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The total crude production from the Dutch East Indies, and the British Protectorates of Sarawak and Brunei for 1932 is given in the following table: Production, which had shown a rapid increase from 1921 to 1929, has since been reduced. Exploitation drilling in the majority of the larger fields was suspended and existing production kept within predetermined limits by beaning. Most of the smaller fields were shut down entirely. The depth of producing levels varies widely in the different fields. Shallow fields with drilling depths to 500 meters (1650 ft.) are still numerous as well as those with depths between 500 and 1000 m. (3300 ft.). A few of the newer fields reach greater drilling depths up to 1600 m. (5250 ft.). The deepest exploration well drilled to date (Djambi field, South Sumatra) reached a, depth of 2332 m. (7650 ft.) before abandonment in March, 1930. The specific gravities of the various crude oils produced range between 0.65 (86.2° A.P.I.) and 0.95 (17.5° A.P.I.). The lighter crudes have a high gasoline content and fuel oils are derived from the heavier crudes, which usually are rich in paraffin. Producing companies are: De Bataafsche Petroleum Maatschappij (Royal Dutch-Shell combine in Dutch territory); The Anglo-Saxon Petroleum Co. Ltd. (Royal Dutch-Shell combine in British territory); De Nederlandsch-Indische Aardolie Maatschappij (De Bataafsche Petroleum Maatschappij-Dutch East Indian Government combine); De Nederlandsche Koloniale Petroleum Maatschappij (Standard Oil Co. of

Jan 1, 1933

United Engineering Trustees, Inc. GEORGE D. BARRON ARTHUR S. DWIGHT R. M. ROOSEVELT The Engineering Foundation H. C. BELLINGER E. DEGOLYER J. V. N. DORR

Jan 1, 1932

In the absence of. J. B. Marks, President, the meeting was called to order by George B. Pryde, Past President. After organization, the meeting was turned over to W. J. Reid, Vice-President for Utah. A moving picture, "The Story of Dynamite," was shown by the courtesy of the E. I. DuPont de Nemours Company. This provoked a great deal of discussion and many questions were asked A. E. Anderson, who described the paper. John M. Jennings read a description of the United States Fuel Company's new Hiawatha tipple, after which the meeting adjourned to Hiawatha for lunch and a visit to the tipple and outside workings of the Hiawatha mine. Thursday, August 27, 1925 Meeting called to order at 9:30 by Vice-President W. J. Reid. Papers were read on the subject of "Mechanical Loading," by A. W. Dickinson, 0. G. Sharrer, and the paper of Glen A. Knox was read by the Secretary. General Superintendent George A. Murphy, of the Spring Canyon Coal Company, read a description of their newly completed tipple. The meeting then adjourned for Spring Canyon where lunch was served, and the afternoon spent in viewing the new. tipple. At 8:00 p. m., D. Harrington read a paper on "Rock Dusting," L. C. Clare read a paper on "Mudite's Relation to Nature," and at 11 p. m. the meeting adjourned. Friday, August 28, 1925 Meeting called to order at 9:30 a. m. by Vice-President W. J. Reid. The discussion of preceding papers continued. Paper by Joseph Harrington on "Domestic Fuel Stokers" was read by the Secretary. The expected paper by Mayne Read was not presented as Mr. Read was unable to get to Price on account of railroad conditions. The papers on the "Theory of Spontaneous Combustion," by C. P. Crawford, were also omitted through an error of the Secretary, and particular attention is called to the publication of these two articles. The new tipple of the Columbia Steel Company was described by T. C. Harvey, after which the meeting adjourned for Columbia, where luncheon was served and the nine visited. After the visit to Columbia, the delegates were taken to the Sunnyside mine of the Utah Fuel Company, where an hour was spent. Upon motion, unanimously adopted, the Secretary was instructed to express the sincere appreciation of the delegates to the Price Chamber of Commerce, the Civic Association, the Price News-Advocate, Harry Keene, Harry Duberstein, John M. Jennings, Chairman of the Committee of Arrangements, and Geo. A. Murphy of the Financial Committee, for the complete success in handling the meeting of the Institute. The report of the Membership Committee on the following was accepted: R. E. Birchard, Standardville, Utah. Dave Brown, Spring Canyon, Utah. William Burt, Hiawatha, Utah. A. J. Coats, Fort Collins, Colo. R. P. Fitch, Rains, Utah.

Jan 1, 1925

By D. T. Iseley, David Bennett, Christopher Norris

The microtunneling test program is one of three main elements of the research project, "Trenchless Construction: Evaluation of Methods and Materials to Install and Rehabilitate Underground Utilities." The other elements of this project include an evaluation and demonstration of horizontal directional drilling technology and pipeline rehabilitation systems. The research is a Corps and Industry cost-shared project funded under the Construction Productivity Advancement Research (CPAR) program. The laboratory partner is the U.S. Army Engineer Waterways Experiment Station (WES), Geotechnical Laboratory. The industry partner is the Louisiana Tech University's Trenchless Technology Center. Industry participants are contributing over half the total cost of this research. The overall objective of the CPAR program is to improve productivity in the U.S. construction industry, thereby enhancing the competitiveness of the U.S. industry in domestic and overseas markets. The Corps and other federal and state agencies will benefit by realizing reduced costs, enhanced safety and better products and services as the products of this research program are developed, tested, and pushed into the market place. The microtunneling demonstrations and evaluations were performed during September and October 1992 at a specially constructed test facility built at WES in Vicksburg, MS. The test facility is 104 m (340-ft) long, 5 m (16-ft) wide, and 4 m (13-ft) deep. It was constructed, ironically, by first excavating a trench into which 6 different types of soils were placed and compacted in approximately 18.3 m (60-ft) long sections. The soil profiles include a lean silty clay (loess), a dry sand, highly plastic buckshot clay, wet sand, clay gravel and silt. The interfaces between soil profiles were sloped to simulate mixed face conditions as the machines exit one zone and enter another, and provided challenges to ground control and alignment and grade control during these transitions. Moisture and density tests were performed on all materials in each lift to assure uniformity. Horizontal inclinometers and 3 levels of settlement plates were installed at key locations in the test bed. The settlement plates and rods and the horizontal inclinometers, together with surface survey points, provided data to evaluate settlement and heave along the test section. Penetration resistance was measured in each section of the completed test bed. Strain gages were installed in 8 of the 0.6 m (24 in.) diam. Hobas pipe sections to measure strains in the pipe as it was being jacked into place. Steering jack loads, jacking thrust loads, and cutterhead torque were also measured during the drives. The results of these tests will be used together with other information to develop guidelines for conducting feasibility studies and selecting proper methods compatible with anticipated conditions.

Jan 1, 1993

By C. Fiore

The annual general meeting of the branch was held in the Witbank Golf Club on Thursday, 28th August, 1969 at 5:00 p.m.

Nov 1, 1969

Established as a Division September 27, 1949 LeRoy Scharon, Chairman C. M. Cooley, Secretary 29 West 39th Street New York 18, N Y

Jan 1, 1952

By Mark Lawshe, Shosie Serata

The conventional method of mining underground, using narrow rooms and wide pillars, has proven to be poorly adapted to the powerful earth forces that act on underground openings. The high costs of roof bolting, timbering, steel arching, and other means of artificial roof support for both room-and--pillar operations and longwall gate roadways attest to the limitations of traditional mine design. Too often, productivity halts when corners buckle under "cutter roof' stresses, floors heave, and roof strata fall. In the 1960s, an entirely different approach to strata control was developed. The Stress Control Method (SCM) stabilizes failing roofs and heaving floors by removing the cause of failure - highly concentrated stresses around the openings. This is achieved by regrouping three to 10 rooms into a single entry where the rooms are separated by very narrow yielding but stable pillars. SCM employs a specified excavation sequence to redirect earth pressures away from the mine openings. The yield pillars act as scaffolding until a permanently protective secondary stress envelope forms around the entry (lower part of Fig. 1). The result is a dramatic improvement in roof stability and a significant reduction in artificial support costs. Some roof bolting must still be done, mainly to meet Mine Safety and Health Administration (MSHA) requirements. Potash, salt, coal mining The stress control method was first applied to the deep potash mines of Saskatchewan. The new SCM design actually increased room widths to enlarge the stress envelope, redistributing the stress concentration away from the entries (Serata and Schultz, 1972). This resulted in a turnaround from pervasive roof fall/floor heave conditions that threatened operations in the eight new mines opened in the Saskatchewan potash field during the 1960s. In addition to eliminating all roof bolting, SCM successfully stabilized ground conditions and substantially increased productivity over the original design (Serata, 1976, 1982). Similarly, the deep Boulby potash mine of Yorkshire, England, was about to close after various conventional means of combating roof failure had failed. An SCM test area was excavated that resulted in the ground quickly stabilizing without need o artificial support (Serata, 1982; Chilton and Laird, 1982). Another example is the Diamond Crystal salt mine in Louisiana. Here, multiple level mining was being extended from the third to the fourth level downward in the Jefferson Island salt dome. Extensive roof failure was occurring at 396 m (1300 ft) ii the third level. SCM design principles in creased the room size from 24 to 49 m (80 t( 160 ft) on the fourth level at a depth of 457 m (1500 ft). This unprecedented expansion of room width was first questioned and later ap¬proved by MS HA, which expressed satisfaction in seeing the improvement documented by field validation data in 1973. The most recent SUM application to salt mining has been in the Michigan Basin a: Goderich, Ontario. The anticipated abandonment of Domtar Chemicals Group/Sifto Salt Division's deep operation in bedded sale was prevented by a multi-year SCM implementation. The success of SCM in dry mining[lecc] to its application to solution mining of salt This adaptation of the method is now being implemented in bedded salt at Fort Saskatchewan, Alberta by Dow Chemical Co., and in St. Clair, MI by International Salt Co. Application of SCM to solution mining in a salt dome is proceeding at Napoleonville. LA by Dow Chemical. The same technology has also been applied to cavern design for the Louisiana Offshore Oil Port (LOOP), and the compressed air energy storage (CAES)

Jan 1, 1989

By M. Hakala, P. Tolppanen

A method to predict stress-induced failures around excavations based on the mapped geological and rock mechanical information is studied. Totally of 7 km of newly excavated tunnels are mapped in the Inmet’s Pyhäsalmi mine in Finland at depth between 1000 - 1440 m from which the results and conclusions from the firstly mapped 1.7 km of tunnels are reported. The main rock type at the site is very hard volcanite having uniaxial compressive strength of 230 MPa in median and GSI-value 72 - 80. The maximum in situ stress around the tunnels is 65 - 85 MPa. The prediction of observed stress failures was preceded by comparing elastic 3-D simulation results of tunnels with three statistical yield criteria. Read’s or Hoek & Brown’s yielding criteria are found the most relevant, though, the effect of anisotropy was found to be significant. The most important parameters having a correlation with stress failures are the tunnel orientation and the foliation of rock. Also, the rock type and its strength, tunnel shape and the geological anomalies have their effects. Further, a delay time on mapping was found to clearly diminishing of the most violent failures. Une méthode de prédiction des contraintes à la rupture autour d'excavation, établie sur la base de cartographie géologique et d'information sur la mécanique des roches, à été étudiée. Au total, 7 km d'un tunnel nouvellement percé à une profondeur de 1000 à 1400 m ont été cartographiés dans la mine de Inmet’s Pyhäsalmi en Finlande. La roche principale du site est une vulcanite dur ayant une résistance à la compression uniaxiale de 230 MPa (valeur moyenne) et une valeur GSI de 72 - 80. La contrainte in situ maximale autour du tunnel se situe entre 65 et 85 MPa. La prédiction des contraintes à la rupture observées a été évaluée en comparant les résultats d'une simulation 3D d'un tunnel avec trois critères de limite d'élasticité statistiques. Le critère de limite d'élasticité de Read ou de Hoek & Brown’s se trouvent être les plus important, même si les effets de l'anisotropie sont significatifs. Les paramètres les plus importants ayant une corrélation avec la rupture sont l'orientation du tunnel et la foliation de la roche. De plus, le type de roche et sa résistance à la compression, la forme du tunnel et les anomalies géologiques ont aussi leur importance. Finalement, un délai dans la cartographie résulte en une diminution nette des ruptures violentes. Basierend auf kartierten geologischen und felsmechanischen Informationen, wird eine Methode untersucht, um spannngsbedingtes Versagen um Ausbrüche herum vorherzusagen. Insgesamt werden 7 km neu ausgebrochenen Tunnels in einer Tiefe zwischen 1000 und 1400 m in der Inmet Pyhäsalmin Mine in Finnland kartiert. Das hauptsächlich vorkommende Gestein ist sehr hart und vulkanischen Ursprungs mit einer mittleren einachsialen Druckfestigkeit von 230 MPa (Medianwert) und einem GSI-Wert von 72-80. Die maximale In-situ-Spannung im Tunnel beträgt 65-85 MPa. Die Vorhersage von beobachteten Spannungsversagen werden weitergeführt indem man elastische 3-D-Simulationsergebnisse von Tunneln mit drei statistischen Fliesskriterien vergleicht. Read's oder Hoek's Fliesskriteria werden dabei als die relevantesten Kriterien angesehen, obwohl auch der Effekt der Anisotropie als wesentlich erachtet wird. Die wichtigsten Parameter für eine Korrelation von Spannungsversagen sind die Orientierung des Tunnels und die Schieferung der Gesteinsschichten. Desweiteren haben Gesteinsart, Tunnelprofil und geologische Anomalien ihren Einfluss. Ausserdem zeigt sich, dass eine Verzögerung bei der Kartierung die ausgeprägtesten Versagen deutlich verringern.

Jan 1, 2003

By F. W. Maclennan

The following is a list of members who died in 1929. It is compiled from reports to the Secretary's office. Biographical sketches published in Mining and Metallurgy are indicated in the last two columns. Issue Containing Year of Date of Biography Election Name Death 1929 Page 1922 Adams, Waldo Peck................................ June 2 September 435 1903 Argall, Joseph.................................... Nov. 1 1894 Austin, Leonard S................................. Oct. 30 December 580 1905 Bailed, Melbourne.........................Dec. 12 May* 280 1890 Barringer, Daniel M.............................. Nov. 30 February* 119 1895 Batchelleb, HENRr R.............................. Jan. 15 March 167 BINGHAM Charles E............................. 1923 Bibkett, E. H..................................... Apr. 23 August 394 1913 Bibbell, Robert W................................. Jan. 18 April 214 1926 Blackmar, Don C.................................. June 4 March* 189 1920 Bosworth, T. O.................................... Jan. 18 April 214 1894 Boucher, Arthur S..........................Jan. 6 May 257 1927 Brandenthaler, R. R.............................. Dec. 14 January* 64 1889 Bretherton, S. E............................... Oct. 4 December 581 1911 Brunton, F. K................ .. Aug. 17 1924 Brush, Charles F................................. June 15 September 436 1925 Cameron, Frank N........................May 16 July 349 1891 Chambers, Robert E.............................. May July 348 1908 Clark, J. Murray................................. Dec. 3 August* 407 1883 Clymer, edward T............................. Oct. 13 December 579 1914 Cofpey, George T................................. May 8 July 349 1901 Cooke, L. H....................................... Aug. 23 November 537 1915 Cousins, George W........................ June 30 February* 118 1886 Cowles, A. H...................................... Aug. 13 September 436 1892 Crane, Theron I............................. Nov. 2 December 579 1921 Crowell, Samuel B................................ Dec. 9 January* 64 1907 Curtis, Bracey.................................... Jan. 31 April 215 1895 de Batz, Rene........ ............... January March* 189 1928 Dines, Tyson S.................................... March May 257 1909 Dufourco, R. G............................. January November 537 1920 Ericsson, Eric A.................................. Apr. 4 April* 231 1920 Erlenborn, Willi....................... Nov. 10 March* 189 1888 Farish, John B.................................... Nov. 14 December 580 1891 Ferguson, Vincent........................... Mar. 8 May 257 1917 Finucane, T. R.................................... Mar. 22 February* 118 1921 Foch, Ferdinand................................... Mar. 20 1885 Foucar, Edouard I....................... Jan. 3 March 167 1880 Francklyn, Charles G............................. Jan. 11 April 214 1925 Friedl, Arthur J.............................. July 9 September 435 1894 Fuller, James W. Jr............................... April 4 1876 . Goodale, Charles W.............................. April 11 May 258 1912 Graybill, John..........................Dec. 11 June* 320 1887 Gritzner, F. A.................................. Mar. 26 July 394 1906 Gunther, C. Godfrey.............................. Deo. 26 February* 118 1900 Haggen, Edward A.............................. Apr. 22 October 488 1895 Hall, Benjamin M........................Nov. 19 March* 189 1913 Harper, Walter S................................. Feb. 15 August 394 1921 Hawley, William S................................ Apr. 20 June 298 1920 Herd, Walter ... . . . ..... Aug. 23 September 435 1889 Holt, Marmadtjke B............................... Dec. 26 February* 118 1928 Husk, Albert D .... ................June 9 December 580 1920 Hutchinson, S. Pemberton......................... Feb. 16 April 214 1926 Ivanoff, Constantine L............................ Mar. 7 June 298 1914 Jaeger, Frederick..... .............July 28 October 488 1891 Jennings, Robert E.......................Oct. 30 December 580 1922 Johnson, Homar L................................. Sept. 17 October 483 1927 Johnson, Homer H................................. Mar. 21 May 257 1918 Jones, Jesse L ..... ...............Jan. 22 April 214 1892 Jones, Thomas J................................... June 18 October 488 1883 JoOet, Cavalier H................................. Mar. 29 May 258 1881 Kennedy, John S . . . . ... ..........Apr. 22 August 394 1881 Kennedy,- Hugh.................................... May 23 September 435 1895 Kebvin, James H................................... September October 488 1926 Kkom, L. J........................................ Dec. 24 January* 64 1905 La Follette, Harvey M........................... Sept. 19 November 538 1920 MacDonald, Alexander............................ May August 394 1904 Marengo, Paolo................................... May 257 1925 Maxwell, Allison R............................... May 7 November 538

Jan 1, 1930

By Melvin E. Poad

INTRODUCTION The placement of concrete by means of pumping is one of the fastest growing practices in the construction industry and is widely used in mining applications. In¬troduced in the United States in 1933, this method has been adapted to many aspects of the industry, especially where space for equipment is either limited or inaccessi¬ble. Concrete pumping offers many advantages over other methods of placement. Using pumping equipment, the concrete mixture can be transported in a single operation from a central mixing site to numerous forms in many remote locations, eliminating the need for rehandling the mixture and reducing the labor costs to a minimum. Pressurized pumping provides an extra mix¬ing and blending step, resulting in concrete of improved uniformity and greater density. In turn, this produces high compressive strength and increased durability. Various placement defects can be reduced or pre¬vented by pumping. The steady and even flow of the concrete minimizes the form distortion and delays caused by vibration problems that are common to fast dumping of the concrete mixes. Segregation, planes of weakness, bleeding, and surface faults are diminished considerably. Perhaps the greatest advantages of concrete pump¬ing are the distances and volumes that can be handled. Depending upon the equipment used, pumping rates of from 8 to 115 m'j/h (10 to 150 cu yd per hr) can be achieved, with a horizontal range of up to 460 m (1500 ft) and a vertical range of up to 140 m (450 ft). Theoretically, almost any concrete mixture could be pumped. However, consistency is an important consid¬eration for efficient operation. Mixes that are too wet, too dry, too hard, or poorly graded can cause difficulties that increase the costs. Similarly, mixes containing aggregate material that is over the maximum size or that is porous and highly absorptive can also result in diffi¬culties and increased costs. For the best efficiency, the mix formulation must have a well-graded sand and aggregate combination, with enough fines to provide flowability. If additional water is required for pumping, additional cement must be used to meet the design criteria of the structure. Each pump manufacturer can provide recommended mix formulations for the specific equipment furnished. Since the 1950s, many improvements have been made in pumping technology. Originally, only rigid steel pipe of 168 to 245 mm (6.625 to 9.625 in.) diam was used; today, both rigid steel pipe and flexible conduit are available in diameters of from 76 to 203 mm (3.0 to 8.0 in.). Improved pumps have been developed, and the introduction of power-coordinated booms has in¬creased the versatility of concrete pumping. The older and more conventional placement methods can no longer compete with the flexibility, efficiency, and econ¬omy of the pumping method, and continued research and development can only improve the system and expand its use throughout the mining industry. PUMPING EQUIPMENT The three major types of concrete pumping units currently available are the piston pumps, the pneumatic pumps, and the squeeze-type pumps. Table I provides a comparison of capacities, horizontal ranges, and vertical ranges for pumps of all three types. Piston Pumps The simplest of all concrete pumps are the piston¬type pumps; they are available from many manufac¬turers. As shown in Fig. 1, each unit consists of a

Jan 1, 1982

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By F. W. Smith

Within the Northern Pennine Orefield, England, intersections of fluorite veins are generally richest in ore, and their recognition enables investigation of the ore potential of the intersecting bodies. A method, based on routine analysis of fluorite for yttrium, has been used with success in mines of the district to predict the location of intersections in partially developed fluorite-bearing ore bodies. This short note is intended as a brief introduction; the details of United Kingdom applications are to be published at a later date. Geology The Northern Pennine Orefield is the second largest fluorspar-producing region in Britain. The associated minerals galena, sphalerite, chalcopyrite, barite, and witherite have all been extensively worked in the past but, with the exception of a single barytes mine, all of these operations are now abandoned. Galena alone is recovered as a byproduct of fluorspar beneficiation. At the present time two companies are producing about 38,000 tpy of fluorspar from high grade vein ore bodies, and a number of concerns are actively prospecting in old lead workings. The mineralization is localized in veins and replacement flats within favorable lithologies in an alternating limestone, sandstone, shale sequence. Since the thickness of individual beds is not great (generally less than 20 m), vein bodies tend to be ribbon-shaped, with length greatly exceeding width and height (Dunham). They thus present difficult exploration targets for conventional diamond drilling techniques. Ore grade varies greatly along the veins, which can change rapidly in character from mineral stained fault gouge to massive mineral bodies up to 10 m in width. It has been known for a very long time that the richer ore bodies tend to be associated with vein intersections in favorable beds. It is assumed that vein fissure intersections provided vertical channelways through which hot ore-bearing brines were able to ascend from deeper cavities in the sequence. Thus investigation of intersections at other horizons generally yields further ore bodies. Technique The relative concentration of yttrium in fluorite is a good indicator, in the North Pennines, of the direction of flow of mineralizing fluids. In all of the ore bodies so far studied in detail in this region, the concentration of yttrium increases markedly in the direction of vein intersections that are believed to have acted as feeder zones to the remainder of the bodies. The reason for this relationship is probably connected with the degradation of soluble rare earths-fluoride complexes by the changes in various physicochemical parameters of the hydrothermal fluids as they entered the ore body cavities from the feeder zones. The simultaneous release of free fluoride ions also contributed to the increased ore values in these intersection zones. Yttrium values in fluorite taken from a longitudinal traverse of the western ore body, Redburn mine, Rookhope, are plotted in [Fig. 1.] The position of a major vein branch and three cross veins are clearly indicated by yttrium maxima. The cross veins themselves contain up to 2.5 m of low grade fluorspar ore. The exact positions of B and C were unknown previous to the geochemical survey. Similar results have been obtained for several other veins and ore bodies in the area. Clean fluorspar was collected as bulk samples from within 300 mm of the vein wall and, since the vein infilling is simple, with mineral growth from the walls inward, analyses of the samples are considered comparable along the length of the ore body. A number of samples taken from the center of the vein show a

Jan 1, 1975

By T. J. Carew

INTRODUCTION The application of micro-computers to the evaluation of an advanced exploration project is illustrated by reference to a large project in the Sudbury Basin region of Canada. The deposit in question is polymetalic (Cu, Pb, Zn, Au, Ag) with a strike length of 3048m (10000ft) and has been explored to a depth of 457m (1500ft) by approximately 1350 diamond drillhole of between 9 - 305m (30-1000ft) in length. The large amount of data involved and the ability to rapidly recalculate geological reserves in response to fluctuating metal prices were key factors in the application of a micro-computer based evaluation system. The evaluation software was implemented at two levels, with data capture implemented mainly on - an entry level workstation comprising an IBM-PC compatible (PC-XT) micro-computer equipped with a 20Mb hard disc, printer and RS-232 serial communication facilities. Data analysis, display, geo- logical modelling and processing was carried out on a high-level workstation based on a Hewlett-Packard 9000 super- micro computer and peripherals, including a 40Mb hard disc, tablet digitizer and plotter. Basic drillhole data was entered either directly at the super- micro or transferred from the PC via the RS-232 link. The evaluation process proceeds once the data is edited and in a "clean" format within a master project database. The software is comprised of a suite of modular, integrated programs based on ease of use, interactive graphics and optimized for the computer workstation. Drillhole data was displayed and plotted as hard copy plots on which geological interpretations were developed. The data was copied into small manageable projects consisting of 100 to 150 drillholes, within which a sectional model was created such that all the models are interconnected along the entire strike with no overlaps. Geological reserves were derived by developing a set of function variables, including Net Smelter Return, within the database. The values of these functions can be displayed on the model sections and "ore” zones defined in accordance with pre-determined cutoff values. Interpolation of the modelled data was then carried out and geological reserves calculated and reported. DATA CAPTURE Drillhole data for the project consisted mainly of existing drillhole logs, although the system allows for the entry of new information as it becomes available. The structure and format (i.e. number and length of variable fields) is user defined. The log for this project comprises an abbreviated section for name, competency, grain size, texture, colour and alteration. Fields are defined for structural information regarding bedding and fault angles and up to seven mineral identification columns and percentages are available. Assay information consisting of the sample number, width and six determinations (Cu%, Pb%, Zn%, Fe%, Au opt., Ag opt.) along with a comments section of 25 characters completes the log. One powerful feature of the data- base is the ability to set up function variables that are mathematical and/or logical combinations of the project variables, enabling the user to develop economic variables such as Net Smelter Return. These functions include current treatment costs, deductions and penalties and can be rapidly recalculated within the database to accommodate changes in metal prices, forecast recoveries and exchange rates (Table 1).

Jan 1, 1987

By N. F. Fiore, M. C. Mangalick

DISLOCATION damping in single crystals is orientation dependent.&apos; If dislocations are distributed uniformly over all the slip systems in a crystal, and if each dislocation contributes identically to the damping process, then a damping loss or decrement 6 arises. In fact, however, for a crystal of axial direction (hkl), dislocations on certain slip systems are oriented to make a larger contribution to the damping loss than are dislocations on other slip systems. Thus the measuring damping, 6 (hkl) is proportional to an orientation factor (hkl), which weights the ability of each slip system in the crystal to contribute to the total damping. If it is assumed that dislocations are distributed uniformly on all slip systems, then for a given orientation of an fcc crystal this factor is determined2,3 by the three nonzero elastic constants Cll, C12, and C44 and the density p. In this communication measured values of elastic constants for two Fe-18Cr-Ni alloys are reported, and the dislocation damping results on two different orientations of each alloy are analyzed for orientation dependence. Elastic constants were determined by measuring the ultrasonic velocities of one longitudinal and two shear waves in (110) direction of the fcc crystals.4 Variation of velocities with the change in orientation is very small for (110) orientation,3 and so this direction is ideal for orientation factor and comparative damping measurements. Experimental Procedures. The nominal compositions of the alloys studied are given in Table I. These alloys were vacuum-melted and vacuum-cast from high-purity materials at U.S. Steel Applied Research Laboratories. The single crystals were grown by the modified Bridgeman technique. The specimens were homogenized at 1350°C for 20 hr, cut to shape and size with a spark cutter, and then annealed at 1100°C for 20 hr prior to use. In the KHz frequency range, the Marx composite piezoelectric oscillator was used to obtain the damping. The ultrasonic pulse technique was used to measure the ultrasonic velocities and the attenuation in the MHz frequency range. M.c.mANGALICK,-- StudentMemberAIME,formerly Research Assistant, University of Notre Dame, is now with Casting Laboratory, Chase Brass and Copper Co., Cleveland, Ohio. N. F. FIORE, Associate Member AIME, is Assistant Professor, Department of Metallurgical Engineering and Materials Science. University of Notre Dame, Notre Dame, Ind. This communication is part of a Ph.D. thesis presented by M. C. MANGALICK to the Graduate School of the University of Notre Dame. Manuscript submitted May 20. 1968. IMD Elastic Constants. Table I gives the values of the ultrasonic velocities. Values of C11, C12, and C44 calculated from these velocities are also given in Table I, and it is evident that there is very little variation in C11, C12, and C44 with nickel content. This slight variation may be attributed to inaccuracies in velocity measurements and practical differences in the orientations of the specimens. For example, in B the time interval for vl was 2.4 ± 0.05 sec giving an error in v1 of 0.235 x or 0.005 in. per sec. The elastic constants obtained by pulse technique were tested by calculating E<110>, Young&apos;s modulus for the (110) direction, from the equation: This value was compared with the value computed from the resonance bar measurements in which where fr is the resonance frequency and I, is the length of the specimen. In all cases, the values calculated from the ultrasonic velocity measurements

Jan 1, 1969

By D. M. Range

Exxon is reclaiming an 80-hm2 (200¬acre) Wyoming uranium tailings basin in accordance with US Nuclear Regula¬tory Commission (NRC) requirements. The reclamation plan is designed to control the tailings and their radioactive emissions for at least 1000 years. Rigorous modeling and analyses of potential erosion by wind and water were used in planning. The use of riprap was minimized and the need for long¬term reclamation maintenance was eliminated. The resulting design is simi¬lar to regions stable from erosion cre¬ated by nature. The concepts used in the design and implementation have wider reclamation application. Highland history Exxon operated the Highland uranium operations from 1972 until mid-1984. Open pit and underground operations mined the roll front, sandstone ore body. The 2.5 kt/d (3000 stpd) mill employed conventional crushing, grinding, sulfu¬ric acid leach, CCD, SX, precipitation, and drying circuits. The mill produced 10 Mt (11 million st) of tailings. An¬other operator began commercial scale in situ leach operations in 1988. The tailings basin was constructed by deposition of mill tailings in a natural ravine dammed with a 30-m (100-ft) high by 1050-m (3400-ft) long clay core earthen dam built entirely downstream of tailings. Figure 1 shows the layout of the tail¬ings basin in 1984. Tailings were de¬posited along the dam face on the east side of the basin as well as in the west end and along the south edge until the early 1980s. This left a deep water pond in the middle of the basin. The deep pond was largely filled with tailings during the last four years of mill opera¬tions by the construction of two tailings causeways and the operation of a float¬ing tailings line. The pond water volume was mini¬mized by recycling tailings water to the mill and by operating a 32-L/s (500¬gpm) evaporation enhancement spray system during the last few years of mill operations. Regulatory requirements Several conceptual tailings basin rec¬lamation plans were prepared during the mine life. In September 1984, Exxon began preparation of a detailed plan in response to an NRC request. Water, Waste and Land Inc. (WWL) of Fort Collins, CO was selected to help pre¬pare the plan that was submitted to the NRC in December 1984. The design had to meet the require¬ments found in NRC regulations (NRC, 1985). These were based on regulations prepared by the Environmental Protec¬tion Agency (EPA, 1983) at the man¬date of federal law (UMTRA, 1978). The NRC regulations contain strict primary standards and a somewhat more flexible set of secondary standards. The primary standards include requiring that: • radiation from tailings be reduced to background levels; • radon gas emitted by the tailings be reduced to 20 pCi/m2/s from the surface of the reclaimed tailings pile; • tailings not be dispersed into the environment; and • these conditions endure without routine maintenance for at least 200 years and for 1000 years to the extent reasonably attainable. The lengthy time requirements pre¬sented a large engineering challenge. Radon is a radioactive, chemically inert gas with a 3.8 day half-life. It is released from uranium tailings by the radioactive decay of radium, which, in turn, is a radioactive decay product of uranium. Radon emissions from ura¬nium tailings can be controlled with an earthen cover thick enough that the radon gas decays to its solid, short-lived radio¬active daughters before the radon dif¬fuses through the cover material into the air.

Jan 1, 1989

By B. Ramaswami, Y. Nakada, R. Schaefer

EARLY investigations1 of the tensile properties of fcc single crystals did not reveal the easy glide region of crystals with axial orientations inside the standard stereographic triangle. With the production of high-purity metal, a region of easy glide became known1 Consequently, modern work of fcc single crystals ordinarily uses high-purity metal. In order to determine whether further purification of aluminum beyond 99.992 pct purity would result in significant alteration of the tensile properties of single-crystal specimens, metal was zone-refined and crystals grown from the purified material were compared to crystals grown from the original material. Some similar studies were made of gold and silver crystals. Aluminum. The material to be zone-refined was cast into bars 1/4 by 1/2 by 8 in. A zone, approximately 1-1/2 in. in length, was passed over each bar eighteen times at a rate of 6 in. per hr. Melting and stirring were carried out by an induction heater under a vacuum of approximately 10"5 mm Hg, the metal being held in a graphite container. he last 3 in. of each processed bar were then cut off and discarded. No quantitative determination of purity was made, but the effectiveness of the zone refining was tested by a method first suggested by Walton, Tiller, Rutter, and winegard3 depending on the relation between cell structure and impurity content. Rutter and Chalmers4 have shown that a cellular substructure is associated with the presence of impurities during solidificatibn, and have shown how solute rejection at the freezing interface can lead to such an association. Thus the observation of cells gives evidence of impurities being present. To look for cells, the last quarter of a zone-refined bar was cut off, remelted, and grown rapidly into a single crystal. It was then heavily electropolished and thermally etched.~ Cells were revealed by the presence of many parallel rows of etch pits. The same process was carried out with material from the first half of the bar, but no cells were detected. These observations give evidence that a significant amount of impurity material was moved to the "dirty" end of the zone-refined bars. Further indication of purification was obtained from recrystallization experiments. It was found that, after cold rolling to a 97 pct reduction in thickness, the material from the pure end of the bar re-crystallized at room temperature, but the material from the "dirty" end did not. Other investigators6 have previously reported that zone-refined aluminum recrystallized at room temperature. The purified material was cast into 1/4 by 1/4 in. bars, and seeded crystals were grown in a (100) orientation and in a single slip orientation where the Schmid factor is m =2. The crystals were then cut to the desired length, annealed for 24 hr at 610°C, furnace-cooled, and etched in Tucker&apos;s etch. The crystals were tested in a floor-model Instron machine, using a gage length of 4 in. and a strain rate of 1/2 pct per min. The resulting curves are shown in Fig. 1. It is seen that the differences between the curves for zone-refined and original-purity materials are small. These differences are within the scatter observed for crystals of the original purity. In particular, the zone refining does not significantly affect the onset of the different "stages" of the curves.

Jan 1, 1964

By T. S. Golosinski

Introduction Typical dragline rigging includes several sets of ropes: hoist ropes, drag ropes, dump ropes and boom suspension ropes. The drag and hoist ropes contribute most to the dragline operating cost. Both work in very difficult conditions and wear out rapidly. As a result, frequent rope changes are required resulting in substantial cost and productive time loss. Although performance of rigging has a profound impact on dragline efficiency, there are no common, industry accepted standards related to its selection, operation and maintenance. Little information on the subject is available and the practice differs widely throughout the industry. The performance of drag and hoist ropes depends on: •rope construction that should closely match requirements of the specific application, rope operating conditions and its maintenance and • correct definition and application of the rope discard criteria so that full advantage of the rope can be taken without endangering safety of its operation. This paper reviews the present practice related to all these factors. Its findings are based on the wide survey of the present industrial practice. Field survey In total, 33 North American dragline applications were surveyed of which 17 were Bucyrus-Erie machines, 10 were Marion and six were Page (now P&H). Most of the draglines were stripping overburden in coal mines, but the survey also included some unique applications such as phosphate mining in Florida and oil sand mining in Alberta. Hoist drive power of the surveyed machines varied from 300 to 7775 kW (400 to 10,400 hp). Drag drives varied from 300 to 5815 kW (400 to 7800 hp). Boom lengths varied from 37 to 110 m (120 to 360 ft). Bucket sizes varied from 5 to 96 m3 (6 to 115 cu yd). The hoist and drag ropes installed on the surveyed machines were supplied by 10 rope manufacturers and differed widely in type and construction. Most common rope strand constructions included standard strands, Seale, Warrington, Warrington-Seale and Filler-Seale. Results of the survey [ ] Working conditions Working conditions of hoist and drag ropes are quite different. The first cope primarily with bending on rope drums, point sheaves and deflection sheaves, and with tension force imposed by the bucket and its load. The drag ropes are exposed to severe abrasion at the pit crest, and to severe bending and crushing at fairleads, all while coping with tension forces imposed by resistance of the bucket to dragging. Bending of drag ropes at house rollers and at rope drums is of minor importance for their performance. Differences in working conditions result in different wear patterns, which make it feasible to install discarded hoist ropes to further serve as drag ropes. The hoist ropes wear primarily due to fatigue breaks of wires, while the drag ropes are subject to severe abrasion of outside wires, rope deformation and internal wire-strand breaking. The last is often indicated by the reduction of rope diameter. The summary of working conditions for the surveyed draglines is shown in Table 1. For simplicity, Table 1 does not include the data on rope bending at the fairleads, nor at the deflection rollers and house rollers. Similarly, Table 1 does not include the data on drag rope bending (other than that on the rope drum), as its influence on rope life is insignificant in view of severe abrasion and internal wear taking place elsewhere. The following abbreviations are used in Table 1: [D - diameter of rope drum. DP - diameter of boom point sheaves. d - rope diameter x - mean value of the measured parameter. s - standard deviation of x B/d - average bending ratio of ropes B/d =(D+ DP)/2d n - operational safety factor of the rope (rope stength) x (number of ropes) n(bucket load + bucket weight)] In terms of D/d ratio, the working conditions of hoist and drag ropes are similar to that in other related industrial applications. However, the safety factors of the dragline ropes are lower, which requires them to carry relatively large tension forces. The last explains the fact that the life of dragline ropes is low when compared to other industrial rope

Jan 1, 1995

By N. A. D. Parlee, I. D. Shah

The solubilities of oxygen in liquid Ag-Au, Ag-Pt, and Ag-Pd alloys have been determined in the range of 940° to 1200°C at 1 atm pressure of oxygen using an improved Sieverts technique. The additions of gold (0 to 57 pct), platinum (0 to 27 pct), and palladium (0 to 22 pct) to liquid silver continuously decrease the solubility of oxygen in liquid silver. Activity coefficients and interaction coefficients of oxygen in these alloys and have been calculated. The interaction coefficientscan be described by the following equations: Approximate second-order interaction coefficients (pM have been calculated. The contribution of the product, pMxNM, to In yo only becomes significant at high alloy concentrations. RECENTLY several investigators1"3 have redeter-mined the solubility of oxygen in liquid silver. In general the recent findings1" for Po, = 1 atm are in good agreement with the equation:4 log %O = 695.4/T + 1.086 [1] It is established1, 2 that the solubility of oxygen follows Sieverts&apos; law reasonably well, at least up to 1 atm partial pressure of oxygen. It should be noted that a very slight deviation from Sieverts&apos; law has been reported by Lupis and Elliott.3 Only recently has the solubility of oxygen in Ag-X binary liquid systems received much attention, and only the Ag-Cu, Ag-Au, Ag-Pt, and Ag-Pd systems have been studied. Par lee, Sacris, and zoellners reported that the solubility of oxygen in liquid silver increases progressively with the small additions of copper (0 to 0.5 pct) up to the point where the copper oxide phase appears. Lupis and ~lliott~ have reported that the solubilities of oxygen in liquid silver decrease progressively with the additions of gold, platinum, and palladium. However, their investigation was limited to 0 to 5 pct of these alloying elements, a range too limited to satisfy the requirements of the research program of this laboratory, where the solubilities of oxygen in the range of 0 to 30 pct were required for the calculation of chemical diffusion coefficients of oxygen in these alloys. The present paper deals only with the solubilities of oxygen in these alloys while another paper will deal with the diffusion of oxygen in these alloys. EXPERIMENTAL METHOD A Sieverts technique was employed which involved an improved design of the reaction tube, Fig. 1, as described briefly in a previous communication.4 Liquid melts (70 to 100 g) were held in Norton Alun-dum crucibles. The induction furnace used had a maximum output of 2.5 kw with very fine control. This induction system had excellent stability and the final equilibrium temperatures measured by calibrated Pt-Pt, 10 pct Rh thermocouples are believed to be within ±2C. These thermocouples were checked against the melting point of pure gold wire, and the results agreed with standard tables (±l°C). Each run involved a careful sequence of operations. The sample was heated to and held at about 900°C in vacuum for 3 to 4 hr with the thermocouple in a retracted position. After melting, the melts were usually exposed to a small amount of oxygen (0.5 to 2.0 cu cm) and then degassed slowly for 1 to 2 min to remove any oxidizable volatile impurities such as carbon and sulfur as CO, CO2, SO2, and SO forth. Argon gas was then allowed to enter the system. The protection tube containing the thermocouple was then

Jan 1, 1969

By N. Mungan, F. W. Smith, J. L. Thompson

Adsorption of polymers and transport, rheology and oil recovery efficiency of their solutions were studied in the laboratory to evaluate the use of polymers in waterflood-ing. While a tenfold mobility reduction was obtained with polymer concentrations as low as 0.05 per cent by weight, the mobility reduction depended on the type of polymer, molecular weight, salinity and pH of water, crude oil and capillary properties of the porous media. Choice of a suitable polymer and a workable concentration will have to be tailored for each application. Little reduction in the residual oil saturation can be expected from polymer flooding. Improvement in the volumetric sweep efficiency is possible hut the extent of the improvement can best be evaluated by properly designed field testing. Some aspects of the field use of polymer floods are discussed. INTRODUCTION Waterflooding is a simple, inexpensive secondary recovery method and is being used widely. Innumerable laboratory studies have been made to unravel the fundamentals of the displacement of oil by water and to find the ways of most efficient oil recovery. These studies and a great many field case histories have revealed that the prime cause of poor oil recovery is the inefficient and incomplete sweep of reservoir volume by the injected water. Sweep efficiency is affected by many factors of which the mobility ratio is an important one. Mobility ratio M is defined here as the ratio of water to oil mobilities: M = (k»/y,r)/(k,JJJJ........(1) In Eq. I, the permeabilities are the effective permeabilities and depend on fluid saturations and, hence, change during the different depletion stages in a flood. A wide practice is to use the effective water permeability at residual oil saturation and the effective oil permeability at interstitial water saturation in Eq. 1. If the mobility ratio is greater than one, the mobility ratio is unfavorable and water, being more mobile than oil, would finger through the oil zone resulting in poor oil recovery efficiency. If the mobility ratio is favorable (one or less) the displacement of oil by water occurs more or less in a pistonlike fashion. In some waterfloods. the mobility ratio is unfavorable and any additives by which the mobility of water can be decreased would favor more efficient oil recovery. The thing to bear in mind, however, is whether or not the improvement in oil recovery is sufficient to more than pay out the cost of the additives needed. For example, materials like sugars, alcohols and glycerine reduce water mobility by thickening the water, but the cost of material requirement precludes any field application. For an additive to be useful in water-flooding, it must bring about a large reduction in water mobility at low concentrations; it must be adsorbed only negligibly; and it must not completely plug up the formation. Some synthetic organic polymers have shown promise of meeting these requirements and have been used in the field.&apos;-W owever, no in-depth studies of the rheological, adsorption and oil displacement characteristics of polymer solutions have been reported. The present work is a study of these properties. EXPERIMENTAL In this work, concentrations are given on a weight per volume basis; 0.5 per cent concentration means 0.5 gm of polymer is dissolved in enough water (or NaCl solution) to make 100 ml. A bactericide, usually 0.1 to 0.2 per cent by volume of 38 per cent formaldehyde solution, was used in the polymer solutions. The NaCl solution was 30,000 ppm. Some properties of the polymers studied are given in Table I. Physical properties of all cores used are in Table 2. Flow behavior of polymer solutions was studied by three consecutive flow tests in cores. First, water (or brine) was injected at constant rate of about 1 ft/D to obtain the water mobility. Then, filtered polymer solution (prepared in water or brine) was flowed through the core. Since the rate was constant, increase in the pressure drop across the core reflects decrease in the mobility. Finally, the core was flooded with water (or brine) to study recovery of mobility. The Alundum cores which were used in

Jan 1, 1967

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The reliable quality assessment of grouting operations is one of the key issues in establishing impermeable grout curtains around shafts, drifts and tunnels. Reliable opera¬tional information on the decrease in the permeability of rock during the grouting process facilitates the assessment of the effectiveness of the completed curtain and assures production quality. Quality control and assurance also fa¬cilitates the early detection and elimination of possible de¬fects in the grout curtain. Thorough control of the grouting process provides the capability to evaluate decreases in the permeability of water-producing fracture zones prior to the beginning or the continuation of excavation of the shaft, tunnel or drift. In order to achieve this objective, post¬grouting hydrodynamic investigations are conducted prior to the initiation or re-initiation of work on the underground workings to measure the decrease in the permeability coef¬ficient of aquifers during grouting operations. 8.1 QUALITY ASSURANCE AND CONTROL OF THE ISOLATION OF ROCK BY GROUTING THROUGH HOLES DRILLED FROM THE SURFACE Quality assurance and control are based on measuring the decrease in the permeability coefficient of water-pro¬ducing horizons as a result of injecting grout into the drill¬holes. For this purpose, the value of the acceptable residual permeability coefficient is calculated beforehand based on the acceptable value of ground water inflow into the shaft, drift or tunnel as specified by the "construction norm and integrated practice (SNIP)" (Anon., 1976). Having ob¬tained the values of aquifer thickness, permeability, and head distribution for each aquifer from flowmetric data (dis¬cussed in Chapter 3), the average value of the permeability coefficient of all aquifers intersected by the shaft is calcu¬lated as [ ] where Ki is the permeability coefficient of the "ith" aqui¬fer; Pi is the pressure of the "ith" aquifer; and M; is the thickness of the "ith" aquifer. The total expected inflow of ground water into the shaft is [ ] where RK is the radius of the contour of influence; RCFB is the radius of the shaft; and p is the coefficient of dynamic viscosity of the ground water. The necessary reduction of the coefficient of permeabil¬ity of the aquifer [ ] where Q is the expected inflow of ground water into the shaft from eq. 8.2; Q.-fin is the acceptable inflow of ground water; and a is the safety factor. By knowing the average permeability value of the rock and the permeability reduction coefficient, it is possible to determine the value of the acceptable residual permeability coefficient for all the aquifers [ ] The values of the permeability coefficient, pressure, and thickness for each aquifer are determined from the data gathered during the hydrodynamic investigations conducted in grout holes prior to beginning the grouting operations as discussed in Chapter 3. After injecting the grout into the first of the holes, the injection into each of the following grout holes is preceded by repeated hydrodynamic investi¬gations. Using the data obtained for each water-producing zone, a graph is constructed where the numbers representing the grout holes in their order of grouting are placed on the abscissa, and the values of the aquifer permeability coeffi¬cient are placed on the ordinate (Fig. 103). This curve char¬acterizes the change in permeability of the given hydro¬stratigraphic unit as grouting operations are completed, one hole at a time. The results of grouting operations are considered to be positive when each sequential observation records a steady decrease in the permeability of the aquifer. The permeabil¬ity measured prior to the injection of the grout into the last hole must be lower than or equal to the acceptable perme¬ability as determined by the application of eqs. 8.1 through 8.4. Table 8.1 presents illustrative data on the hydrogeo¬logic properties of aquifers at a mine shaft prior to the initiation of grouting operations. Data also are presented on the reduction in the permeability coefficient during the in¬jection of the grout. The shaft in question intersects eleven aquifers. Based on data concerning aquifer pressure, thick¬ness, permeability, and the expected inflow rate of water for each of the strata, the average permeability coefficient is calculated according to eq. 8.1 as K = 0.2 x 10-12 m2. The

Jan 1, 1993