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By Heinrich H. Becker, Vladimir L. Konyukh, Sergey P. Koltsov, Valery A. Davidenkol

Ramifide conveyor transport of the biggest russian coal mine "Raspadskaya" is simulated by Petri nets. There are 6 longwalls with accidental output. 12 belts, and 5 transfer bunkers at this mine. It is necessary to coordinate transport links for recon- structed conveyor system. Many tasks were solved by dynamic simulation, such as estimation of transport lateness, influence of transfer bunkers on an unevenness of total coal flow, maximum filling of each bunker, maximum time of main conveyor refusal, co-ordination of unloading for dependent bunkers, choice of main conveyor. The problem of mine-wide conveyor transport animation for dispatcher is formulated in reference to mine "Komsomolets". OBJECT OF SIMULATION The biggest russian coal mine "Raspadskaya" is situated on a south part of Kusnetski coal basin (West Siberia). One mines about 20.000 tons of coal out of flat seams by 2.2-4.5 m thickness in a day. There are 6 fully- mechanized longwalls by output 720- 3600 tons in a shift in mine. Coal haulage to surface is carried out by 12 belts having 1.6 - 4.3 km length. For a smoothing over un- evenness of coal flows 5 transfer bunkers to 300 tons were intro- duced. The designed conveyor system is divided into two parts (Figure 1). In the upper part some coal from Face 1 comes through belt road C, into transfer bunker 8,. From the opposite side that bun-

Jan 1, 1996

By David Caiden, Geoffrey Fosbrook, Jaidev Sankar, Donald Phillips, Jon Hurt, Anil Parikh

This paper will cover the progress being made on Phase 1 of New York’s Second Avenue Subway line. It will include a description of construction progress on the first contract to be let, which consists of a 248 m (815 ft) long launch box, 3.9 km(2.4 miles) of TBM tunnels and a tunnel sump; the packaging strategy for the remaining work; and the progress through Final Design of the remaining packages, including the two large span rock cavern stations.

Jan 1, 2007

By J. Sanko

When faced with the task of designing a process control system the control engineer must consider the following: 1. The type of process to be controlled and the manner in which the process must be controlled (the algorithm). 2. The type of equipment to be used to implement the control system. 3. A means of describing the control system for use in the following areas: a. Project engineering. b. Procurement of the control system equipment. c. Plant startup. d. Operator training and reference.

Jan 1, 1981

By Monserrat Rebolledo, John L. Uhrie, Candice Simms

Jetti Resources has developed a catalyst which overcomes the passivation of primary sulfides, enhancing the heap leach recovery of copper from chalcopyrite. It is important to emphasize that heap leaching of copper sulfides, whether with or without the Jetti catalyst, is a biooxidative process. We have demonstrated the efficacy of Jetti’s technology in bioreactors, with a chalcopyrite-dominant porphyry ore. In the presence of Jetti’s technology, optimal bioleaching conditions significantly increase copper extraction by 2.6 times that of the baseline.

Feb 1, 2025

By A. M. Marabini, G. Belardi, V. Alesse, E. Spaziani

The objective of this investigation was to verify the effectiveness of commercial and synthetic depressing agents in a complex oxidized zinc mineral flotation flowsheet using sodium sulfides and amine. The flotation tests were run on a finely-disseminated calamine ore grading 16.96% Zn and containing smithsonite, hemimorphite, cerussite, calcite, quartz and iron oxide, present essentially in colloidal form. The reagents used in the testing can be grouped into four classes: guar, starch, celluloses and acrylic polymers having various degrees of substitution and molecular weights. The depressants exhibiting the greatest selectivity towards the gangue (considered as a whole) are tripolyphosphate, polycarboxylic acids, methylcarboxylate guars and ethyl hydroxylic starch. In the rougher stage, the Zn recovery ranged from 70% to 80%, and the Zn grade ranged from 27% to 29%. In the Zn-Fe2O3 separation, the compounds ensuring the most efficient separation are the tripolyphosphate and the polycarboxylic acids. In the Zn-SiO2 separation, the compounds that exhibit the best separation efficiency are tripolyphosphate and guar of methylcarboxylic sodium salt. In the case of the Zn-Ca separation, methyl-carboxylic guar exerts the best depressing effect and does not adversely affect the Zn recovery. The separation provided by ethyl hydroxylic guar is also good, although not as good as the preceding reagent. Regarding the Zn-Al separation, the tripolyphosphate, the celluloses and the carboxylate guars are the best depressants. A hypothesis is advanced to interpret the depressing action of the most effective reagents.

Jan 1, 1995

By Richard M. Levine

Now that the former Soviet Union has divided into 15 countries, it is necessary to analyze separately the industrial mineral position of each of these countries. Industrial mineral data The former USSR was one of the world's largest producers of industrial minerals and had some of the world's largest reserves. Production and reserve figures, however, must be qualified. Even when these figures were reported by the Soviets, it was not clear that the production figures were for products that would be equally marketable in a competitive market economy. The reserve figures are based on the Soviet system of reserve classification, which is not directly equatable with the US Geological Survey and US Bureau of Mines classifications of economic reserves in market economy countries. The Soviets reserve classification system, in terms of which reserve figures will be given, is based on the reliability of the quantity of the material in place. The Soviet system assigns each occurrence to one of seven categories A, B, C1 and C2 are considered reserves; P1, P2, and P3 are considered resources. Mining in the former Soviet Union was authorized on the basis of the reserves in place in categories A, B and C, which had to be in prescribed ratios. However, the categories A, B and C do not necessarily correspond to Western concepts of material economically exploitable under present market prices with existing technologies.

Jan 1, 1992

By A. Halim, J. Gustafsson, A. van Wageningen, J. Johansson, · J. Lööw, K. Kocsis

A major part of the European Union’s (EU) project Sustainable Intelligent Mining System (SIMS) is investigating the development of diesel-free/carbon–neutral underground mines in order to ensure sustainable underground mining in the future. Replacing diesel machines with electric vehicles in underground hard rock mines has been widely acknowledged by the mining industry worldwide as a critical step to improve working conditions by reducing diesel exhaust–related contaminants, to reduce mine ventilation electrical power cost by reducing mine airflow quantity, and to reduce mine greenhouse gas emissions. All of these are major requirements to achieve sustainable future underground mining practices. A field trial of Epiroc’s 2nd generation of Battery Electric Vehicles (BEVs) at Agnico Eagle Finland’s Kittilä mine was conducted during 2019–2020. Vehicles tested were MT42 mine truck, ST14 Load-Haul-Dump (LHD), and Boomer E2 jumbo drill rig. This paper outlines the improvement of the working conditions observed in the field trial, and the opinions of the mine personnel at Kittilä mine on using BEVs instead of diesel machines. Measurements of atmospheric contaminants and air temperatures taken during the field trial clearly demonstrated a significant improvement of working conditions when BEVs were operating as opposed to diesel machines. This field observation was supported by the opinion of the majority of the Kittilä mine workers. However, some remaining concerns must be addressed before BEVs can replace diesel machines.

Oct 25, 2021

By C. D. Breeds, F. E. Valencia, J. H. Pye

The "FIRST AUTOMATIC SHOTCRETE TECHNIQUE" (F.A.S.T.) system, provides safe, efficient methods of remotely lining, shafts/raises. Two systems are available; one system will be used to shotcrete a 3.04m (10-ft.) diameter ventilation raise at a mine in Mexico; the other system incorporating remote control and sensing devices will to used to shotcrete ventilation raises from 2.75-4.9m (9-16 ft.) in diameter. This paper describes development of the system, from idea to field use.

Jan 1, 1981

By W. J. Wuest, M. J. DeMarco

The US Bureau of Mines (USBM) recently developed a rock-mass classification system for evaluating the structural competency of bedded roof strata. The system is called the coal-mine roof rating (CMRR), When compiled with the companion Lotus 1-2-3 spreadsheet, CMRR data are a useful tool with applications to longwall gate-road pillar design, safety-hazard mapping and artificial roof-support analysis. This paper gives a review of some applications under actual conditions found in current US mining operations.

Jan 1, 1997

By William H. Meadows

This information describes the ventilating and methane gas testing procedures utilized during the raising of a mine shaft using a Robbins 81-R drill. An overview of the raise boring process is presented to illustrate the ventilating and gas testing procedures. A mine shaft project, 4.88 m (16 ft) in diameter and approximately 457.2 m (1500 ft) in length, will be the focal point of the presentation, with reference to similar projects for additional comments on variations in procedure.

Jan 1, 1980

By K. J. Shoop, S. J. Ripke, T. C. Rao, T. C. Eisele

This study investigated the use of a 7.6-cm inside diameter, 344.2-cm-tall horizontally baffled column to reduce the ash content of a low-sulfur bituminous coal. Tests were carried out to determine how the baffled column responded to changes in feed rate, reagent dosage and feed quality. The tests were also carried out to compare the columns performance to a conventional column installation in an operating plant. In this paper, the results of this investigation are reviewed and discussed.

Jan 1, 1999

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Phillip Crowson

The annual surveys of corporate exploration expenditures carried out by Metals Economics Group (MEG) show that estimated total spending rose to a record $13.2 billion in 2008 ($14.4 billion including expenditure on uranium exploration). This compares with the previous peak of $5.2 billion in 1997. MEG does not adjust its figures for the effects of inflation but recognizes that the substantial increase in budgets of recent years has not been reflected in a proportionate rise in actual activity on the ground (MEG, 2009). In recent years, costs were boosted by rising prices of fuel and other inputs, and acute shortages of qualified staff and drilling rigs. The inflation rate for exploration substantially exceeded increases in general price levels as measured by national indices. Just how much exploration costs in-creased, however, is probably not widely appreciated. Figure 1 compares the MEG estimates of total annual dollar spending in Australia, Canada and the United States combined with their drilling activity, measured in meters drilled, as reported by national agencies. These three countries accounted for 39 percent of MEG?s estimate of total spending on minerals exploration in 2007 and 40 percent in 2008. Expenditure is shown both in money terms, as reported by MEG, and adjusted to real terms, with the U.S. implicit gross domestic product (GDP) price deflator. The comparison is of index numbers based on 2000 as 100. Comparisons for the individual countries are shown in Fig. 2 to 4. National figures for drilling activity and expenditures cover some minerals like iron ore, coal and industrial minerals (29 percent of Australian spending in 2007 to 2008 and 10 percent of Canadian in 2008) that are not included in the MEG estimates of spending. Such exploration has risen more than the total since 1997 so that the index of drilling activity in the figure may be overstated for recent years on a like for like basis with the MEG data. MEG only includes spending on uranium exploration in recent years whereas the national agencies include it throughout. Figures 5 and 6 com-pare national and MEG estimates for Australia and Canada. No comparable figures are available for the United States. The national estimates have been converted to U.S. dollars at the appropriate annual average exchange rates, and the Australian annual totals have been summed from the published quarterly data. Not only is spending cyclical, but so are unit costs. Australian and Canadian data show that total exploration spending per meter drilled rose by 2.5 to 3 times between 2001 and 2008. These rises may partly reflect increasingly complex drilling methods, and an increased share of assaying and testing in total exploration costs. In Canada, however, diamond drilling accounted for a lower footage of the total in 2005-2007 than in the preceding few years, and spending on exploration, as opposed to deposit appraisal, made up a higher proportion of the total. Figure 7 compares index numbers of exploration spending per meter drilled in U.S. dollars for Australia and Canada. These figures raise a number of questions. First, is the relationship between total expenditure and meters drilled in the three countries for which data are available typical of the global experience? Secondly, has there been a significant change in the make-up of exploration spending in recent years? The Canadian data do not suggest that there has been, but is Canada typical? Has drilling become less important relative to other exploration techniques, or is it just as important as ever in confirming the existence of viable ore deposits?

Jan 1, 2009

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

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

Jan 1, 1995

By A. B. Cecala, G. W. Klinowski

Introduction The US Bureau of Mines has designed and evaluated total mill ventilation systems (TMVS) at two mineral-processing operations. Both systems have proven proficient at cost-effectively reducing respirable dust levels throughout the mills. An 11.8-m3/sec (25,000-cfm) system installed at a clay-processing mill provided about 10 air changes per hour (acph). This system reduced respirable dust concentrations by about 40% throughout the mill building. The second evaluation was performed at a silica sand operation. Tests were performed with 23.6 and 47.2 m3/sec (50,000 and 100,000 cfm) of ventilation to the mill building, corresponding to 17 and 34 acph. Average millwide respirable dust reductions were 36% and 64%, respectively. This work was done to determine general guidelines and design criteria for operations that may be interested in implementing TMVS. TMVS should be designed to draw clean make-up air in near the base of the mill. This air provides general purging and may clear some dust-laden areas throughout the mill structure. It is discharged at or near the top of the structure. Without the TMVS, dust can exit the building through open doors, windows, etc. and contaminate outside plant workers. This dust can also be recirculated back into the mill building causing additional contamination problems. TMVS is not a stand-alone system. It is a supplemental technique to assist other localized dust control systems (Mody et al., 1987). Testing Total mill ventilation systems were evaluated at two processing operations. For both tests, the primary goal was to determine the reduction in respirable dust throughout the structure. Test site No. 1. The first evaluation was performed at a clay-crushing and screening mill in New York. The mill's base dimensions were 39.6 x 9.8 m (130 x 32 ft), giving a volumetric capacity of 4250 m3 (150,000 cu ft). The TMVS was designed to provide 12 m3/sec (25,430 cfm) of ventilation air to the mill, representing about 10 acph. This was accomplished by three 4 m3/sec (8475 cfm) roof-powered exhaustors that were evenly spaced across the roof. Three wall louvers provided inlets for make-up air at the mill's base. Two evaluations were performed at this mill. The first occurred for two days in December when outside air temperatures ranged from -12.2° to 4.4° C (10° to 40° F). Temperatures with the wind chill factor dropped as low as -34° to -40° C (-29 to -40° F) during the early morning. The second evaluation was in April when outside air temperatures ranged between 10° to 26.7° C (50° to 80° F). These two evaluations verified the effectiveness of TMVS in lowering respirable-dust concentrations throughout the mill. Gravimetric-sampler respirable-dust concentrations in December ranged from 0.22 to 2.39 mg/m3 under normal conditions, compared to 0.13 to 1.55 mg/m3 with TMVS. In April, respirable dust concentrations ranged from 0.29 to 4.84 mg/m3 under normal conditions compared to 0.21 to 2.37 mg/m3 with the system on. TMVS also increased visibility throughout the mill building. Dust reductions for gravimetric and RAM-I instruments are listed in Table 1. Each value was determined by comparing the average concentration for the entire day of testing for normal and TMVS values. The mean/standard deviation values for gravimetric and RAM-1 results were 42.5/17.6 and 37.1/16.8, respectively.

Jan 1, 1996

By Thomas S. Kovach

In today's competitive energy market, every effort must be made to assure the lowest possible production cost. 'rhi s. paper describes the cooperative effort between the coal producer and the utility customer to develop and install a fully integrated mine information management system at a 2.5 million tpy lignite mine in east Texas. A description of each item of the modules is included along with the problems encountered during development and how they were corrected. The system configuration is also described.

Jan 1, 1989

By D. F. Doerr, E. Blalock, D. R. Bush, K. D. Cohen

"BACKGROUND The design, development, fabrication, and test of a cryogenic supply for a refuge alternative is part of a larger effort funded by the National Institute for Occupational Safety and Health and managed through a Space Act Agreement with the National Aeronautics and Space Administration at the Kennedy Space Center, FL to create cryogenically based life support equipment responsive to the needs of miners as required by the Miner Act of 2008. This work is done by a joint contractor/NASA team. In particular, the Cryogenic Refuge Alternative Supply System (CryoRASS) serves to provide a respirable air source to miners confined for 96 hours in a mine refuge alternative. This system stores liquid air in a zero-loss cryogenic reservoir (Dewar) until an emergency requires use. The system does not require electric power when in use by miners. Because of the very cold nature of the liquid air, considerable heat stress relief and reduction of humidity is provided in addition to the required replenishment of oxygen. Moderate dilution of CO2 also occurs.INTRODUCTION The design of this new cryogenic refuge alternative supply system (CryoRASS) uses NASA/Kennedy Space Center technology as a development baseline. The design stores cryogenic (liquid) air that is warmed in an air handler box and the resulting gaseous air is introduced to the chamber that not only causes air circulation, but relies upon environmental and human heat to expand the liquid air to a gaseous state. Therefore, the primary benefits of the liquid air source, besides atmosphere maintenance (oxygen), is refuge cooling and dehumidification. This results in reduction of human heat stress and this occurs without the need for electrical power in the refuge chamber. This reduction is a major factor in ""apparent heat"" and the human perception of comfort. Another benefit with introduction of ""air"" vs only ""oxygen"" is the dilution of carbon dioxide. The volume of new air will displace existing refuge air and this air laden with humidity/CO2/low O2 passes out of the refuge relief valves. These RVs should be strategically placed to vent stagnant air preferentially. An initial system test was conducted in our Florida facility in a production 10-person hard chamber for 96 hours. Last year (2014) the system was tested with a 23 person inflatable refuge alternative in Bruceton. For this test, the CryoRASS air handler was redesigned to offer superior characteristics when installed in a Built-in-Place refuge alternative. This report describes the results of the modified system being tested in the Pittsburgh Experimental Mine while attached to a 30 person refuge for 96 (+/-) hours."

Jan 1, 2016

By L. M. Anisdahl

Ventiflex is a flexible ventilation duct system with a global market within mine and tunnel ventilation. The system is distributed in a wide range of qualities and dimensions. It is very important that ventilation systems are designed correctly in mines and tunnels. An incorrectly dimensioned system might result in bad air quality and too high costs. The objective of this study has been to assess some important design parameters for ventiflex comprising the friction loss factor for the duct wall itself, as well as the shock loss factors and the leakage factors for the different couplings systems. This has been obtained by carrying out a large scale experimental program on ducts with all the three different couplings: Velcro, zip and steel-ring. Pressure profiles and air flow rates have been measured, and based on these data the different factors have been calculated. The following results have been obtained: A friction factor l = 0,012 (K = 10) for the duct wall, a shock loss factor z = 0,0125 for velcro- and zip couplings, a shock loss factor z = 0,005 for steel-ring couplings, and a leakage factor f* = 0 mm2/m2 for all three coupling types. The latter result proves that no leakage is possible for a correctly installed duct with proper alignment. The results also confirm that ventiflex complies with class S, or better, according to the SIA-norm. However, whether ventiflex should be designed as class S, A or B according to the SIA-norm, depends on the particular application.

Jan 1, 2009

By Luisa Alfieri, Giorgio Fantauzzi, Sergio Notarianni, Spencer Cavagnet, Alessandro Damiani

"The new tunnels of the Copenhagen ""Cityringen"" will be excavated, at Kongens Nytorv square, crossing over the existing M1 and M2 metro lines. After the crossing, the tunnels will underpass the historical building of the Magasin du Nord and adjacent masonry buildings, with a minimum tunnel cover of about 4.5 m. The tunnels have an internal diameter of 4.9m and are being excavated through earth pressure balanced Tunnel Boring Machines in mixed face.The Magasin du Nord estate is characterized by a very complex structure, with different blocks built in different times. An assessment of the condition of the building and its sensitivity to tunnellinginduced ground movements has been conducted through a fully coupled soil-structure interaction analysis; based on the results of this analysis, mitigation measures have been designed.INTRODUCTIONThe Copenhagen Metro Cityringen is expected to be running in July 2019 and will consist of two single tracks in a twin tunnels, each approximately 16 km in length, for transporting up to 243,000 passengers on an average work day. The Cityringen includes 17 underground stations with island platforms, 4 crossover facilities and 3 construction and ventilation shafts. The tunnels is being excavated using earth pressure balanced Tunnel Boring Machines and have an internal diameter of 4.90 m.The intersection with the existing M1 and M2 metro-line will take place at Kongens Nytorv and at Frederiksberg. According to the design alignment, in Kongens Nytorv the new tunnels will pass over the existing M1 and M2 ones; as a result of this, the tunnels will underpass the historical building of the Magasin du Nord and the adjacent masonry building hosting a Mc Donald’s with a minimum cover of about 4.5m.In the framework of Risk Assessment for sensitive buildings, the effects of tunneling-induced ground movements on the Estates ØS0437 (McDonald’s building), ØS0428 (Magasin du Nord) and ØS0004 (Hviids Vinstue) were evaluated. To achieve this goal a fully coupled soil-structure interaction has been considered and the entire excavation process has been reproduced in a 3D model where the abovementioned buildings have been modeled. In this way, a risk assessment analysis was performed evaluating directly the effects of the excavation in terms of settlements and variations of the internal forces as far as the excavation process proceeds. In figure 1, a picture of the construction of Kongens Nytorv station, with Magasin du Nord and Mc Donald’s building on the background, is presented."

Jan 1, 2016

By Frédéric Flament, Daniel Hodouin, Claude Bazin

A control strategy is proposed to meet the following specifications of flotation control: minimize the variations of concentrate and tails grades and recovery while limiting the amplitude of variations of the manipulated variables. The strategy consists of determining future control actions which optimize a performance criterion over a limited time horizon. The desired dynamic behaviour of the controlled flotation plant is specified using reference models. The strategy is tested on a 2-computer plant simulator working in an environment which emulates the industrial conditions.

Jan 1, 1993