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By William R. Yernberg

The combined 79th Annual Minnesota Section of SME Meeting and 67th Annual University of Minnesota Mining Symposium was held April 11 and 12, 2006, in Duluth, MN. About 300 mining professionals attended, including engineers, executives, mining managers and suppliers, from throughout the United States and Canada. This year’s theme was “Success in the New World of Mining.” The SME Minnesota Section meeting included a keynote session, a special projects update, a general managers update and technical paper presentations. The University of Minnesota Mining Symposium program featured a panel discussion on Minnesota’s relationship with China. The combined meeting included a trade-show consisting of 64 booth spaces. Tradeshow demand exceeded expectations, and an additional 10 booth spaces were added in the registration lobby. Also included were an SME short course on pump basics and social events.

Jan 1, 2006

By Maurice Grieves

GC. Waterman Mr. Grieves' paper on "Shaft Sinking Today--A Boring Business Tomorrow" in the Dec. 1981 issue of MINING ENGINEERING is an excellent description of recent improvements in speed and costs of shaft sinking. However, shaft boring techniques are not a recent development. In 1935-36, the Idaho Maryland Co. (gold), Grass Valley, CA, bored the vertical 1.5-m-diam Idaho No. 2 ventilation and supply shaft to a depth of 346 m. The equipment was designed and perfected by Branner Newsome and was, I believe, described in "Transactions." Pickands Mather later bored a 1.8-m-diam shaft to a depth of about 305 m using Newsome's design. The Idaho No. 2 shaft cut through hard gabbro, diabase dikes, soft serpentine, very incompetent ankeritized serpentine and a strong fault zone. The completed shaft did not require timber support; the fault zone was cemented. Equipment consisted of a rotating 1.5-m or 3-m core barrel with a slotted bottom which cut a 76-mm kerf. The cutting agent was chilled steel shot introduced into the slot as needed. The driving mechanism and core barrel were lowered into the shaft, the former secured to the wall and the inshaft "engineer" operated the motors which controlled core barrel rpm and advance. After an advance of 1.5-3 m the equipment was lifted out of the shaft, a half(?) stick of powder cut off the drilled core, and it was hoisted to the surface with a cable attached to an eye bolt at the tope of the core. A "shift" consisted of three men: hoistman, equipment operator (down the shaft), and a surface laborer. Advance was variable as equipment and techniques were perfected. Near the end of the job advance was, as I remember, 1.5-3 m per shift. Costs were, as I recall, about $115/m. Up to 4 m cores were lifted out in one piece and swung to the dump by a stiff leg derrick. The Newsome equipment was relatively inexpensive to build and operate and his method should be (more than?) competetive with the methods described by Grieves. W.E. Hawes The author, in mentioning the South African developments of the cactus grab, ignores the parallel development of the Cryderman mucker in this hemisphere. Perhaps a slightly longer history of blind hold drilling would have been in order. Blind shaft boring got a major boost at the Nevada Test Site, as part of the nuclear weapons development, when it was essential to be able to quickly assess depth for weapons testing. One of the early civilian attempts to utilize this technology occurred when Kerr McGee drilled the shaft at the Section 19 Mine at Ambrosia Lake, NM. This twin masted drill rig belonged to a subsidiary of Kerr McGee, not Shaft Drillers. This was a difficult hole, due to bentonitic shales that decomposed. Later advances in drilling fluids eliminated this problem. The Conoco project referred to is not entirely correct, and needs minor amplification: The depth of the development (not pilot) shaft drilled was 684 m, at a diameter of 3.05 m, with a 2.16 m ID hydratatic casing being installed to a depth of 669 m. Conoco Inc., not Challenger, executed the drilling, however Conoco did use Challenger Drilling Company's Rig 14-S. Rig 14-S is unique in that it is specifically designed for shaft drilling, rather than being an oil field rig used for this service. All work was planned and directed by Conoco personnel. Deviation of the aforementioned shaft from vertical was 31 cm. Two smaller shafts (pilot shafts) were drilled with the same rig. These smaller ones were drilled 1.8 m in diameter, to depth of 665 m and cased with .91-m-diam casing, to depth of 650 m to be used as pilot holes for 5.5 m diameter shafts. The total elapsed time for all drilling, casing, cementing, and rig moves was 50 weeks. The constraints mentioned of vertical accuracy, torque, wall collapse, rig set up time, lining time etc., pumping out, etc., may not be as serious as the author implies. First, vertical accuracy: The record achieved by Conoco indicates that with reasonable care, good accuracy can be obtained, primarily by having a fair amount of weight in tension, rather than on the drill bit. Second, the amount of torque that can be applied to the drill stem has been greatly increased by the introduction of larger diameter drill stems, such as the new Hughes rig uses. Third, risk of collapse of shaft walls is minimal with the proper mud program. Basically, the walls are supported by the weight of the mud and it is not removed until some type of casing is installed and grouted in. Here in the real challenge of blind drilling shafts to devise a more economical method of lining bored shafts than using steel casing. The remaining issues were minor at Conoco, provided adequate planning and scheduling takes place. An excellent paper, "Shaft Drilling--Crownpoint Project" by Hassell H. Hunter, presented at the Fifth Uranium Symposium in Albuquerque in Sept. 1981 contains details of this project. A combination of blind boring, with enlargement by either mechanical means (boring machine, or tunneling type shield) or slabbing down, with muck removal through a larger bored shaft would seem to be the future trend in shaft development. The economics, especially for wet conditions, favor this concept over conventional sinking, as has been demonstrated at Kerr McGee --Churchrock, UNC--Churchrock and Conoco at Crownpoint.

Jan 3, 1983

By S. J. Neethling

The flotation froth structure and motion determines the amount of entrained gangue that is collected to the concentrate. Despite this important role in the overall performance, the behaviour of the froth is still ill understood. To date, predominantly empirical models have been developed to describe entrainment. A fundamentally based model is described and used here in an attempt to improve the understanding of entrainment. This general froth simulator (UMIST FrothSim) allows the modelling of a wide range of flotation conditions, as it takes account of a large number of the physical phenomena that occur within flotation froths. The model firstly describes gas motion and bubble coalescence, and the water motion based on gravitational, viscous and capillary effects. The motion of the solids distinguishes between various solids classes based on hydrophobicity, particle size and density and models them by including the effects of Geometric and Plateau border dispersion, particle settling and the motion of the water. The aim of this paper is to show the applicability of this model to explain the observed entrainment and collection of gangue. This is done by comparing model predictions with the experimentally observed relationship between the gangue and water recoveries. The model predicts the identical trends, which are explained in terms of the interaction between the linear effects of water motion and hindered settling and the non-linear effect of particle dispersion.

Jan 1, 2001

By Fred C. Bond

The perennial subject of comminution is deserving of a fresh approach if one can be found. This is attempted here by directing attention primarily to principles rather than to theories. While theories attempt to explain how a certain process acts, principles delineate the laws which the action-must follow. Comminution may be defined as the progressive separation of particulate matter. At the present time this is accomplished commercially by applying energy in the form of compressive and, shearing forces. Such forces strain and distort the particles they contact. When-the applied stress locally exceeds the breaking strength at the weakest flaw in the particle, a crack tip forms. The surrounding strain energy flows to the crack tip, which rapidly extends through the particle as a break. The minute changes in shape and the resulting energy flow cause the failure of other highly stressed flaws, with the consequent formation of a progeny (1) of new smaller particles out of its destroyed progenitor. The applied mechanical energy is transformed into strain energy as the particle is distorted, and then into heat when it fails. Only a tiny fraction of the order of 1/1000 of the applied energy necessary for breakage appears as new surface energy in the product particles. The comminution problem is the relationship between, energy input and product particle size. Its solution is of great-economic importance,

Jan 1, 1961

By M. Baumann

Operating and maintaining the heavy equipment used in open-pit mining operations is an ongoing challenge for mines. Typical operations employ a combination of shovels, wheel loaders, and haul trucks to excavate and transport the material for further processing. As this is the first point of contact for the material after blasting, it is often a bottleneck of all downstream processes, thus, equipment uptime and efficiency is crucial. Automated systems for bucket tooth monitoring, payload measurement, proximity detection, and rock fragmentation analysis will be examined. The impact of these systems on operation safety, efficiency, and reliability, in real-world scenarios will be investigated. The process of managing all the above fleet of equipment in a mine can be streamlined by connecting the equipment monitoring systems to a central server. Using the central server, the distribution of this information can be automated to ensure that the right people see the data in a timely manner, improving both proactive and reactive responses to threats to safety and productivity.

Feb 23, 2014

By G. B. Amicucci

Golden Sunlight Mines Inc. operates an open-pit gold mine and cyanidation plant about 11km (7 miles) northeast of Whitehall, MT. Golden Sunlight is a subsidiary of Placer Dome US Inc. Mining is accomplished with conventional open pit drilling, blasting and hauling methods. The mining rate is about 102 kt/d (113,000 stpd). Gold occurs as fine-grained disseminations in a pipe-shaped breccia and in immediately adjacent country rock. The ore body is low grade, averaging 2 g/t (0.059 oz per st) of gold. The metallurgical plant uses conventional carbon-in-pulp (CIP) technology. Daily throughput averages 6.5 kt (7000 st). Recoveries are enhanced through a sand tailing retreatment facility while a preaeration process reduces operating costs. The low grade ore dictates that maximum mill throughput at the desired grind size be maintained. Therefore, in October 1988, an Armco Autometrics (now Denver Autometrics) particle size monitor (PSM) 400 was commissioned. The PSM 400 is an on-line instrument. It measures particle size and percent solids in a slurry stream. The PSM pulls its own sample, deareates the sample and passes it through diametrically opposed ultrasonic sensors. The unit uses ultrasonic sound to measure the deareated slurry stream's particle size and percent solids.

Jan 1, 1991

By D. K. Denton, M. J. Friedel, D. F. Scott, T. J. Williams

Spokane Research Center personnel have been investigating the use of seismic tomography to monitor the behavior of a rock mass, detect hazardous ground conditions, and assess the mechanical integrity of a rock mass affected by mining. Seismic tomography can be a valuable tool for determining relative stress in deep ( > 1220 m) underground pillars. If high-stress areas are detected, they can be destressed prior to development or avoided. High-stress areas can be monitored with successive seismic surveys to determine if stress decreases to a level where development can be initiated safely. There are several benefits to using seismic tomography to identify high stress in deep underground pillars. The technique is reliable, cost-effective, efficient, and non-invasive. Also, investigators can monitor large rock masses, as well as monitor pillars during the mining cycle. By identifying areas of high stress, engineers will be able to assure that miners are working in a safer environment.

Feb 24, 1997

By E. M. Spokes

A.B. Rushton The work and resulting publications of the Committee on Underground Coal Mine Safety and The President's Commission on Coal were appropriate at the time of publication. The statistics cited in the publications appeared to be valid, and the actions proposed by the committee had merit. Although the actions proposed to achieve further improvements in safety may have been worthy of repeating, to associate statistics (Table 1) that are seven to eight years old with company names is bad. Statistics are readily available from the Mine Safety and Health Administration (MSHA) and are usually within three months of being up to date. A review of the most recent Coal Controlling Co. Report will reveal significant changes in the rankings. The more recent statistics should have been cited, or at least compared with those used in the article. Westmoreland Coal Co. is listed as having a 21.1 disabling injury rate for the period (1/78-12/80). The disabling injury rate for the Company for the year ending 1985 was 7.4 for the underground mines and 6.25 for the total facilities. Through March 1986, the disabling rate for the underground mines was 2.86. The management and employees of Westmoreland Coal Co. are committed to safety. The improvements achieved over the past five years attest to that commitment. We are justifiably proud of the achievement and dislike being misrepresented as in the April 1986 article. Reply by E.M. Spokes I regret that the lapse of time since the original work gives a picture of safety practice that is no longer true today. That Westmoreland has reduced its disabling injury rate by 65% to 7.4 in 1985 is most commendable, and I congratulate you for this accomplishment. It is typical of the progress sought by the National Academy of Science and National Research Council when they funded the original study and published "Toward Safer Underground Coal Mines" in 1982. If a significant number of companies do as well as Westmoreland, the report will have served its purpose. Since the paper was primarily a condensation of the original report, it would not have been appropriate to seek access to later statistics. The purpose in seeking its publication at the time of presentation was to reach those who might have missed the original report. The fact that publication was delayed has not voided that purpose and may have reminded some who were aware of the original that its recommendations still apply to improvement of underground coal mine safety.

Jan 6, 1986

By David A. Brence, Andrew D. Marti, Richard P. Stahura, Ronald J. Valentine

The development of an effective belt-cleaning system for a new conveyor system seemed a fairly conventional assignment. But a closer examination of the project showed a number of “fine print” details: •High altitude location (with resulting cold weather and low humidity conditions). •Long conveyor length. •Troublesome material (very sticky and capable of hardening like concrete). •Limited water. These difficulties are spelled out the challenge of developing a cleaning system for the production conveyor2 (PC2) belt at Climax Molybdenum’s Henderson Mine.

Jan 1, 2005

By J. V. Ghelarducci, M. A. Sharpe, D. G. Chedgy

This article presents a novel design, called the Climax, High Capacity Processor (HCP), for pre-combustion cleaning of coal. This design is particularly suitable for the plant capacities of 400 ~ 660 TPH. The HCP is different from the conventional designs with an equivalent throughput in three major aspects: 1) the use of one large capacity HM cyclone to separate coarse and intermediate raw coal in a single circuit; 2) the elimination of the conventional secondary magnetics recovery system by use of the high efficiency Climax, magnetic separators; 3) the use of "single line" piping with a minimum requirement of ancillary devices. In comparison with conventional designs with similar capacity, the HCP plant costs much less in capital investment, operation and maintenance.

Feb 24, 1997

By Fernando Humberto Muge

One of the problems that arises in the early stages of the mine planning before the determination of the mining exploitation method is to predict what will be the recoverable ore reserves in tonnage, grade and metal quantity, if the select ion is based upon block grades. The estimation of these recuperation or recoverable functions can act as an important planning tool in what concerns the appraisal of the consequences of ore-waste discrimination of individual blocks in a selective mining operation. Estimation methods of all these functions are discussed together with cross validation tests using the available information concerning a preliminary drilling campaign accomplished during the exploration stage of an iron orebody located in the North of Portugal (Moncorvo Orebody).

Jan 1, 1982

By Glen Church, Don W. Deere, Christoph M. Goss

The Standley Lake Dam Rehabilitation Project in Colorado included the complete replacement of the outlet works. This required two 1.8m (6 ft) diameter microtunneled wet lake taps, 380m and 186m (1246 ft and 636 ft) long, a permanent valve shaft 12.2m(40 ft) wide and 32m (105 ft) deep, and a 3.5m (11.5 ft) diameter (962 foot long) outlet tunnel driven with a roadheader. The tunnels were driven in low-strength, expansive, and slickensided claystone. The paper addresses the site investigation, shaft construction, roadheader performance, shotcrete placement, tunnel convergence, long distance microtunneling, and wet lake taps.

Jan 1, 2005

By Chee Theng

"Getting minerals out of ores is messy, but mineral processors learn to have fun with it. Celebrating the innate joys of the industry was the theme of this year’s Colorado MPD Conference. At the same time, attendees dug deeper for insights into problems that continue to dog the community, including tailings dam crises, water resource challenges, and a dwindling number of academics to teach and train the workforce (see box).Tailings dam crisesMatt Fuller of Tierra Group International Ltd. brought some of the audience to tears with his impassioned talk on “The Infamous Legacy of Upstream Tailings Dams,” which won the conference’s Lou Cope award for best paper. The talk was especially poignant coming on the heels of the January tailings dam failure in Vale S.A.’s Córrego do Feijão iron ore mine in Brazil, the latest in a series of catastrophic dam failures over the past decade that includes the Fundão dam failure in 2015.Noting that the two dams share a common characteristic — both are upstream tailing dams — Fuller explained the three basic dam construction methods and why certain upstream tailing dams are susceptible to failure (Fig. 1). The downstream, centerline and upstream methods denote the direction the centerline of the dam crest moves with each successive raise in elevation of the dam. With the downstream dam, the foundation for future raises is essentially the same as that of the starter dam, he explained. With the centerline dam, half of the dam is constructed on the same foundation as the starter dam, and with the upstream dam, each successive raise is built out over the top of the tailings.“Downstream dams are expensive to build because each successive raise is bigger, so the sustaining capex to build them keeps growing,” Fuller said. Many decades ago, mine owners realized they could construct a starter dam, put their tailings through a hydrocyclone, generate a coarse sand fraction and stack the sand over the top of the tailings to build the next raise. “Then, there is no sustaining capex, it all comes out of operations,” Fuller added. The footprint is smaller, and it is a favorable means of retaining the tailings."

Jul 1, 2019

By Robert W. Miller, Rhoda S. Kriesel

INTRODUCTION With each passing year, the general public becomes increasingly aware of the potential hazards associated with many products and services previously considered unharmful. This new consciousness is especially evident in the workplace, both as the result of advances in scientific understanding and the influx of concerned, responsible professionals into the labor movement. For example, data produced by epidemiologists doing retrospective mortality studies which related mortality to toxic substance exposures has often been the prime motivating factor behind reductions in exposure standards. This certainly has been the historical case with many radiation exposure standards, including the limitations set for "Working Level" exposure. The concept of "Working Level" was introduced in PHS Publication No. 494 as a result of both the difficulty in relating radiation units to biological effects and the complexity of interaction of radon daughters with the physical environment. WORKING LEVEL One working level is defined as any combination of radon daughters in one liter of air that will alternately release 1.7 x 105 MEV of alpha energy during decay of 210pb (RaD). Its usefulness is that working level can be readily measured in both the field and the laboratory. Because of this, existing occupational exposure standards for radon daughters in both the U. S. and other countries use working level as their basis. In the U. S., the last twenty years has witnessed a significant reduction in radiation exposure standards from 120 working level months to 4 working level months per year. New international standards are being proposed that would be only slightly higher than these current U. S. standards. POTENTIAL EXPOSURES TO RADON DAUGHTERS The universal use of nuclear power, along with the transfer of nuclear technology to developing nations, has significantly increased the demand for fuels. As a result, uranium mining operations have expanded and increased numbers of miners are regularly exposed to radon daughters. In addition to potential exposures in uranium mines, other deep mines such as tin, gold, platinum and tungsten, also have potential exposures due to the distribution of uranium and its accompanying radium throughout much of the earth's crust. The energy crisis has also prompted interest in general population exposure to radon daughters in buildings. Many homeowners, using additional insulation and caulking to seal cracks and conserve energy, have reduced indoor ventilation to the point where working level exposure could become significant, especially in homes with unventilated crawl spaces. In fact, several serious exposure situations have occurred where mine tailings were used for building materials. MEASUREMENT TECHNIQUES Historically, rapid and convenient measurement techniques have been actively sought to improve the ability of ventilation engineers to limit mining exposures through ventilation control. A number of measurement techniques have been developed, including the Kusnetz method (further elaborated by Shalaynev) and techniques developed by Rolle and Tsivoglou. All, however, suffer from similar handicaps that prevent their usefulness in the mining environment. Their main limitation, in terms of minimizing employee exposure, is the elapsed time from the start of sampling until the results can be calculated. It is usually 40 to 90 minutes. This time delay presents both serious economic limitations as well as the potential for exposing workers to high radon daughter concentrations until calculations are completed and corrections made. Further reducing the utility of the Kusnetz and Tsivoglou techniques is the need for tedious calculations that increase the possibility of human error along with the need for cumbersome sampling equipment that is not ideally suited to the mining environment. INSTANT WORKING LEVEL METER To alleviate these problems, the idea of an Instant Working Level Meter was conceived by several groups. Several such instruments were proposed, built, and tested with disappointing results. Although the concepts behind them were sound, the instruments simply failed to measure working level without excessive distortion. In addition, their cost was relatively high, gamma background presented a problem, and sophisticated computations were necessary to determine working level. In short, these systems lacked the design engineering that would make them suitable for the demands of a mining environment.

Jan 1, 1981

By K. Shoji

Residence time distributions (RTDs) were estimated by water tracing in a number of wet overflow ball mills (diameters 0.38 to 4.65 m) producing dense, coal-water slurries. In open-circuit mills of 0.38 m diameter and various length-diameter (L/D) ratios, the mean residence times of solid were also determined from measured mill holdups. Holdup increased with increased mill feed rate, but the mean residence times of coal and water were still equal to each other. The experimental residence time distributions were fitted to the Mori-Jimbo-Yamazaki semi-infinite, axial mixing model, and the dimensionless mixing coefficient was determined for each of 25 tests in single- and two-compartment mills. This coefficient was found to be independent of the feed rate but linearly proportional to the D/L ratio. The mixing coefficient was smaller for two-compartment mills than for single-compartment mills, showing that there was reduced mixing introduced by the diaphragm separating the compartments. Equations are given to scale residence time distributions for changes in mill diameter and length.

Jan 1, 2008

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module (Fig. 1). The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information. Key for Fig. 1 A-Helium-3 neutron detector B-Wax moderator for neutrons C-Irradiation chamber D-Funnel for receiving ore E-Gate F-Transducers G-Biological shield for neutron source H-Transfer tube for neutron source I-Neutron source J- Microprocessor K-Accumulator for hydraulics L-Control panel M-Electro-hydraulic servo valve to control movement of the source N-Electro-hydraulic proportional valve to control movement of the gate O-Hydraulic power unit Irradiation and Counting Equipment Irradiation of the ore is achieved using a 1 X 10 7 neutron*s-1 californium-252 neutron source. The source is contained in an iron capsule and pneumatically transferred between the irradiation/counting and storage chambers. The irradiation/counting chamber possesses 12, 600-mm (24-in.) long helium-3 thermal neutron detectors embedded in an annulus of wax (Fig. 1). Two sets of six detectors are each interfaced to a preamplifier, amplifier and single-channel pulseheight analyzer system, and the two systems are connected to a scaler. A special timing unit controls the transfer of the source and the on and off switching of the detector assembly, in sequence with the movement of the source. Results and Discussion In the experimental demonstration, using the 107 neutron.s-1 source, a 10% uranium ore was measured to ± 10% (2o total error) in 10 cycles and ±6% (2o) in 100 cycles. Since a larger source was not available, it can be statistically predicted that a 109 neutron-s-1 source could be used to measure a 0.1% ore to ± 10% (20) in 10 cycles. The uranium can occur as discrete nonhomogeneous particles in the ore, and the error caused by such a distribution of particles is ± 4% (2o), as determined by

Jan 1, 1984

By Verya Nasri, Mehdi Bakhshi

The development of the Montreal metro extension Line 5 -Blue Line- toward the northeast are located between the existing Saint-Michel station and Anjou. It consists of 6-km long two-track tunnel, five new stations and six auxiliary structures. Most of the tunnel (5.5 km) will be excavated by a 9.6 m diameter shielded rock TBM and supported by a precast segmental lining system. TBM will be fitted for probing and preexcavation grouting to control water inflow into the tunnel during the excavation. This paper presents major design aspects of the TBM tunnel including alignment, geological condition, spaceproofing, precast segmental lining, and geotechnical instrumentation and monitoring.

Dec 1, 2022

By M. E. Gershon

I would like to congratulate Professor Gershon on an idea that is obviously valuable and is certainly new to me. It has, however, to be further developed, in that one can get fallacious results by limiting oneself to positional weights. In the example problem, the positional weight of one block was 8.6 (the 8.0 value in the paper was incorrect) for 36 units in the cone, for a mean block value of 0.239. The positional weight of the second block was 4.4 for 15 units, for a mean block value of 0.293. Surely one would expect the second block to be rated higher. Additionally, the time value of money adjustment cannot be determined by depth, because the time that a block is mined is determined by the mining sequence, not depth. Figure 1 shows a slightly altered version of Professor Gershon’s pit. The overall value of the pit is the same, as are the total values under each of his blocks. The values in the right-hand cone have been reversed top to bottom, and the values in the rest of the pit adjusted to keep the central cone total unchanged. All values have been multiplied by 10. Figure 2 shows the sequence of mining, at three blocks per year, for Professor Gershon’s pit, and Fig. 3, the sequence for a pit under the other block. Table 1 shows the value and discounted value (157. ) of Professor Gershon’s pit, and Table 2 shows similar values for the second pit. It is apparent that the second pit schedule is more valuable than the first.

Jan 12, 1987

By H. I. Kim

The leaching of a complex matte in chlorine and chloride media was studied. The matte was prepared from manganese nodules by the reduction smelting-sulfidation route. From XRD analysis, the major phases of the matte were identified as CuFeS2, CuS2, (FeNi)9S8, (FeNi)S2, Ni9S8, Ni3S2, (CoFeNi)9S8 and Co metal. Investigated were the chlorine, hydrochloric acid, hydrochloric acid-oxygen, ferric chloride, ferric chloride-oxygen, cupric chloride and cupric chloride-oxygen systems. The results are discussed in terms of reagent concentration, retention time, leach temperature and gas sparging rate. The chlorination process resulted >99% metal extraction efficiency within a short retention time and at atmospheric temperature. However, there were no preferential extractions of Cu, Ni and Co with respect to Fe. The cupric chloride-oxygen system produced leach liquors with the lowest Fe content and with a high base-metal extraction. Sulfide sulfur was converted to mixtures of elemental sulfur and sulfate in all the studied systems.

Jan 1, 2009

By L. You

Theoretical and analytical analyses of three mitigative measures- plane fitting method, trenching, and tension cable led to the development of a total protection plan to eliminate or reduce damages to the residential structures subjected to surface movements caused by underground longwall mining. The protection plan was applied to twelve residential houses and two external garages. The success rate is overwhelming. This paper describes the methods employed for protection and results of those case studies.

Jan 1, 1986