the process developed around the beach material is therefore risky as it differs from the live tailings in a number of aspects, with the effects on the process unknown. Titanium Corporation successfully obtained an intermediate product through a ‘bulk sampling plant’ coupled directly to an Oil Sands operator’s froth treatment tailings pipeline, during the summer of 2005. Subsequent further processing of this material resulted in poor separation or concentration of the heavy minerals due to a number of reasons. A more effective process flowsheet had to be developed.The newly derived process is basically a marriage between mineral sands technology and oil sands technology that can economically concentrate and recover the heavy minerals and entails the incorporation of a flotation process that takes advantage of the entrained bitumen in the feed to separate the heavy mineral fraction from the quartz gangue This paper will describe the newly derived process as well as the successful results of a subsequent on-site concentrator pilot campaign completed during the 3rd Quarter of 2006."

Jan 1, 2008

The failure of a seal in an underground coal mine can have devastating consequences. Traditionally seals have been constructed as flat walls. Large hydraulic dams, on the other hand, are constructed as concave arches. This paper presents an initial analysis of the potential strengthening of seals by concave construction as well as the impact of additional cost in terms of labor and material. A simple finite element analysis of three-dimensional solids under uniform loading allows comparisons between the strength of a flat wall and a curved arch. The models are examined for reasonable behavior and displacements at 20, 50, 100, and 200 psi loads.

Jan 1, 2007

By Claude Bazin, François Robichaud

"The control of the operation of semi autogeneous grinding (SAG) mills is mainly achieved through adjustments of feed rate, water addition and speed of rotation. With the introduction of image based sensors it is possible to have on-line measurements of the size distribution of the ore feeding a SAG mill. This information could be used for the development of a control strategy to regulate a SAG mill feed size distribution to a set-point. The control of the size distribution is carried out by adjustment of the speed of the ore extractors underneath the ore stockpile. The potential and difficulty associated with the implementation of such control are assessed using a multivariable statistical analysis of data collected from the operation of the grinding circuit of the Agnico-Eagle Goldex mine in Val d’Or, QC. A control strategy for the ore feed size distribution is assessed using a simulator of the ore stockpile operation.INTRODUCTIONThe popularity of autogeneous (FAG) and semi autogeneous (SAG) grinding mills has been increasing over the last 20 years (Napier Munn et al., 1996). These mills use the ore or a combination of ore and steel balls as the grinding media. Since the ore is used as a media for the size reduction the mill ore feed size distribution impacts the performance of the mill. Several papers were recently published on this subject (Dance et al.,2011; Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011) showing the advantages of preparing the feed size distribution to maximize the circuit throughput or mill efficiency (Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011). With the availability of particle size sensor based on image analysis of the ore on a conveyor (Partanen et al., 2001; Bobo et al. 2004; Edwards et al., 2002; Wei & Craig, 2009) it is possible to control the size distribution of the ore feeding autogeneous or semi autogeneous mills (Partanen et al., 2001).This paper describes a method to control a size distribution index to a set point via the manipulation of the speed of the extractors that withdraw the ore from the stockpile feeding the grinding circuit. The paper consists of 4 sections. The first section describes the considered grinding circuit. A statistical analysis of the collected data on the grinding circuit is presented in the second section. This analysis demonstrates the necessity of using real-time adaptation mechanisms to develop a control strategy for the size distribution of a grinding circuit. The third section describes the method used to simulate the ore stockpile operation and the last section describes the preliminary development and assessment of a control strategy of size distribution of the ore feeding a SAG mill."

Jan 1, 2013

By Rebecca Hardy

The escalating demand for coal, in addition to the steady depletion of reserves, has produced increasingly difficult mining conditions for coal mining companies within the United States, resulting in mining above or below previously mined areas. The safe and productive extraction of coal in a multiple seam environment requires specific design technology for effective mine planning. Presently, one of the most useful design approaches for multiple-seam feasibility analysis is using numerical modeling techniques. Therefore, numerical modeling programs LaModel 2.1 and Lam2D were used to determine the feasibility of mining the Herrin (No. 6) coal seam over an existing room-and-pillar mine in the Springfield (No. 5) coal seam in the Illinois Basin.

Jan 1, 2006

By R. Bitmead, R Barker, A. Drake

The opencut coal mine is located about 50 km west of Brisbane, the capital city of Queensland in Australia and produces about 140 000 tonnes per annum of high grade steaming coal utilising a small team (ten) of personnel to mine remnants of the Aberdare and Bluff seams left by previous underground mining. This, in itself, is not remarkable except that one of the seams has commenced spontaneously cornbusting resulting in temperatures in excess of 55’C commonly being encountered with portions of the p’it being wreathed in steam. To extend mine life, the mine required a blasting system that would: . Safely blast in temperatures below 100°C . Significantly reduce overall unit costs . Not increase the current vibration impacts on nearby housing . Not increase flyrock.

Jan 1, 2002

Vale Inco?s Coleman Mine 153 Orebody is a narrow vein deposit with a dip that varies from 30 to 70 degrees. It has a strike length of 1200-feet and the current mining extends from the 4250 level to 5100 level. The primary mining method is mechanized cut-and-fill although drift-and-fill is also used in wider sections of the ore zone. The ore body is divided into six main mining horizons with an average of three mining blocks on each cut. There are some interesting rock mechanics challenges related to mining upward from the 4550 level and these challenges are increasing as the mining front progresses toward the 4400 level. The mining of the 4550 block is currently at cut 13 whereby the crown pillar thickness is approximately 60-feet. Significant factors including increased stress levels in the crown pillar, the change in the orebody dip at this level and a recently indentified geological structure in the footwall, have all combined to introduce seismic activity in the footwall of 4550 level. Closer studies revealed a consistent trend of seismicity that was further examined with core logging and Acoustic TeleViewer technology. The data from these assessment tools disclosed a geological structure that is influencing the mining environment. This paper reviews the seismic analysis for the 4550 level mining zone and the subsequent procedures used to locate the new geological structure. The paper also discusses the key mining practices used to successfully mine cut 13 of 4550 level and briefly reviews the studies that are underway to up-date the mine design for continued extraction of high-grade ore in a safe and productive manner.

May 1, 2009