Search Documents

By Brain W. Lawrence

Sublevel stoping generally is a large-scale open stoping method. It sometimes is referred to as long-hole or blasthole stoping. This method usually is applied to strong ore bodies that require minimal support and are surrounded by strong country rock. The ore body should be fairly regular in shape and well-defined. The dip of the footwall normally should be sufficient to allow broken ore to gravitate freely, although the method has been adapted successfully to mine some of the flatter ore bodies. Sublevel stoping is not dependent on the width of the ore body. Widths of less than 6 m (20 ft), however, make the utilization of long-hole drilling techniques more difficult. The main criteria for sublevel open stoping are competent ore and stable host rock, regular ore boundaries, and a footwall dip that exceeds the angle of repose of broken ore. Basically, the method entails providing access to the ore body at various subintervals between the main haul- age levels in order to drill and blast the intervening ore. Stope drilling is carried out from drilling drifts on the sublevels, and the ore is blasted in slices towards an open face, which generally is vertical on the downholes and may be inclined towards the open face for the up holes. The blasted ore gravitates to the bottom of the stope and is collqcted through drawpoints. Fig. 1 illustrates a typical sublevel stoping mine. The method requires extensive ore body development with relatively high capital expenditures. How- ever, much of the development is in ore, and production costs are comparatively low. Productivity rates are in the 13.6 to 27.2 t (15 to 30 st)/manshift range. The drilling, blasting, and loading operations are performed independently, and equipment utilization is high. Large outputs can be obtained with few units and limited personnel. Dilution with waste rock may occur if ore boundaries are irregular or if caving occurs, but 100% of the ore within the stope usually is recovered. Pillar recovery sometimes is a problem. It is at this point that most of the waste dilution occurs. At the time data was gathered for this report (1973), the sublevel open stoping method was used by more than 21% of the metallic ore mines in the US (9 out of 42 mines producing 1088 + t/day (1200 + stpd) and provided more than 12% of the total metallic ore production from these mines. DEVELOPMENT Access to a sublevel open stope mine is gained by a slope or shaft that is normally sunk in the footwall of the ore zone, away from any possible effects from blasting or other production operations in the stoping process. Main level interval is selected and usually varies from 45.7 to 121.9 m (150 to 400 ft), depending on the vertical extent of the ore body, the number of drawpoints required to maintain output, and the eventual stope height. The main haulage drifts are located at the bottom of the stope horizon, either directly in ore or in the footwall, with access crosscuts at intervals. Raises are driven up through the ore to connect to the level above and provide access and ventilation to the sublevel drilling drifts, which are driven horizontally in the ore for the length of the stope. Depending on the drawpoint system to be used, the stope is either undercut by a raise-and-cone system, from a scram or slusher drift, or simply provided with a drilling drift from which access is made at intervals to the main haulage level. A raise driven either at the end or in the middle of the stope (if stope access is from both ends) provides a start for opening up a vertical slot across the stope to establish a free or open face for long-hole drilling and blasting. Sublevel stoping has proved to be a relatively safe and economical system.

Jan 1, 1982

By Lei Fu

In the wake of the collapse of the ceiling panels in the I-90 section of the Central Artery/Tunnel (CA/T) Project in 2006, the long term performance of tunnels became the focus of the tunneling industry. At the time of the accident, the tunnel had been open to traffic for about 10 years. This paper will discuss design methods and considerations to improve the long term performance and durability of a tunnel. The major problems found during the inspection of several tunnels will also be presented and the causes for these problems will be discussed.

Jan 1, 2008

By Fang-tu Wei

The purpose of developing this system is to monitor the operating status of the fully mechanized mining system in order to raise it's reliability. The completely mechanized coal mining system is a complex dynamic system, involving many sub-systems. It is the key of the mining process as well as a connected system composed of each working procedure. Any hitch in breaking, loading or transporting will interrupt the mining process. The analyses of the operating data has proved that the reasons causing the interruption of the extracting processes mainly come from the inner of coal face. Hence, after the installation of the MINOS system in England, the delays of belt conveyor sub-system were reduced. But the production increment was not as much as expected.

Jan 1, 1986

By Armando Z. Remacre

The purpose of this paper is to attest the effectiveness of conditional simulation technique's for determining the principal pre-mining control parameters during mine planning. These parameters include: determining the selective mining unit size, the optimum pre-mining grid spacing and the number of mining faces to be used. The research work was developed in an iron ore deposit located in the Minas Gerais State, Brazil. Phosphorus was considered the master control variable for the research work. This paper provides the structural analysis (variography) , a description of the steps the were taken for setting up the numerical model, the kriging studies which were applied to the numerical model and the final optimum grid spacing, block size and number of mining faces.

Jan 1, 1992

By Ugo Preti, Gianfranco Ferrara, Mark C. Williams, Thomas P. Meloy

Theoretical and experimental lockedness profile curves are the same when the model assumptions are met. They appear different because of the way the data is presented. Continuous theoretical lockedness profile curves present all the data, including the wings of the curves giving them the deep U-shaped configuration. Experimental lockedness profile curves, because of the measurement methods, are discretized, obscuring the wings. If the continuous theoretical curves are discretized, they are similar to the experimental curves. When a float sink test is used to measure an interval of locked particles, the integrated averaging of the separation process obscures the wings of the continuous theoretical curve. To disclose the wings with experimental data, narrow density or measurement intervals must be used.

Jan 1, 1992

At periodic intervals-typically at the end of each month, quarter, and year- the accounting department in each division of a company distributes a set of reports to the division's managers. These reports describe the financial performance of the division and its major departments over the time period. They identify the revenues generated from the division's products and the costs incurred in their production. The results from individual departments are combined to determine divisional results, and the results from individual divisions are in turn consolidated to determine the results for the overall corporation. Cost reports for each of these activities show actual amounts, budgeted amounts, and the differences between the two. This reporting is carried out for the individual period and for the year to date.

Jan 1, 1999

By Wang Xuewen, Bao Qingbao, Xie Jiacheng, Yan Zewen

This study puts forward a virtual memory cutting method based on the prediction of AFC shape. The integrated key technologies include the generation of virtual teaching cutting curve, the height adjustment strategy of front and rear drum, cutting height adjustment strategy, the reciprocating cycle of cutting technology and so on. Through the two algorithms, SVM and ELM, the form of the next working cycle can be obtained by rolling prediction using the form information of AFC in the first several circulations. Based on the obtained dynamic fluctuation of the form in the vertical and horizontal working faces, the position and cutting information of the shearer during its operation can be controlled and predicted in advance, the memory cutting template can be adjusted and modified in time. Employing this method, the more stable, accurate and timely memory cutting of the shearer under the uneven coal seam could be realized.

Aug 16, 2021

By Richard Dunmore

Most coal seams in the United Kingdom contain firedamp, some of which is released into the ventilating airways as the strata around a working relax. The close sequence of seams in carboniferous strata is usually such that seams in adjacent roof and floor strata are also disturbed and potentially large quantities of firedamp can enter the workings from these sources. Since emission depends on the disturbance rate of the seams, higher coal production rates associated with increased mechanisation in the colliery accentuate the problem. Firedamp concentrations in UK underground workings are subject to statutory regulations1. To enable production to continue uninterrupted and in safety the problem is tackled by ventilation engineers by the provision of adequate ventilation in the airway and by the technique of firedamp drainage. In the UK the latter usually takes the form of cross-measure boreholes in roof strata, and increasingly also in floor strata, drilled from the gateroads, especially from return airways. In this way firedamp emitted from disturbed seams in adjacent strata can be extracted directly to the surface or to some safe position outbye, and the airways of the working district are not contaminated. Despite the use of firedamp drainage, considerable quantities of firedamp enter the ventilation air- stream and the concentration levels may be seen to be related to the coal production activites in the district. Figure 1 shows a typical recording of methane concentration obtained by a BM1 methanometer at a position outbye in a return gate where firedamp drainage was operating. The general rise in level throughout the working week may be seen, together with peaks occurring during production shifts and the gradual decay of emission over the weekend. In view of the fact that ventilation and firedamp drainage contribute a considerable proportion to the operating costs of the colliery, it is essential that the ventilation and firedamp drainage requirements be calculated as accurately as possible at the planning stage of the working. In the past the ventilation engineer has had to rely heavily on experience in neighbouring workings but with the development of mines in previously unworked areas there is clearly a need for a methodical approach to meet the requirements of planning. The work described in this paper represents such an approach and it is hoped that successful development and application of the method of prediction will lead to the widespread adoption of a routine procedure. The current development of a method of prediction of firedamp emission into longwall workings is based on the theoretical treatment of gas emission from coal seams proposed by Airey2 in 1971, which was based on his study of gas emission from broken coal 3 which was published in 1968. In common with other experimental methods of prediction, the method enables a calculation to be made of the total make of gas emitted from all sources of gas in the strata per tonne of mined coal.

Jan 1, 1980

By M. Rossi

Geologic ore controls are usually interpreted and modeled from cross sections and plans, as they are critical to construct meaningful and accurate ore reserves models. The geologic boundaries are often modeled as two-dimensional surfaces or three-dimensional solids and used as hard boundaries to constrain the grade estimates. In many cases, the geologic model is the most important factor in estimating the mineralized tonnage predicted by the reserves model, since it is generally drawn to separate the volumes of mineralized and waste material within the deposit. Because geologic models are interpreted from widely spaced drill hole data, the accuracy of the contacts of the different zones modeled could be poor. The lack of accuracy in the definition of geo-logic zones can have a significant impact on a resource model. In this paper it is shown that the geologic interpretation can and should be checked for potential biases. In particular, an unbiased overall volume of the original drill hole information with respect to the resulting solids and codes assigned to the block model is sought. In addition, geostatistical conditional simulations can be used to simulate the geologic attributes modeled, thus providing an alternative view of the geology, and a measure of the uncertainty of the geologic boundaries. The simulated geologic model provides an assessment of the impact of alternative geologic interpretations on the over-all uncertainty of the reserves model. In the case study presented here, the uncertainty of the geologic model is quantified for a difficult-to-model, high-grade copper deposit in northern Chile.

Jan 1, 2004

By J. P. A. Hortley

In conventional electrolytic copper production, looped copper starter sheets, on which the copper is deposited, are supported by hard-drawn copper bars. These bars have cross-sectional areas of between 1 and 1.4 in2. The suspension bars serve two purposes: firstly to support the cathode, which can weigh up to 150 kg, and secondly to introduce the electric current. IYI examined these two requirements in detail to see whether a better design could be achieved. Considering firstly the current carrying requirement, it wee noted that under normal refinery conditions, each cathode carries between 400 and soak. If one applies the generally accepted rule that 1 in of copper can carry 1000A4 a suspension bar need theoretically contain only 1/2 in2 of copper. This contention is supported by the fact that the 500A current passed through the cathode also appears to pass quite satisfactorily through the two support loops, which have a cross-sectional area of approximately 0.35 in 2. Turning now to the strength requirements, the cross-section of hard-drawn copper bar required to support a 150 kg cathode is relatively small, and easily met by the present designs. It is not, however, the static load applied by the 150 kg cathode immersed in the electrolyte which causes the damage to the hard-drawn copper, or even the increased static load of a 100 kg man standing on one foot in the centre of the bar; the damage is actually caused by the dynamic loads applied to the bare when they are dropped, thrown into containers, lowered too rapidly into the cell, or otherwise mistreated. The important requirement is therefore that the bar does not permanently bend. Its proof stress of limit of proportionality, rather than the amount that it deflects under static loading, should be the strength criterion. With the hard-drawn copper normally used for the manufacture of suspension bars, it has been observed that the copper becomes annealed during use, particularly at the end of the bar where the current is introduced. Thus, after a relatively short time, the cathodes are in effect being supported by annealed copper bars.

Jan 1, 1978

By C. E. Schmidt

Homestake crushing and grinding, practice-was completely revised during the period of 1951-1953. During the conversion, both crushing plants at the head frames the Ross and the Yates, were changes at each plant involved replacing the two 8K gyratories with a rebuilt, The flow sheet single thirty inch gyratory for primary crushing to four inch. The gyratory discharges to a seven foot Symons standard cone crusher set at 1-1/8 inch opening. The cone discharge goes to Tyrock screens fed by vibrating feeders. Screen openings, are l/2 inch by 5 inches: at the Ross. Plant, and 3/8 inch by 5; inches at the Yates. Screen oversize goes to a seven foot Symons cone brusher converted to the short head type; the crusher setting is 5/16 inch. Crusher, discharge; is closed circuited with the Tyrock screens, The above crushing installation delivers a minus 5/8 inch rod mill feed to the mill bins. The crushing revision eliminated the-180 heavy South African type stamps, which had previously reduced the 2 inch crushing plant product to a minus 5/8 inch feed for the 10 rod mills, and 8 ball and pebble.; mills, other revisions in the mill flow sheet, resulted in 3 identical units of about 1350 tons per day, capacity per unit, Each unit consists of an open, circuit 6 foot by 12 foot rod mill, fed from two 24 in by 48 inch vibrating feeders of the four feeders available per unit. A gathering belt delivers the ore discharge of the feeders to the Transportoweigher, which in turn discharge s, into the rod; mill feed spout. Rod Mill discharge is the feed for the 9 foot by 11 foot ball mill, The ball mill is closed circuited with a 16 foot by 30 foot Dorr Classifier. The Homestake make amalgamator, for coarse gold recovery in the ball mill circuit, made' necessary a 36 inch spiral conveyor between the classifier rake discharge and, the combination drum and double scoop feeder on the ball mill See Figure l - the flow sheet of the two stage grinding units.

Jan 1, 1958

By Luzheng Chen

The study aimed to evaluate the separation performance of a full-scale continuous centrifugal concentrator in the reconcentration of fine hematite from high-gradient magnetic separation (HGMS) concentrate in a tailings recovery plant, and to investigate the usefulness of the technology in industrial applications. The effect of changing the operating variables on concentrate grade and recovery was studied. The results indicate that among the variables considered in the study (water spray pressure, centrifugal acceleration, feed volume flow rate, feed percent solids and the reciprocation and spacing of water sprays), the change in centrifugal acceleration affects the recovery the most and continuous concentration can be achieved only when the water spray has a sufficient pressure. In addition, the reciprocation and spacing of the water sprays have a significant influence on the concentrate grade. Although performance of the full-scale concentrator is slightly inferior to pilot-scale trials, due to limitations in the scale-up design and the fluctuations of feed material characteristics, this concentrator provides a potential way to achieve continuous concentration and obtain a high-grade concentrate from tailings.

Jan 1, 2010

By Artur Skoczylas, Paweł Stefaniak, Sergii Anufriiev, Wioletta Koperska, Maria Stachowiak

In many underground mines the excavated ore is produced during the blasting process. Drilling machines that are used to drill the blastholes, where later the explosives will be mounted, are crucial for this task. In order to optimize the currently used techniques, great emphasis is placed on increasing the efficiency and safety of blasthole drilling operations. The natural consequence is the development of monitoring systems and algorithms for processing operational and historical data. In this article, we present the result of the analysis of drilling data carried out as part of the IlluMINEation project. The paper proposes an algorithm for work cycle detection for drilling machines, together with the estimation of general operating parameters and diagnosing the drilling process on the basis of long-term data. Based on the available data, selected components of the drilling cycles have been indicated. Several approaches based on the parameters of the defined cycles are proposed. These approaches can be used in the future to create a full-fledged system for examining the effectiveness and safety of mining drilling rigs.

Jun 25, 2023

By J. Kemeny, K. M. Kim

A site-specific blasting model has been developed for the Asarco Mission mine. The main inputs to the model are the in situ block size (F80), the post-blast fragmentation (P80) and the intact tensile strength (To). The output from the model is the specific blast energy (ESE). Modern techniques are being used to obtain the input parameters for the site-specific blast fragmentation model. In particular, image processing software is used to obtain both the in situ block size and the post-blast fragmentation, and the Schmidt hammer is used to obtain the tensile strength at numerous locations throughout the rock mass. An initial model was developed using five test shots in an area of the mine that contains argillite and is relatively homogeneous. This process also included calibrating Schmidt hammer readings against actual Brazilian tensile measurements. The model for argillite will be updated as more tests are conducted, and the model will be extended to the four other primary rock types at the Asarco Mission mine.

Jan 1, 2011

By David A. Summers

Historically, the presence of methane has been a problem, mainly in and around the working areas of active coal mines, and only in these area has drainage been considered. Drainage, where practical, has been achieved through the drilling of holes forward into the coal and the surrounding strata from the working area. These holes generally have been short in length, although where methane drain- age operations around a longwall face have been undertaken, the holes have had to be longer in order to adequately drain from the center of the face into the access gate roads. In recent years, attempts have been made to degasify the coal seams in advance of mining, without disruption of the mining cycle. This is done by drilling much longer horizontal holes through the coal in advance of the working area. Under the aegis of the Federal Government, methods have also been developed for draining seams of coal of their methane content in advance of mining, but from shafts sunk from the surface, without using the active area of the mine as the location for the drill holes.

Jan 1, 1979

By Kendrick C. Hardcastle

The practice of transporting an ore or matrix from a pit area to the plant process area by means of a hydraulic pipeline has long been used by the mining industry. The industries which have classically used hydraulic transportation of ore to the plant include the phosphate mining industry, the heavy minerals mining industry, the clay mining industry, and the sand mining industry. As the advantages of this method become better known throughout the mining industry as a whole, and as techniques and expertise improve, its use will undoubtedly spread. Its use is presently being contemplated in iron mining diatomaceous earth mining, and several industries where massive removal of overburden is a predicate to economical mining. The widely recognized advantages of hydraulic transport at ion of ore from the pit area to the mines are: 1) Its relatively low initial capital costs. 2) Its speed of relocation at the mine site. 3) The washing and scrubbing action given to the ore by its hydraulic passage into the pit pump and the subsequent trip through booster pumps down the pipeline to the plant.

Jan 1, 1969

By Gary Repetto

Hiring trends in the construction materials industry have remained steady through the years, with some seasonal and geographical variances. However, after several years of a depressed market, the base metals industry has begun to actively hire only over the past two years, due to the continued strength in the prices of copper, lead, zinc, molybenum, etc. The expanding precious metals industry, which for the most part escaped the depressed markets of the mid-1980s, actually preceded the base metal companies in hiring of engineers and operations personnel. The coal industry has not seemed to rebound much from the cutbacks of the early 1980s. And the uranium industry, which began its decline a year or two earlier than the other mineral industries, has remained nearly nonexistent in the United States. In nearly all of the mineral industries, new hires have been mostly in the engineering and operations areas, not exploration.

Jan 1, 1990

By Graeme J. Jameson

A form of column flotation cell is described in which the contact between the feed and the air stream is made in a mixing device at the top of a vertical downcomer. The air-liquid mixture flows downward to discharge into a shallow pool of pulp in the bottom of a short cylindrical column. The bubbles disengage and rise to the top of the column to overflow into a concentrate launder, while the tails are discharged from the bottom of the vessel. The main advantages of the device described are that the overall height of the column is reduced to about one metre (3 ft), and the column can be self-inducing with respect to the air supply. Experimental results are given from tests on a feed stream to a conventional zinc cleaner circuit. Using wash-water to reduce gangue entrainment, the column gave high concentrate grades and high recoveries.

Jan 1, 1988

A form of column flotation cell is described in which the contact between the feed and the air stream is made in a mixing device at the top of a vertical downcomer. The air-liquid mixture flows downward to discharge into a shallow pool of pulp in the bottom of a short cylindrical column. The bubbles disengage and rise to the top of the column to overflow into a concentrate launder, while the tails are discharged from the bottom of the vessel. The main advantages of the device described are that the overall height of the column is reduced to about 1 m (3 ft), and the column can be self-inducing with respect to the air supply. Experimental results are given from tests on a feed stream to a conventional zinc cleaner circuit. Using wash water to reduce gangue entrainment, the column gave high concentrate grades and high recoveries.

Jan 1, 1987

By Michael J. Martinson

A powerful methane explosion in the mid-shaft pump station of a main service and ventilation shaft killed several persons on the station; more workers were killed when a mancage travelling down the shaft fell to the shaft bottom under a pile of broken concrete slabs from a brattice wall destroyed by the explosion venting into the shaft and a tangled mass of twisted steelwork from the equipped compartment. The newly-commissioned shaft serves a recent extension of an established mine on the Orange Free State (OFS) goldfield in South Africa. After outlining the circumstances in which the paper was written, and reviewing the nature and extent of the methane hazard in OFS mines, the paper describes the physical layout of the shaft and summarises evidence given by witnesses at the official inquiry concerning events prior to the explosion. No attempt is made to apportion blame for the incident.

Jan 1, 1989