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By F. Varley

Researchers from the National Institute for Occupational Safety and Health (NIOSH), in cooperation with mine operators, conducted a study of heat stress exposures among mine rescue workers in underground mines. Mine rescue workers face extreme heat loads as they enter hot and poorly ventilated environments, particularly when they are wearing breathing apparatuses. The burden of wearing a closed-circuit breathing apparatus (CCBA), the inability to drink fluids for extended periods of time, and the potential for ventilation disruptions and fires combine to increase the risk of heat illness during a response to an emergency. In this research, ambient environmental conditions and heat strain indicators were measured using conventional ventilation monitoring tools during mine rescue training exercises. Heat strain was measured with an ingestible temperature-sensing pill that measured core temperature continuously. In addition, a heart-rate monitoring chest strap was used to indicate physical strain resulting from physical activity and heat. Both core temperature and heart rate data were trans-mitted to a remote recorder worn by a rescuer and time-stamped. Team activities were also observed to determine the contribution of work rate to a person’s total heat load. The effectiveness of several engineering controls, such as cooling the air inhaled by rescuers through an apparatus, wearing cooling vests, and supplying water through the CCBA masks, were investigated. Administrative controls, such as limiting the duration of exposure and controlling the work rate through forced-rest regimens, were also evaluated. Results to date are discussed, and tools for evaluating team condition and estimating the length of time it would be safe for rescuers to work are proposed.

Jan 1, 2004

By Bruce D. Hanson

A computer model capable of simulating the basic production system in a variety of underground mining methods has been developed at the Twin Cities Mining Research Center. This Bureau of Mines paper describes the simulation of the basic production system at the working face, which includes drilling, blasting, mucking, and roof bolting. The model is event-oriented; that is, the status of the system is updated at the occurrence of an event. All event or activity times, including failure and repair of equipment, are treated as stochastic variables. A location dimension, in addition to the time dimension, is used, allowing the simulation of multiple location systems. To aid the user in determining the efficiency of his system, the model provides several reports. An event or activity report gives the frequency of occurrence total active time, and cumulative delay statistics for each event. An equipment report delineates the availability and utilization of all equipment. A production report provides information on various performance measures such as tons produced and holes drilled. The cost report gives the cost of various elements in the system. Operating data from an actual mining situation were used in validating the model. In addition to the description of the model, examples of simulation experiments which were performed on a room-and-pillar mine system are given. Use of the model to determine methods for improving the efficiency of the system is discussed.

Jan 1, 1974

By Chi-Ming Chang

Alloy powders are commonly used in the purification of zinc sulfate solutions for use in standard electrowinning practice. The primary goal of this research was to study the use of zinc-lead alloys for similar reactions in the galvanic stripping of impurities from organic solutions. The effects of various parameters on the electrochemical behavior of commercially produced zinc-lead alloy powders were determined. A qualitative evaluation and summary of the influence of the more pertinent parameters on iron removal from di-(2-ethylhexyl) phosphoric acid (D2EHPA) by galvanic stripping is given. Water concentration, oxygen content and reaction temperature were the principal factors affecting the reaction rate and process efficiency. Activation energies were also calculated and indicated that the reaction changes from chemical to diffusion control .with increasing temperature in the range of 35" to 45°C. The results showed that zinc-lead alloy gave faster reaction rates with less sensitivity to process variables than SHG zinc as a reducing agent for galvanic stripping. The role of lead could be as cathodic sites for the reaction which enhances the zinc dissolution or anodic half reaction. The morphology of the zinc-lead alloy surfaces, before and after the reactions, supports this assumption.

Jan 1, 1998

By LeEtta M. Shaffner, John R. Marks

INTRODUCTION In the last twelve years, underground stoping methods at the Homestake Gold Mine have evolved from open cut-and-fill with jacklegs, cribbed raises and electric slushers to ramp-based mechanized cut-and-fill (MCF) and vertical crater retreat (VCR) with dieselpowered loaders, trucks and drill jumbos. The evolution was swift. Unfortunately, ventilation practices have had trouble keeping pace. Stope ventilation is still too dependent on auxiliary fans and coolers. A single large MCF stope might contain up to eight headings, each of which requires significant ventilation resources when a diesel loader is present. Air is shortcircuited all too frequently through mined-out VCR panels. These factors and the recent rapid decent of the center of mining prompted a study to improve ventilation practices and to more accurately project future requirements. BACKGROUND In 1992, the underground portion of the Homestake Gold Mine produced 8336 kg of gold from 1407 ktons of ore (268,000 oz from 1,551,000 st). The ore is hosted in low to medium grade meta-sediments. Stoping took place from 420 to 2347 m below the surface (1400 to 7700-ft levels). The weighted center of mining was 1662 m deep (5450-ft level) in 1990 and is projected to be 1890 m (6200-ft) in 1993. The deepest level is 2440 m (8000-ft) where the virgin rock temperature (VRT) is 56.1°C (133F). The mine is ventilated by 504 m3/s (1,069,000 cfm) measured at mid-exhaust-circuit density. The air-conditioning system includes an 8.1 MWR (2300 ton) controlled recirculation plant, a 2.0 MWR (580 ton) chilled water plant, a 1.0 MWR (290 ton) exploration drift refrigeration plant, 28 spot-coolers totaling 3.4 MWR (960 tons) and 35 spray coolers totaling 1.5 MWR (420 tons). The mine employs 117 diesel units with a total nameplate rating of 7961 kW (10,672 hp). These units include twenty-four 1.5m3 (2-yd) loaders, twenty-six 2.7m3 (3.5-yd) loaders, fourteen 3.8m3 (5-yd) and two 7.6m3 (10-yd) trucks, and assorted utility vehicles and drill jumbos. THE STUDY In April 1991, the University of Nevada-Reno (Mackay School of Mines) and the South Dakota School of Mines and Technology cooperated with Homestake on an MCF study. Mackay instrumented a stope with thermocouples, air velocity meters and hygrometers (Duckworth, 1992). South Dakota Tech conducted a finite-elements computer analysis of a back-filled MCF stope with/without light-weight shotcrete insulation on the sidewalls and back (Chellam, 1992). Homestake conducted an empirical analysis of deep-level MCF stoping. This paper describes the Homestake study. Twenty-three of the forty MCF stopes deeper than 1800 m were surveyed at least once during the last quarter of 1992. The stopes not included were being cable-bolted or back-filled. Figure 1 shows the complexity of one of the ramp-based stoping areas included in the survey. This particular stope has six separate production headings. A baseline wallrock heat load was derived for each stope or heading from psychrometric calculations. Broken ore, waste rock fill and fissure water were noted when present and the effects included in the baseline heat load. Other heat sources often mentioned in the literature such as metabolic heat from workers, heat from explosives and small electric loads were neglected. Fan heat was considered part of the ramp & crosscut heat load and thus not included in the study. Diesel heat was addressed separately. RESULTS Survey results, plotted as the heat flux against VRT, are shown in Figure 2. The equation for the regression line is: W/m2 = 2.1236*VRT - 75.405 The 0.31 correlation coefficient is poor which implies that the results should be used cautiously. Previous experience strongly suggests that differences in productivity are most likely responsible for the scatter in data points. A rapidly advancing stope will have a

Jan 1, 1993

By G. H. Kesler, J. E. Oberele, C. E. Dryden, J. H. Oxley

Heat-transfer rates were measured in a model of a multifilament vapor-deposition bulb fo the preparation of high-purzty metals. Local transfer coefficients for heat transfer frow the filaments to the circuates process stream were determined as a functiol of gas flow rates and bulb, inlet, and filament, geowzetvies. These results a-e then convertted to n mass-t7,ansfer basis to provide a method of correlating vapor-deposition rates and to predict the performance of commercial urzits. The firm1 cor?,elating equation for mass transfer was found to be- wheve k, is the mass-transfer coefficient; D, is bze gas difficsivity : P is the system pressure; R is the gas constant; T is the gas temperature; Df, DB, and Di ave tlze diameters of the filament, bulb, and inlet, respectively; k is the gas conductivity; C, is the specific heat of the gas, M is the wzolecular zleight of the gas; Npr . N,,, NRe a7'e the Prandtl, Schmidt, and Reynolds numbem of the system, respectively; LB is the height of the bulb; and SB and Sf are the total suyiace area of the bulb and filaments, respectively. ONE of the more conventional methods of preparing very high-purity metals is the vapor deposition of the metal upon heated filaments. The reactions which can be employed in such a process generally fall into two classes: 1) Pyrolysis of a volatile metal compound. Generally because of their instabilities, metal iodides are used; however, other metal halides and hydrides have been employed. These processes are frequently carried out under vacuum conditions, and the filaments are maintained at high temperatures. 2) Hydrogen reduction of a volatile metal compound. A metal chloride is normally the preferred feed for these processes, but other halides are sometimes used under special conditions. These hydro- gen-reduction processes are usually operated at atmospheric pressure. Many of the current processes for the production of semiconductor materials are actually vapor-deposition processes. However, the vapor-deposition process itself is relatively old, and almost all metals can be prepared in a state of very high purity by the use of vapor-deposition techniques. Loonam has recently given an excellent review of the iodide process for metal deposition,' and Owen has started some detailed studies with a carbonyl system.' A more general treatment of all vapor-deposition processes has been outlined by Powell, Campbell, and Gonser.3 Despite the fact that vapor-deposition techniques have been employed since 1890, there is a surprising lack of fundamental information regarding the transport characteristics of even the simplest deposition system, the heated filament. The purpose of the investigation described in this paper was to extend some of the earlier data obtained at Battelle to predict deposition rates and the uniformity of deposition in large-diameter, multifilament deposition bulbs under forced convection conditions.4'5 A number of other investigators have already presented a simplified treatment of the deposition process under non-flow conditions, i.e., pure diffusion.6"9 To obtain information on the transport characteristics of a deposition bulb, a model of a typical plant deposition unit was constructed, and local heat-transfer coefficients from the filaments to a circulating air stream were determined under various conditions of air-flow rates and bulb, inlet, and filament geometries. These results were then converted to a mass-transfer basis, and consequently provided a method to correlate experimental vapor-deposition rates and to design commercial deposition units. EXPERIMENTAL WORK Description of Model. The apparatus which was used in this work was a scaled model of one type of deposition bulb. It was constructed of Lucite and was easily altered to give several combinations of diameter of bulb liner, inlet size and location, number of filaments, and outlet geometry. The model is shown in Figs. 1 and 2. The bulb shell was made octagonal for convenience in construction and observation, and the bulb liners were cylindrical. Air, introduced through one of several different jet inlets, was used as the test fluid in these model studies. Filaments for deposition of metal were simulated by metal rods 1/4-in. in diam. Two of the rods could be heated by passing electrical current through them ¦ while the remaining rods could not be heated. This procedure minimized the amount of heat which was

Jan 1, 1962

By Ze-Qiang Xie, Bai-hui Wang, Feng Chen, Ling-zhi Yang, Yu-feng Guo

In this paper, based on the big data in steelmaking process, equation about the relationship between steel composition (especially the oxygen content) and the element yield is established by using mathematical analysis method to analyze the factors influencing yield. Meanwhile, combining the alloy yield data-base that were established based on self-learning of historical data and computer technology in steelmaking process, a method of yield prediction of alloy elements is studied. To provide support for this method, alloying elements yield prediction software is developed to realize the rapid and intuitive site operation guidance, and to realize the precise control of dosage of alloy material.

Mar 1, 2017

By W. B. Berwald, T. W. Johnson

"The United States Bureau of Mines, in cooperation with the Natural Gas Department of the American Gas Association, has been studying actively for the past several years the flow of natural as through high-pressure pipe lines. 4 This report deals with that part of the study involving the design of pipe-line systems consisting of lines laid parallel to each other. The lines are referred to as the original line and the parallel line. The paralleling lire may or may not extend the entire length &f the original line; however, the parallel line is connected at its two ends to the original line.As a result of the large expansion of the natural-gas industry almost all markets within economic distance, from the developed as fields have at least one pipe line supplying them with natural gas. Small increases in the consumers' demand for natural-gas can be met usually by increasing operating pressures, but any marked increase in demand necessitates paralleling sections of the existing systems. It is believed that to a large extent future expansion in the natural—gas industry will consist of paralleling the present lines. The several formulas given in this report have been developed to aid in the design of such systems. Curves and tables were prepared from these formulas to simplify the calculations materially and to facilitate determination of the sizes of pine to be used in paralleling various diameter lines to give the most economical design. The formulas are presented without detailed derivations, as a complete report on the flow of natural gas through pipe lines is being prepared which will contain the derivations of these formulas. Examples included here show the detail of application of the several formulas to problems of design of parallel lines."

Jun 1, 1934

By E. Kamon

This report summarizes completed research pertaining to the reaction of human subjects exposed to Self-Contained Breathing Apparatus (SCBA) stressors during escape conditions (viz., resistance breathing, C02 breathing, hot air breathing, and static positive facepiece positive pressure breathing) and the development of an automated breathing and metabolic simulator (ABMS). Volume I contains summarizes of the studies performed under this contract. Volume II is an appendix which contains more detailed descriptions of these studies.

Jan 1, 1984

By David Schlain

"INTRODUCTION Electrolytic manganese is produced by electrolysis of a solution of manganese sulfate and ammonium sulfate in a diaphragm-type cell, metallic manganoese being deposited on the cathodes and sulfuric acid produced at the anodes. The spent electrolyte is used to leach fresh ore that has previously been subjected to a reducing roast; the resulting leach liquor is purified and returned to the cell.The leach liquor produced at the Bureau of Mines pilot plant 3/ at Boulder City, Nev., contains 135 to 140 grams of ammonium sulfate and 32 to 36 grams of manganese (as sulfate) per liter, and has a pH of about 2.5. After neutralization to a pH of 6.4 to 6.6 with ammonia and thickening, 0.10 to 0.15 gram of hydrogen sulfide per liter is added to the liquor to precipitate the heavy metals, and after 15 to 30 minutes the precipitate is filtered off. Residual sulfides and colloidal material are removed by ferric hydroxide formed by adding a small amount of ferrous sulfate to the solution at a pH of 6.5 to 7.0 and passing air through it. The solution is then filtered and clarified as a precoat filter. Finally, 0.1 gram of sulfur dioxide per liter is added, and the solution is electrolyzed in an anode-in-diaphragm cell. The spent electrolyte coming from the cell contains approximately 10 grams manganese, 135 grams ammonium sulfate, and 45 grams sulfuric acid per liter."

Oct 1, 1946

By David L. Veith

Caustic scrubbing followed by silica flotation with cationic ether primary amine proved effective in upgrading commercial magnetic concentrates to supergrade products containing 2 percent SiO2, or less. Samples representing 41 million long tons per year of commercial magnetic concentrates from the Lake Superior region were tested by the Bureau of Mines. Bench-scale and pilot plant research indicated an estimated potential of 32 million long tons per year of superconcentrates to be available. Two pilot plant investigations yielded superconcentrates containing 1.1 and 1.9 percent SiO2 that were derived from commercial concentrates containing 7.0 and 4.2 percent SiO2, respectively. Corresponding iron recoveries were 90 and 93 percent at reagent costs of just over 33 cents per long ton of superconcentrates produced. Additional sources evaluated on a bench scale resulted in iron recoveries of 72 to 95 percent in products containing 1.7 to 2.0 percent Si02. Although reagent cost ranged up to 60 cents per long ton of superconcentrate, it is probable that these costs would be reduced in plant applications where benefits are derived from material and reagent recycle.

Jan 1, 1974

By Martti Hagfors, Mervi Hokkanen

The Finnish Defence Forces has destroyed obsolete and spoiled ammunition in open surface mass detonations at Hukkakero fell, in municipality of Kittilä, in Lapland since 1988. The de-struction method was learned from Sweden in 1987. The disposal camp is organized every au-tumn and it lasts four to six weeks. A gross weight of an average charge varies from 150 to 250 tons and the net weight of explosives from 15 to 25 tons. As a booster there have been used 2 to 3 tons of emulsion explosives. The explosives mainly consist of trinitrotoluene (TNT) containing smaller a mounts of amatol and hexotol. The Finnish Defence Forces Tech-nical Research Centre (PVTT) has studied the environmental effects of these mass detona-tions since they began in 1988.

Jan 1, 2011

By George E. Fish

This report describes investigations conducted by the Bureau of Mines from 1957 through 1959 in the Roseland anorthosite-titanium belt of Nelson and Amherst Counties, Va., for the purpose of studying the titanium resources of the area. Three titanium deposits containing much saprolite ore derived from the decomposition of anorthosite and diorite were selected for detailed study Following a reconnaissance of the reported titanium occurrences in the belt. These three deposits occur in or are associated with anorthosite. Most of the occurrences are situated on or near the contact between the anorthosite and the gneissic country rock. The saprolite deposits are as much as 1.5 miles long, 1,000 feet wide, and 100 teat deep and contain significant reserves of titanium-bearing material estimated in excess of 20 million tons. Where it was penetrated by drilling, the unweathered rock beneath the saprolite contains nearly the same titanium values as the saprolite.

Jan 1, 1962

By Kenneth M. Maas

The Bureau of Mines inventoried the land in southwestern Alaska to classify ownership and availability and compared this land with known mineral terranes and mineral deposits. This report discusses and quantifies the potential for favorable location of mining claims for metalliferous (locatable) mineral deposits. Of the nearly 24 million acres that constitute the study area, approximately 3.6 million acres or 17% are available for mineral location, 3.4 million acres or 14% are available with restrictions, and 17 million acres or 71% are unavailable. Mineral terranes are known to underlie about 7.8 million acres or 33% of the study area; 3 million acres or 38% occur on land available for mineral location, 1.3 million acres or 17% occur on land available with restrictions, and 3.5 million acres or 45% occur on unavailable land. The State of Alaska has opened over 91 % of its land in the study area to mineral location; in comparison, the Federal Government allows mineral location on 0.07% of the land under its jurisdiction. There are 117 known mineral deposits or occurrences within the study area, including 77 lode deposits and 40 placer deposits. Of the 41 lode deposits occurring within a mineral terrane, 2 occur on available land.

Jan 1, 1988

By Ludwig Schmidt

The consumption of gasoline in the United States varies materially with the seasons; during some of the summer months it exceeds the rate of production and in the other seasons the production of gasoline exceeds the consumption, The usual practice of the larger refiners is to store surplus stocks during the winter season in order to meet the heavy demands during the summer months. The following table shows the stocks of gasoline at refineries on the last day of April for the past five years: 1919 593,616,170 gallons 1920 643,5526644 " 1921 747,222,900 " 1922 892,567,766 " 1923 1,336,417.871 "

Jan 1, 1923

By D. P. Lindroth

The thermal diffusivity above 298 K, up to 400 K and at pressures from ambient down to 10-10 torr was measured for the following six igneous rock types: tholeiitic basalt, obsidian, rhyolite, granodiorite, gabbro, and dunite. The thermal diffusivity data were obtained by the laser flash method. Extreme values range from a low of 5.3X10-3 cmF/sec for rhyolite at 383 K and 2X10-7 torr to a high of 13.4x10-3 crrf/sec for dunite at 298 K and 760 torr. Data are also included on a commercial fused silica, which was used as a standard.

Jan 1, 1974

By Sheldon P. Wimpfen

EVER since the Phoenicians roamed the known world in quest of metals to harden their helmets and precious metals and gems to adorn their ladies, many other nations have sought metals in the far corners of the earth. During the past 75 years American engineers have been among the foremost in seeking for and developing ore deposits. They have been responsible for many of the principal discoveries and for the application of modern methods to mines in the far comers of the earth. American mining engineers with a zest for travel and an insatiable appetite for metals have also introduced throughout the world many articles made of metal which we consider essential to everyday life. Production and consumption of these metals in many foreign countries has brought to these countries an increased standard of living and added tax revenue to their governments. But unfortunately engineers from this country are no longer given ready access to all the countries of the world. This is not only unfortunate for American engineers but also for the countries that have adopted the exclusion policies and especially so since time has proved that in countries where the American engineers have gradually been excluded there has been a significant decline in the progress of the mining industry. Removal of these barriers to professional men would aid immensely in the world-wide development of the mineral industries. New areas brought into production under the guidance of American engineers with their exceptional know-how would provide more jobs for nationals of the countries involved and a better standard of living for the nations as a whole.

Jan 1, 1946

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Doug Harding

The Metropolitan St. Louis Water District’s Project Clear is a 28-year program targeting wastewater capacity throughout St. Louis, MO. The extensive program includes the Jefferson Barracks Tunnel, running parallel to the Mississippi River at 2 m ID and 5,400 m long. A rebuilt Robbins Main Beam TBM began boring but hit challenging conditions about 2,400 m in, including water inflows. A larger 4.1 m Robbins Main Beam machine will complete the remaining 3,050 m in limestone, dolomite, and shale. The TBM will utilize a high-powered, high breakout torque cutterhead along with enhanced 360-degree probing & grouting capabilities to detect any karst formations.

Dec 1, 2022

By Christophe Bragard, Jason Choi, William H. Hansmire

This case history provides a physical benchmark of tunneling technology progress comparing an old hand-mined tunnel to a modern 20-ft-diameter EPB TBM that mined in the same ground. This rare opportunity to see how tunnels were constructed many years ago occurred when an old 10-ft-diameter storm drain tunnel in the way of os Angeles subway construction was demolished and replaced with obas berglass pipe. Steel rib and lagging initial support was exposed showing solid-sawn 2" × 12" × 48" redwood lagging. This paper documents the as-built old tunnel, how it was successfully replaced, and some lessons learned.

Jun 23, 2024

By Jos. Kalbhenn

PART of the plan of re-allocation evolved by the engineers of the Dominion Steel and Coal Corporation in 1932 to retard the rapid advance of the submarine workings seawards was the opening of Dominion No. 20 Colliery, in the Harbour seam, by a pair of cross-measure tunnels from the workings of an older mine working the Phelan seam, which lies 430 feet below the Harbour seam. The tunnels, 8 ft. by 13 ft. in size, were started away from the old workings of Dominion No. 2 at a point 1.5 miles beyond the shore line and were driven on a rising grade of 1 per cent for a distance of 5,335 feet, tapping the Harbour seam at a point 2 ¼ miles off-shore. For transportation and ventilation, shafts and roadways already existing in the Phelan seam were made use of during the stages of development and up to the present time. One of the two tunnels served as a main haulage road and an intake airway, and the other as a return airway and travelling road. However, as the workings of this new colliery expanded, the need for additional airways arose as was anticipated when the plans for the development were formulated. To meet this need, work was started on October 1st, 1942, to establish a new opening between Nos. 2 and 20 Collieries in the form of a vertical shaft to provide a passageway for the ventilating air current of some 250,000 cubic feet per minute on its return to the upcast shafts in No. 2 Colliery, thus making both tunnels available for intake airways. In the interest of safety, speed, and economy, the shaft was raised from the Phelan to the Harbour seam. It is 20 ft. by 12 ft. in size and, with the exception of three timbered compartments across the narrow end, is untimbered. One of the three timbered compartments housed the ventilating tubing and pipe lines, the centre one served as a cage-way for hoisting men and material during the operation of raising, and the third one contained a ladder way. The rest of the shaft was filled with debris during the period of shaft raising.

Jan 1, 1944