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By Manchao He, Xiaoding Xu, Yubing Gao

Based on the geological conditions of the 15,600 working face in Pingba Mine, this paper establishes the mechanical model of the stress distribution in the pre-working face under the influence of mining pressure based on a hemispherical stress distribution theory. The analytical formula for the stress distribution is determined. A numerical model for the mining pressure under the influence of key layer is also constructed, and dynamic stress interruptions under various roof-cutting relief schemes in different horizontal positions are simulated. The stress distribution of the stope under different pressure relief schemes is analyzed and compared with the theoretical results. Results show that blasting roof cutting blocks the transmission of the abutment pressure to the working surface to the coal and rock mass in front of the working face, reducing the significant working face mining pressure influence range, which decreases as blasting pressure relief position moves toward the working face. Stress accumulation in the coal in front of the working face also increases as the blasting pressure relief position moves toward the working surface. Engineering practice results show that blasting cutting can effectively block the propagation of mining pressure, and the stress impact range is shortened by approximately 1/3.

Jul 31, 2021

By B. J. Vorster

The South African Iron and Steel Company Ltd operates a large Iron Ore Mine in the Cape Province of the Republic of South Africa. The Mine is situated on the farm, Sishen, approximately 600 km west of Johannesburg. The annual capacity of the Mine is 100 million ton of ore and waste and the present production level is 60 million ton. The Mine operates continuously for 6 days a week. The primary equipment used to maintain the above production are 55, 154 t trucks (Wabco and Unit Rig) and 14 P & H shovels. During 1979 an urgent feasibility study into the application of a trolley assist system was launched at the Mine. This was due to the unrealistic escalation in the cost of diesel fuel and the appeal from the South African Government to reduce the use of fossil fuels. Tests and economic evaluations proved the system technical feasible. The estimated discounted cash flow return exceeded 50% and the diesel fuel consumption would be reduced by 120 million litres in the next 10 years. Management subsequently approved the implementation of the system at Sishen. The construction of 7,7 km of line and the conversion of 66, 154 t trucks were completed in March 1982.

Jan 1, 1985

By Rolph Linder

It is generally known and accepted that the efficiency of the blast furnace process, expressed by the symbols specific production and coke consumption, is influenced by a large number of variables connected to the type of burden, the alloy contents of the produced iron and the conditions of production as composition and preheating of the blast, the furnace heat losses and the possible use of fuels added on the tuyere level. It is furthermore obvious, that the coke consumption of the process will be influenced by specific production which could be said to be the result of two counter-acting influences. With high production, realized through intensive blowing, the passage time of the ore through the furnace is shortened leading to a decrease of the indirect reduction taking place in the upper part of the shaft within the temperature range of 300 to 1000°C. Simultaneously the external heat losses (constant in time) will be smaller calculated on each ton of iron and the resulting change in coke consumption will be the effect of both those influences. It is certainly of interest to be able to predetermine the effect of those variables before a change in the production practice is carried out.

Jan 1, 1964

By Xin Jingmin

The present trend toward concentration of mining activities into fewer sections of a coal mine and application of automation and remote control technologies favor utilization of the longwall mining method. Longwall mining also offers the potential for improving safety and productivity, better ground control and higher levels of recovery. The production potential of a longwall operation is determined by the interaction of several factors including the rate of face advance, the thickness of extraction, the speed of cutting and hauling equipment, the physical dimensions of a longwall panel, and the reliability and availability of the equipment. Some of these factors are dictated by the environment in which the extraction process is to be undertaken. Some other factors, however, can be manipulated in order to optimize the production of the system. The objective of this paper is to develop a stochastic model to simulate the production operations in longwall mining. The activities and parameters that affect the production potential of the longwall operations and the relationships between the activities are identified. An attempt is made to verify these relationships through the use of existing data. The model is then used as a computer simulator to investigate the effects of various activities and parameters on the production capacity of a panel.

Jan 1, 1983

By Vinton A. Garbesi

INTRODUCTION The placing of the monolithic concrete lining of the Bi-County water tunnel illustrates the state of the art in the materials, equipment, methods and productivity available to the industry for this type of construction. This paper describes the management and engineering rationale used by Clevecon, Inc., a subsidiary of S & M Constructors, Inc., Solon, Ohio, in developing the concrete mix designs, the design of the specialized equipment, and the methods used in the manufacture, transport, and placing of the lining. Also discussed are the changes to the reinforcing steel design instigated to facilitate its handling and placing, as well as to enhance the quality of the lining. The intent is to bring to the attention of designers the effects certain design aspects have on the construction of reinforced concrete tunnel linings, as well as indicate the areas where research and development effort should be expended to further reduce costs for this type of construction. GENERAL DESCRIPTION The Bi-County Water Tunnel West is a water transmission main owned by the Washington Suburban Sanitary Commission of Hyattsville, Maryland. The facility was designed by A. A. Mathews Division (formerly STRAAM Engineers, Inc.) CRS Group Engineers Inc., Rockville, Maryland. It is to be utilized as the central feeder of potable water for the northern suburbs of Washington, DC, extending some 4,597 meters from a riser shaft at its western terminus to the eastern interface, with a similar tunnel which extends eastward about 5,180 meters; the majority of its length being under Interstate Highway 1-495. The excavated and finished diameters are 3.8 and 3.05 meters respectively. The reinforced concrete

Jan 1, 1981

By Frank McKinley

This paper discusses the basic requirements covering the area between the initial discovery of a deposit and a decision as to whether or not the deposit will support an economically sound mining venture. The procedure for gaining the necessary information and the sequence of steps is in .some respects parallel to gaining like information on deposits other than the sand type. For the sake of simplicity the stages and sequence are presented as follow: first, the initial reconnaissance drilling; second, the exploration drilling to permit initial ore body estimates; third, metallurgical testing; fourth, the estimate of plant and operating costs; and last, the economic balance. The objective of the reconnaissance drilling is to collect sufficient information to reach a decision as to the advisability of further exploration or the abandonment of the area. This initial phase will comprise, in addition to some drilling, a thorough reconnaissance of the area, a study of the best geological and topographic maps of the region, and a study of any literature available, all under the guidance of a geologist.

Jan 1, 1958

By L. Rintala, P. Latostenmaa, T. Kalliomäki, O. Forsén, I. Lehtiniemi, M. Lundström, J. Aromaa

Conductivity is a typical physico-chemical property that has a substantial effect on electrical energy consumption in the electrorefining process. The objective of this study was to examine the validity of regression models predicting the conductivity of industrial copper electrorefining electrolytes. The models were based on data measured with synthetic solutions on a laboratory scale. The variables included in the models were temperature and the concentrations of copper, nickel, arsenic and sulfuric acid. Both first-order and combined effects of the variables were investigated. The measured data were analyzed using MODDE 8.0 modeling and design software. The validity of the models was investigated by comparing the predicted values with tankhouse conductivity measurements taken at the Boliden Harjavalta Copper Refinery in Pori in Finland. During the tankhouse tests, the conductivity and temperature of the industrial electrolyte were measured, and the composition of the electrolyte was analyzed. The measured conductivity values were compared with the predicted conductivities calculated using the developed models. The conductivity model developed in this paper was shown to be more accurate than previous models and suitable for industrial use.

By R. H. T. Garnett

Recovery performance may be measured in absolute terms or in relative ones that relate the actual mining outcome to what was expected. Consideration is possible in terms of the control exercised by the operator. The eventual mineral output is the only figure known with complete certainty. The paradox is that, while performance is judged against the sampling, the reverse also applies. The R/Efactor and its four component factors are the best overall measures of performance. The choice of equipment for sampling and mining is very important. The sampling process may introduce errors and the sample density may lead to poor mining block selection. The resulting recovery performance is most visible on a regression curve. Changes in the cutoff grade must also be taken into account to avoid unrealistic production targets. Estimation methods incorporating different mathematical assumptions and techniques may cause over-or underestimation of the grade and volume. If the selective ability of the mining equipment does not match the size of the reserve blocks, the performance will be compromised. The treatment plant efficiency, the throughput rate and the degree of bedrock clean-up also affect performance. The maximization of throughput does not yield the optimum results. Marine conditions, which are outside the operator's control, affect not only operational availability but also performance. Examples are provided from actual operations for various commodities, showing the relative magnitude of the different effects upon performance.

Jan 1, 1999

By Anshumali Gangwar

This paper outlines the chance constrained binary integer program developed by the author in his doctoral dissertation at Columbia University, New York, 1973. The optimum planning of a mining program involves the determination of a production schedule which maximizes the net present value of the mining property. The actual mining process has an inherent stochastic and discrete nature. In particular, the demand for the mined product and the grade of the mineralized material in the ore are the two most important stochastic mining variables. The decision to mine or not to mine any block of a mining property is a binary integer variable. This paper outlines the formulation of pit production scheduling problem as a chance constrained binary integer program. The deterministic equivalents of these chance constraints are nonlinear in binary variables which have subsequently been linearized. The planned objective is to meet with a stipulated high degree of reliability, all the random demands as they materialize with a random supply of the commodity being marketed as determined by the eventuating random grade values of the exposed ore blocks and the mining system's production constraints such that the discounted total production costs over the entire planning horizon are minimized or the profits maximized. The emerging surface of the mine in each scheduling period is determined by the values taken by the binary decision variable associated with each block of material in each scheduling period.

Jan 1, 1982

By J. R. Aggson

In an effort to minimize cutter roof failures in longwall development entries of a West Virginia coal mine, a three-entry system was developed with the center entry arched into the strata above the coal seam. The purpose of the arched entry was to modify the horizontal stress envelope associated with the three-entry system and improve roof conditions in the adjacent outside entries. Contrary to the predictions of elastic analyses, the presence of the arched entry improved roof conditions by eliminating cutter roof failures in the outside entries. In-situ stress determinations and numerical modeling techniques were employed to quantify the impact of the arch and investigate the mechanism of stress relief associated with the arched entry. Stress determinations and underground observations indicated that the roofs of the outside entries were within the zone of stress relief generated by the arch, and that failure or movement along bedding plane discontinuities played a major role in the formation of the stress relieved zone. The numerical model incorporated horizontal joints to simulate bedding plane discontinuities. The properties of the joints were adjusted until roof stresses predicted by the model were in agreement with measured stress profiles. This process resulted in a calibrated analytical tool that could be used to evaluate alternate mining geometries.

Jan 1, 1988

By Al. Valero, A. Domínguez

This paper appraises the energy requirements for mineral production through an exergy approach taking into account the long-term decline in ore grades. In this context, the exergy required for restoring resources from the dispersed state of the so called Crepuscular Earth into the physical and chemical conditions in which it was delivered by the ecosystems in light of the available technology, is known as exergy cost. The exergy cost is the point of reference in order to evaluate in a single variable, characteristics such as composition, concentration (ore grade) and the state of technology. The ratio between the exergy cost and the minimum thermodynamic cost required to accomplish the same process is called unit exergy cost and provides a measure of the irreversibility of the analyzed processes. Otherwise, the ore grade is one of the main issues which affects directly the costs associated with mining the ore. Hence, the unit exergy costs to produce minerals from progressively lower ores grades by the current technology is analyzed. The unit exergy cost curves are presented for some important minerals such as aluminium, chromium, cobalt, copper, gold, lead, manganese, nickel and zinc, in order to know the trend when the ore grade decreases.

Jan 1, 2011

By Doug Harding

The Niagara Tunnel Project (NTP) is a 10.4 km (6.5 mi) long, 14.4 m (47.5 ft)bored tunnel that will run under the City of Niagara Falls from the Upper Niagara River to the Sir Adam Beck Power Station. The completed project will enhance the capacity of the Sir Adam Beck Power Station by adding 500 m3 (17,657 ft3) of water per second through the tunnel. The paper describes the project in which the world’s largest hard rock TBM at14.4 m (47.5 ft) in diameter is being used to excavate the tunnel. Features of the project include the use of a High Performance (HP) Main Beam TBM, state-of-the-art ground support and tunnel logistics systems. A continuous conveyor is used for muck transport from the TBM to the disposal area. The paper will also discuss the optimization of the TBM system as the tunnel advances through the varying geology. In June 2004 the Ontario Government announced that Ontario Power Generation(OPG) had been given approval to proceed with the third tunnel under the city of Niagara Falls—the first two tunnels were constructed in the 1950s using drill and blast techniques. The third tunnel, appropriately named The Niagara Tunnel Project (NTP), consists of a 10.4 km (6.5 mi) long TBM-bored, concrete-lined tunnel with a finished diameter of 12.5 m (41.1 ft). In August 2005, OPG announced the selection of Strabag AG of Austria, one of the largest construction companies in Europe, as the contractor to build the NTP Project at a price of CAN $600 million. In September 2005, Strabag A.G. selected The Robbins Company of Solon, Ohio to design, manufacture and deliver a new 14.4 m (47.5 ft) diameter High Performance state-of-the-art Main Beam Hard Rock TBM to be used in the excavation of the tunnel. In order to meet the aggressive construction schedule as requested by OPG/Strabag, the TBM needed to be supplied as “ready to bore” within a 12 month period after the contract was awarded to Robbins.

Jan 1, 2007

By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger's and Kick's laws (Bond, 1952). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed.

Jan 1, 1994

By X. P. Zheng

Barium fluorphlogopites were synthesized in the laboratory in order to investigate the effects of composition and crystal parameters on surface properties and floatability. The negative surface charge of mica is determined by the barium content, the crystal cell volume, the distance between layer structure and the Ba-O bond length. The higher barium content, the larger the cell volume and distance between the layer structure and the longer the Ba-O bond, the more negatively charged the surface is and hence the more easily adsorbed to cationic collectors. When an anionic collector (sodium oleate) is used, barium oleate forms on the mica surface and hence chemical adsorption occurs.

Jan 1, 1993

By William F. Riggs

The basic theme of this symposium and panel Is Rotation Pads: Are They Optimized? There Is a. reason for phrasing the title In the form of a question. There Is not only the technical competency which we must address; there Is the operating philosophy that must be evaluated on the part of both the customer and the supplier. Customers desire reagents which are trouble-free and capable of providing that extra amount of selectivity or recovery. When they receive ft, after the supplier has provided several years of Internal research, one of the first concerns/complaints Is the price of the product. This has a tendency to rapidly reduce a supplier's support level In the future. Suppliers are equally guilty from another perspective. When they approach a customer to Introduce a product, they often attempt to market by offering only a price Incentive. They then wonder why a customer doesn't respond Immediately to the incentive. They are often oblivious to the fact that the reagent cost is such a minor aspect of the operating budget, and the customer has many more pressing problems on a day-to-day basis In comparison to the reagent cost. We need to establish the understanding that reagent cost Is an Inconsequential cost of operation, and yet has such a disproportionately high Impact on the success of the entire operation. This understanding Is required by both the customer and the supplier. We say to each other,' why are we discussing this since this has been obvious for some time?' The reason is relatively simple in that we talk about it, acknowledge it, and yet we do not adhere to it. The supplier provides a product along with test data containing statistics, analysis, recovery, grade and cost calculations while most of the time ignoring the operating technique which must be applicable In the plant In order to optimize the product. He expects the reagent to be substituted In the plant for the existing reagent and ft works or does not work after trying several variables. The operating management Is equally guilty, In order to best explain this to both the customer and the supplier, ft becomes necessary to review the basic purpose of the major reagents utilized In flotation. A collector is basically to Impact selective, maximum water repellency on the surface of a particular mineral, The frother has the purpose of providing a chemically stabilized membrane on the surface of the bubble at the air-water interphase. This, then, provides a host environment for the attachment of the collector-coated mineral to a bubble. The depressant functions In the reverse of the collector and must demonstrate the same or greater degree of selectivity than expected of a collector. The key area which has been Ignored Is the rate by which these reactions occur and Interrelate. This has a very specific effect on the operating technique and the compatibility of the chemistry, equipment, and the operator himself. Researchers, suppliers, and customers provide reams of data to demonstrate how their products or design produce, for example, higher kinetics, more selectivity, or more recovery. The Information is often true. After all, we are all learned men and laboratory and actual plant data do not lie. However, we must remember the theme of this symposium and panel: Flotation Plants: Are They Optimized? and Optimizing of Flotation Reagents? The direct, honest comment to the two titles is very simple. OF COURSE THEY ARE NOT The plants, equipment, and reagents had better not be optimized or else we are in trouble. The Issue of this panel discussion is to approach this subject from a slightly different or perhaps mainly Ignored aspects of optimizing reagents in flotation. When we have reagents which provide higher kinetics, more selectivity, and better recovery, how do we use them? Since each reagent has a different physical characteristics of froth, rate of recovery, volume effect on the compatibility of equipment, and many more aspects too numerous to mention, the question which has been severely Ignored Is, 'What degree of study and cooperation by both the supplier and the operating management has been conducted In order to prepare the operator for maximizing the performance of a reagent In relation to the rest of the system?" Prior to testing a new reagent, how much time Is spent to bring the actual operator(s) Into the program to make them feel part of the program? How much time is spent explaining to the operator on the float floor how to possibly take advantage of a reagent with faster kinetics or one which Is Inherently more selective? What

Jan 1, 1993

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By James R. Aggson

This paper summarizes the research effort that was conducted as part of a cooperative agreement between the U.S. Bureau of Mines, Denver Mining Research Center and the Pittston Company. The objective of this investigation was to further the understanding of the floor heave ground control problems that have plagued underground mining of the Beckley coalbed in southern West Virginia. Floor heave is both economically and structurally undesirable. Floor heave is accompanied by a redistribution of stresses and loads that are associated with an underground opening. This redistribution of stresses may cause roof or pillar problems that otherwise would not have occurred. The floor heave that occurs in the Beckley coalbed is not cons1dered to be caused by "squeezing" or plastic flow of the materials involved. The "squeezing" type floor heave is normally associated with relatively weak floor members that contain significant amounts of clay mineralization. The floor members in the Beckley coalbed are competent, relatively strong materials which fail in a manner that can best be described by the term "buckling." This buckling-type failure is indirectly related to the time-dependent deformation properties of the materials involved, but is considered to be more of a slender column, or panel, type failure than a "squeeze" or "flow"-related failure. The Beckley coalbed has had a history of floor heave ground control problems. The Glen Rogers mine, which opened in the 1920's, experienced extensive floor heave. In a report dated April 1929, James P. Keatley of the West Virginia State Department of Mines described the floor heave in the Glen Rogers mine. Several observations listed in the Glen Rogers report that appear to be consistent with current observations follow: 1. Water plays no apparent part in the floor heave process. 2. No evidence of gas has been found in the floor. 3. The thickness of the overburden, which varies from 183 to 396 m (600 to 1300 ft), has no relationship to the floor heave. In the conclusions of the 1929 report on the Glen Rogers mine, Keatley states, "A statement in brief as to the cause (of floor heave) would lie that an undetermined natural condition doubtless augmented by former and present mining methods causes the bottom to heave." As this report will show, Mr. Keatley's insight into the problem was most accurate. The "undetermined natural condition" is the existence of a biaxial, horizontal, compressive stress field. This stress field is in excess of that which would be expected from gravity loading. This stress field and the material properties of the floor rock combine to cause the floor heave that has been experienced. The floor heave that occurs in the Beckley seam generally occurs as arching, with a tension fracture, near the center of the floor span or as a break near the rib followed by vertical deflection (of the floor) at the rib and sloping of the floor across the entry. Examples of the floor heave are shown in Figures 1 and 2. The two types of floor heave are merely different manifestations of the same basic failure mechanism. When the entries experiencing floor heave were identified on a map of the mine under investigation, it was apparent that there was a relationship between the floor heave and the direction of the entries. The main entries, which are used for ventilation and haulage, experienced the majority of the floor heave. This directional relationship suggested a directionally related loading mechanism, or more specifically, a biaxial, horizontal stress field that exceeds what would be expected by gravity loading. Since the main entries were oriented at a bearing of N25oW, it was hypothesized that the maximum compressive component of the horizontal stress field was at 90° to these entries or at a bearing of N65oE. An in Situ testing program was developed to test this hypothesis and investigate the floor heave problem.

Jan 1, 1978

By Terry M. Tilley, Gregory P. Sherry

Purpose This is a unique project because the mining and tunneling industries have become highly mechanized and cannot normally accommodate a project where surface disturbance must be minimized or eliminated when there is less than 10 feet of ground cover. More traditional less mechanized methods must be employed and the skills and techniques for accomplishing the work have become all but non-existent. Therefore, this article highlights those areas of specialized skills and techniques. Methods The article will describe the design approach, the method(s) of ground support selected, the method(s) of excavation selected, the planning, scheduling and materials selection process. In addition, the article will describe the actual production characteristics and the results of the project. This approach will concentrate on productivities, adaptations to actual conditions and the quality of the end product. The design approach basically was to insure that the temporary support methods were adequate for the excavation and shoring stage of the project. The final concrete liner was designed using conventional structural design methods with the goal of constructing a tunnel that will last 100 years. The methods of support were spiling and variations of spiling in order to provide advance temporary ground support. In this project spiling is steel channel cantilevered and driven into the ground ahead of the working tunnel face which supports the ground allowing the face muck to be removed with overhead protection’. Conclusion This will be a discussion of the applicability of the methods. The actual performance of the excavation and shoring methods will also be evaluated. The final concrete liner completion results, the actual production rates and the conformance with budget and schedule will be addressed.

Jan 1, 2005

By E. Piesch, M. Urban, G. M. Hassib

INTRODUCTION In radon dosimetry diffusion chambers with passive [a]-detectors are now in current use to investigate the long-term accumulated indoor exposure of radon and daughters. For the diffusion chamber developed at Karlsruhe the electrochemical etching technique (ECE) was applied to reveal [a]-particles in polycarbonate plastic detectors. Compared to conventional etching, the main advantages of the ECE are the large track diameter, simple track counting in a single field of about 1 cm2, stability of the latent tracks to humidity and temperature and ability for a long-term accumulation up to 1 year. The electrochemical etching system was designed to etch in one run about 101 foils by using a 200 W high voltage generator. The commercial available radon dosimeter system is applied for a country wide program of measuring radon in dwellings. Because of an increasing interest of radon monitoring in mines, environmental monitoring and personnel dosimetry the properties of several track etch detectors have been studied. In the comparative study, Makrofol E, LR 115 and CR 39 were used in the diffusion chamber to investigate the etching techniques, the relative efficiency, the measurable [a]-energy range, and the counting statistic for the measurement of radon concentration. THE KARLSRUHE DIFFUSION CHAMBER The design of the Karlsruhe diffusion chamber is presented in Fig. 1. The upper part of the chamber is closed by a fibreglass filter to separate radon from aerosols, heavy dust particles and radon decay products. Inside the diffusion chamber [a]-particles from radon and daughters are detected by two passive radon detectors, one on the bottom of the chamber and the other in front of the filter. With the fibreglass filter applied the diffusion time is about 1 minute for radon. Therefore, also short-term variations in the radon concentration are recorded by the track detector. Mainly because of the short halflife of thoron (56 s) only about 7 % of the thoron concentration is expected in the diffusion chamber. Radon monitoring provides the estimation of the long-term accumulated mean value of the radon concentration with a lowest detectable concentration of 0.1 pCi/1 after 3 months accumulation by using polycarbonate track etch detectors, an electrochemical etching technique and track counting in a single field of 1 cm2. The diffusion chamber was optimized for the polycarbonate track detector with respect to chamber geometry and detector position [1]. Table 1 presents data for 3 different diffusion chamber designs. The radon sensitivity of the diffusion chamber depends on the [a]-energy discrimination and the angular response of the [a]-track detector. The difference between foil No 2 and foil No 1 can be explained by an additional plate-out effect of radon daughters on foil No 3. The high sensitivity of foil No 3 compared to that of foil No 2 is caused by an increase of the effective volume and the effective area for [a]-particle from radon and from plated-out decay products, respectively. Additional experiments have been performed in order to investigate the electrostatic properties of the plastic material and the properties of the diffusion filter [1]. The properties of bare a-detectors were compared with those inside the diffusion chamber. In contrast to laboratory results, dust particles deposited on the detector surface or in the environment of the detector obviously affect the reproducibility of the bare detector significantly. After 4 months exposure in dwellings the scatter of 10 dosimeters exposed in the same room exceeds a factor of 3 compared to 15 % for the diffusion chamber. COMPARATIVE STUDY OF TRACK ETCH DETECTORS Detector types and evaluation technique In the comparative study the properties of the following plastic track etch detectors have been applied: - LR 115 from Kodak Pathé, France, is a red dyed cellulose nitrate layer of 13 µm thickness on a 100 µm thick polyester support. The holes appears as bright spots in a dark red background. In an optical microscope counting of through etched [a]-tracks which must penetrate the detector layer - CR 39 from Pershore, England, is an ally1 diglycol carbonate plastic of about 300 m thickness. The normal etched [a]-tracks are counted by an optical microscope and the ECE [a]-tracks by a microfiche reader.

Jan 1, 1981

By C. Cassios

The paper covers the principles of Silver & Baryte Ores Mining Company (S&B) integrated Environmental Impact Assessment study for the whole of the extractive activity at the Fokida prefecture, over a time horizon of a quarter of a century. More specifically the paper includes: - The investigation and the evaluation of the environmental and socioeconomic variables of the specific area of concern. - The identification of the present and future impacts of the extractive activity. - An evaluation of the results of related reclamation activities. - Activity planning following a holistic approach, based on the principles of Sustainable Development, including environmental, social and economic aspects. - The monitoring of the procedure?s implementation.

Jan 1, 2003