Search Documents

By R. H. King, D. R. Schricker, R. E. Cameron

Mine managers rely on their experience, section foreman's daily reports and occasional time studies to make important production, maintenance and forecasting decisions. This information is often subjective and imprecise, consumes expensive engineers time, is often biased and is restricted to short time periods. Managers need better information. Furthermore, if a recently purchased machine is being evaluated, a large increase (>10%) in productivity may be obvious, but small increases are usually masked by other variables. In today's highly competitive markets, cumulative small changes are important, especially since managers have already exhausted most available, obvious avenues of improvement. A technique to separate these charges from effects of other variables and to quantify the target effect is necessary. CSM researchers are addressing this need by developing an intelligent decision support system (IDSS). It provides key production and maintenance information to support management decisions by recognizing patterns in mining machine sensor data and representing knowledge about mine management, operations, layout and equipment. A continuous miner section at Western Fuels-Utah Inc.'s Deserado Mine in Rangely, CO is the test facility for the research. Power transducers located in the sections load center provide kilowatt versus time data from continuous section equipment. The IDSS pattern recognition algorithms classify the kilowatt versus time patterns into individual machine operating modes. It produces an output similar to a typical industrial engineering time study but continuously, automatically and without an engineer in the section. It is relatively easy for a time-study engineer to identify the stop and end points of a cycle element and classify it by looking at the machine and glancing at a stop watch. The problem is that monitoring system hardware and software do not have these unique capabilities. Consequently, IDSS algorithms must automatically define the begin and end points of machine tasks and identify the task by recognizing patterns in kilowatt versus time data. However, many mines have file drawers of time-study type data that are not very useful. So, IDSS contains the knowledge representation subsystem (IDSS/KRS) to produce short reports containing only key information. This is difficult because mine environments continually change and information requirements differ for each management level. Consequently, IDSS/KRS must learn and adapt to new conditions in a dynamic mining environment. Western Fuels-Utah owns and operates the Deserado Mine, located about 19 km (12 miles) northeast of Rangely, CO (Upadhyay, 1988; King and Eros, 1988). One longwall and two continuous miner sections produce about 900 kt (1 million st) of coal annually for the Deseret Generation and Transmission Cooperative's 4000 MW power plant located near Bonanza, UT. Three mine monitoring systems control the underground mine, the preparation plant and the unit train load out. Pattern recognition IDSS/PRS produces a listing of begin and end times for each machine operating event as shown is Table 1.

Jan 1, 1991

By K. Norrena

Grade control in open pit mining requires the specification of dig limits that account for mineral grades, economic costs, and selectivity of mining equipment. Visual identification of ore rock types is ideal, but this is not always possible in lower grade disseminated deposits where ore/waste contacts are not visually discernible. In this case, conventional grade control practice consists of a two-step procedure (1) create a map of mineral grades at some selective mining unit scale, and (2) determine practical ore-waste boundaries or dig limits on the basis of the gridded block grades or assay information. This procedure is laborious, depends on a subjective assessment of where the boundary should be, and may be economically sub-optimal. We pose the determination of dig limits as an optimization problem and solve that problem with the technique of simulated annealing. Simulated annealing has the unique advantage of being able to combine multiple non-linear constraints into a single objective function. We use maximum profitability and the ability of the equipment to mine the proposed dig limits as constraints in the determination of optimal dig limits. We require a map of expected profit for eah block. Geostatistical techniques are recommended, for mapping expected profit but not necessary. Geostatistics will provide a quantification of the uncertainty in the grades within rock types using all available blasthole samples and exploration drilling. Some variant of L-optimal estimation or kriging can be used to determine the block-by-block classification that is economically optimum. We also need the expected profit for each block. It is unrealistic to assume free selection, that is, each block cannot be extracted independently of its neighbors. The optimal balance of ''accepting dilution'' and ''wasting ore'' is achieved to maximize profit subject to equipment constraints.

Jan 1, 2001

By Lit-Mian Chan, David H. Carlson, John H. Johnson

This paper presents an apparatus and procedures to implement a methodology for controlling ambient air quality in a confined working environment, such as an underground mine, an underground storage area, or a tunnel. The air quality control methodology is directed at limiting the quantities of exhaust pollutants emitted by diesel-powered vehicles into the confined space by identifying vehicles that are excessive polluters. The methodology is based on tailpipe and ambient air measurements in the confined space, and the ambient air threshold limit values (TLVs) for the various diesel exhaust pollutants. The methodology consists of two computational methods; a deterioration criteria computational method to identify excessive polluting vehicles, and an engine performance computational method to determine engine performance and to provide information needed to diagnose engine faults. The methodology is designed for implementation in conjunction with a preventive maintenance program. The first part of the paper describes the mine air and tailpipe emissions measurement apparatus and techniques used to obtain data. This is followed by sections discussing the deterioration criteria and the engine performance computational methods. A section discussing some common engine emissions-related faults is also presented. Lastly, a proposed procedure to implement the methodology including a possible hierarchical structure for mine personnel involved to implement the procedures is presented.

Jan 1, 1993

By Rajiv Hadkar, M. D. McKinley, J. B. Hiskey, B. J. Scheiner, G. W. Warren

Flyash from municipal solid waste incinerators containing roughly 13 to 15% Zn, 2% Pb and 0.2% Cd is an attractive material for simultaneous metal recovery and detoxification. Results have shown that Pb in flyash is present as PbSO4 and PbC12 and that about 70% of the Zn is present in the form of soluble amorphous compounds. Leaching results in a number of acid and basic lixiviants show that acid leaching is clearly more effective than basic leaching which is limited by low solubility of Cd and Zn. Cementation in acid solution with pH < 1 promotes redissolution of the precipitated metal. Successful zinc electrowinning from chloride leaching solutions has also been demonstrated, although Pb2+ and Na+ can be potentially detrimental. The presence of Pb2+ can prevent deposition of metallic Zn, forming zinc oxychloride, ZnC12.4Zn(OH)2, instead. High concentrations of Na+ can also result in precipitation of NaCl. This investigation has shown that hydrometallurgical treatment of MSW flyash through leaching, cementation and electrowinning is a potentially feasible method for simultaneous detoxification and recovery of metal values. Several issues have also been identified which would require further investigation for final process development.

Jan 1, 1993

By Thomas Oberndorfer

Recent publications indicate that there is a rising interest in the field of mining method selection, and increasing efforts are being undertaken to develop satisfying solutions. Yet, computer assistance for mining method selection (MMS), or in other, terms assistance in the (pre-)feasibility state of a project, still lags behind the computer application in other mining activities. This is true for research as well as for application. This paper is intended as a contribution to the development in computer-assisted mining method selection, and focuses especially on underground hard-rock mining. It is structured in four main parts. The first deals with general aspects of mining method. selection within the process of mine planning, and with problems and requirements which derive from this specific situation. In the second part some numerical MMS-approaches hitherto published are briefly introduced, and investigations on these approaches are discussed. Based on .these experiences, in the third part an analysis of the complex system MMS is presented. It focuses on heuristics, influencing factors, objectives, assessment-routines and mining methods, and emphasizes the essential necessity of a profound mining method description (MMD) specially developed for the requirements of MMS and numerical processing. Accordingly the fourth part of this paper presents a mining method description model developed at the Department of Mining Engineering, Montan-University Leoben (MUL), Austria. It describes the consequences of the investigation, in particular the new mining method description model, on &apos;the design of a MMS-system. The paper closes with a outlook on the possibilities of an successful MMS-system being realised.

Jan 1, 1992

By C. H. Perbix

[On Pier D in the Port of Long Beach, California, the Port of Long Beach has built a new bulk cement distribution terminal. The facility is operated by Pacific Coast Cement Corporation, who designed the complex and leased it from the Port of Long Beach on a 40-year lease. At one bulk cement is unloaded from vessels carrying from 12,000 to 30,000 t. The unloading is performed with pig speed mechanical unloading equipment and delivered into a concrete silo complex with, a capacity of 50,000 t. From the silos, bulls cement is loaded into trucks for delivery. The facility was completed in September 1981.]

Jan 1, 1983

By J. F. T. Agapito, R. R. Goodrich

Yield pillars have been used for many years to help reduce stresses near mine openings and improve roof and floor stability. A yield pillar is often defined as a pillar that fails but retains residual strength. Stress transfer occurs through the roof and floor after the peak strength of the pillar is reached. High stresses are transferred from around the openings onto abutments that can be barrier pillars or unmined ground. This mechanism, often referred to as pressure arching, is possible as long as the width of yield pillar mining (panel width) is less than the critical width above which stresses cannot be carried by the overburden. Significant stress transfer also can occur due to small amounts of pillar and/or floor yielding before the peak strength is reached. This is accomplished in a quasi-elastic manner with little or no visible roof and pillar fracturing or floor heave. Long-term stability may be achieved when stresses and mechanical properties are favorable to prefailure yielding. This paper gives practical examples where improvements in stability and resource recovery were achieved with this mechanism. Yielding was assessed by comparing measured and calculated vertical pillar stresses. Results indicated that calculated stresses in the pillars were 25% to 40% higher than the measured stresses, demonstrating significant arching load transfer to the abutments. Prefailure yielding is probably often present but unintentional in both development and production areas. Better recognition and use of this mechanism should lead to design improvements, as mines become deeper.

Jan 1, 2003

By Dr. -Ing. Rainer Alles

Introduction Belt conveyors usually run in a straight line. The correct alignment is a major criterion for reliable functioning and optimum running properties. Any misalignment of the conveyor, especially mistracking of the belt, can have many negative effects. As a consequence of this, damage can occur on the belt and conveyor, some material can he thrown off at the edges and serious running problems can arise. In the worst case the belt can come off its supporting structure so that operations are brought to a halt. The absolutely straight alignment of a belt conveyor is actually impossible so that on closer observation every belt conveyor contains more or less sharp curves. Whereas on non-movable conveyors the deviations from the straight can he kept within the range of practical measuring accuracy by careful installation, movable conveyors in surface mining or belt systems in difficult terrain often contain curves that are easy to see with the naked eye. Below ground, too, alignment cannot always achieved to a desirable extent or is changed by geological faults over the years. Despite these unwanted curves, most belt conveyors perform their function reliably. In view of the belt tension, the belt tends to take the easiest path. This in general is the shortest distance between the pulleys on which the belt is tensioned and bent. This means for example that for a curve radius of 10,000 m over a flight of 100 m the greatest deviation from the ideal line is as much as 12 cm. And a figure of this magnitude cannot normally he accepted. In such cases, use is made of guide devices to properly train the belt if it doesn&apos;t center itself anyway. If it&apos;s possible to control the belt position over such unwanted curves. it can also he expected that the belt tends to follow the desired curved path within certain limits. I&apos;d now like to cite a few examples to illustrate the ways and means of guiding a belt conveyor in curves. Cost-effectiveness The advantages of belt conveyors, in particular for longer conveying distances and mass flows have been known and undisputed for some time. However, there are often obstacles standing in the way of the concept of straight routing. If a belt conveyor is to bypass elevations or existing buildings or to follow the seam of underground deposits or a predetermined path, then substantially tighter radii and longer curves are desirable or necessary.

Jan 1, 1993

By James E. Lawver

Response surface methodology is a well-known and powerful technique for determining optimum conditions in flotation systems. One disadvantage is the onerous task of the numerical calculations and curve platting. This paper describes a system in which the investigator simply selects the range of several flotation variables after which the design of the experiment, data analysis, generation of response surface equations, and response surface plats are automatically carried out using a digital computer and a computer-driven curve plotter. Several typical response surfaces of important variables in phosphate flotation are presented as examples. An overview of the flow charting of the system is provided in the addendum.

Jan 1, 1978

By Ross W. Smith, Salim Akhtar

Flotation of oxide minerals, particularly silicates, with cationic collectors differs substantially from classic flotation of sulfide minerals using sulfhydryl collectors. In the case of cationic flotation of these minerals, in general: 1. There is a greater importance of electrostatic interaction between mineral and collector. 2. Longer-chained collectors are required; as a rule at least 10 carbons must be present in the hydrocarbon chain. 3. Moderate concentration of collector is required. 4. Attachment between collector and mineral appears to be less strong even though the system may exhibit a large contact angle. 5. Required conditioning time before flotation is usually relatively short. 6. There is sensitivity to slimes. 7. The collectors usually possess frothing power. 8. Selectivity of cationic collectors for specific minerals is usually poor. Almost any of the oxide and silicate minerals may be floated using cationic collectors. However, specific response to these collect- ors varies with mineral type. The following classes and minerals are

Jan 1, 1976

By Richard L. Breithaupt

INTRODUCTION A shuttle car is a self-propelled rubber-tired haul¬age vehicle designed specifically for underground min¬ing, primarily in coal mines (see Figs. 1 and 2). The drive mechanism is located on the sides of the vehicle, and the load is carried in the center. A chain-and-flight conveyor distributes the load when the car is being loaded by a loading machine or continuous miner and also discharges the load onto a conveyor belt or into a mine car. As the name implies, the shuttle car shuttles back and forth between the working face and the un¬loading point and is not required to turn around. Shuttle cars have served as primary haulage in this country since 1940 and still perform this function in most coal mines today. TYPES OF SHUTTLE CARS Shuttle cars may be battery-powered, diesel-powered, or cable-reel electric powered. Battery-powered cars were used between 1940 and 1950, but are almost out of existence today because of their limited power. Diesel cars have been used in coal mines, but are used mostly in noncoal mines where rules and regulations are not as stringent. Also, objections by organized labor have limited the use of diesel-powered shuttle cars in coal mines. Today, the largest number of shuttle cars are cable-reel cars. Electric power is fed through a cable stored on a cable reel that may have a capacity from 152 to 244 m (500 to 800 ft), depending upon the size of the cable and the cable reel (see Table 1). This, of course, limits the travel of a shuttle car. However, time studies have shown that for the most efficient haulage cycle, the length of haulage should not exceed 91 m (300 ft). As haulage distance increases, the waiting time for the loader or miner becomes longer, and haul¬age efficiency suffers. Cable-reel cars are manufactured in standard hand and opposite standard hand models. A standard car has the operator&apos;s control on the left front side and the opposite standard car has the opera¬tor&apos;s controls on the right front. Generally a loader or miner operates with two shuttle cars, one of each type. This allows the trailing cable to lie on either side of the entry and prevents the cables from fouling each other. Shuttle cars are designed to operate with either dc or ac power. Cars using dc power are designed for 250-v supply and ac cars are designed for either 440, 550, or 950 v, three-phase voltage. Most coal mines that are relatively level can use dc cars where dynamic brak¬ing is not required. For mines having steep grades (in excess of 15%) ac cars are used because ac traction motors have inherent dynamic braking and are safer when tramming downgrade. A recent innovation in shuttle car design in the last five years has been the introduction of an ac/dc car. This design utilizes ac input power through the cable, and the power is rectified on the car to 250-v do for traction motors. The pump motor and conveyor motor are powered by ac voltage. The main advantage of this design is efficiency-a much smaller cable may be used and the smooth operating characteristics of the dc traction motor is obtained. This prevents shock loading on the traction drive train, which is an undesirable feature of ac traction motors. SHUTTLE CAR FEATURES Elevating Discharge Today, all shuttle cars are equipped with an elevating front conveyor that allows the shuttle car to discharge into a mine car or belt feeder without the use of ramps. Hydraulic Power System Today, all shuttle cars are equipped with an hy¬draulic power system. An hydraulic pump is driven by an electric motor which provides hydraulic power for steering, elevating the front conveyor, and powering the

Jan 1, 1982

By E. S. Forbus

A process for producing reductive bleach in-situ in kaolin slurries has been developed. The bleach is formed by mixing an iron sponge powder (-325M) with slurry treated with SO2. The process has been successfully demonstrated in production, and applied to several grades. The color, brightness, and viscosity of the bleached product is equivalent to or better than that of product bleached with commercial sodium hydrosulfite . The tested products can be bleached at a cost, which is 38-54% of those associated with the use of commercial bleach.

Jan 1, 1993

By M. Rossi

Geologic ore controls are usually interpreted and modeled from cross sections and plans, as they are critical to construct meaningful and accurate ore reserves models. The geologic boundaries are often modeled as two-dimensional surfaces or three-dimensional solids and used as hard boundaries to constrain the grade estimates. In many cases, the geologic model is the most important factor in estimating the mineralized tonnage predicted by the reserves model, since it is generally drawn to separate the volumes of mineralized and waste material within the deposit. Because geologic models are interpreted from widely spaced drill hole data, the accuracy of the contacts of the different zones modeled could be poor. The lack of accuracy in the definition of geo-logic zones can have a significant impact on a resource model. In this paper it is shown that the geologic interpretation can and should be checked for potential biases. In particular, an unbiased overall volume of the original drill hole information with respect to the resulting solids and codes assigned to the block model is sought. In addition, geostatistical conditional simulations can be used to simulate the geologic attributes modeled, thus providing an alternative view of the geology, and a measure of the uncertainty of the geologic boundaries. The simulated geologic model provides an assessment of the impact of alternative geologic interpretations on the over-all uncertainty of the reserves model. In the case study presented here, the uncertainty of the geologic model is quantified for a difficult-to-model, high-grade copper deposit in northern Chile.

Jan 1, 2004

By R. Ciccu

Permeable reactive barriers (PRB) are one of the most promising systems for groundwater remediation. PRB are constructed with the purpose to intercept the contaminated flow and to degrade it as a big reactive filter. The most common type of reactive barriers is realized by excavating a trench and by filling it with a substance able to treat the contaminants dissolved in the water. Another type of PRB is obtained by realizing a line of wells in which chemically active compounds are injected. A new method for barrier construction has been studied at DIGITA?s laboratories. It makes use of the waterjet technology to increase soil permeability and to inject active compounds. This technology is especially suitable in the case of deep contaminated aquifers where traditional construction methods are difficult to apply. The paper describes the experimental apparatus especially designed and constructed and the activities carried out to investigate the relationship between waterjet operative parameters and the penetration of the jet in the soil. In conclusion, on the basis of the experimental results, the economical feasibility of the technique is evaluated.

Jan 1, 2008

By William Bogan, Mark Vaivoda, Richard Daniele

Many factors must be considered when performing engineering and construction work for a major minerals project. Dravo&apos;s Afton copper smelter in British Columbia illustrates some of the E&C factors considered when setting up this new operation. In March 1978, Afton Mines Ltd., a Tech Corp. subsidiary, began blister copper production at the world&apos;s first integrated copper ore mining, milling, and smelting operation to use Top Blow Rotary Converter (TBRC) process smelting technology. This key TBRC smelting element is licensed by INCO Ltd. to Dravo Corp. The TBRC process uses a single vessel of modular design. It rotates for mixing and provides complete processing capability, from concentrate through blister copper. The Afton facility is about 16 km west of Kamloops, a town in south-central British Columbia. To make its operation as economic as possible, Afton officials considered several options for the copper smelter, the first in modern times in the province. Inability to extract precious metals-primarily gold-from the concentrates ruled out hydro-metallurgical processes. The final choice was between a conventional electric furnace/ Peirce-Smith converter combination and the TBRC process. The TBRC process was chosen for its versatility in processing the various Afton concentrates and grades and its 40% lower capital cost compared to the electric furnace converter system. The TBRC smelter -including concentrate slurry receiving, drying, storage, smelting, off-gas cleaning, and copper casting-had an installed cost of about $25 million. In December 1974, a pilot program was conducted by Teck, Dravo, and INCO at INCO&apos;s Port Colborne, Ontario, research facility. The pilot program demonstrated the TBRC&apos;s ability to smelt, convert, and clean copper slags. After the successful pilot effort, Dravo performed additional research to find an effective fixation system for high grade, intermittent SO2 gases. Once the go-ahead was given, design and procurement for the Afton copper smelter was provided by the Denver Division, Dravo Engineers and Constructors. Commonwealth Construction, Vancouver, constructed all Afton facilities. Afton&apos;s single furnace smelter is designed to produce 25-kt/a blister copper in 363-408 kg cake form. The product is sold to England for electrolytic refining. The TBRC process is only one of several innovations proposed by Dravo and chosen by Afton to insure the most modern smelter in North America and meet all British Columbia&apos;s environmental restrictions. Conventional smelting methods use a separate furnace for each process step. The TBRC process employs but one furnace. The Afton TBRC vessel measures 6.4 m in length and 4.3 m in diameter. The furnace is supported on a set of motor-driven rollers. These permit the vessel to rotate at variable speeds to 40 rpm. Rotation provides good mixing, high reaction rates, and effective heat transfer from the refractory sides to the molten bath resulting in a low thermal gradient. The furnace can tilt through a full 360° for charging, firing, and pouring. Smelting at Afton is accomplished in sequential fashion. Flotation concentrate and limestone flux are initially charged to the hot, empty vessel. Using the water-cooled, oxy-fuel lance, the initial charge is smelted. Once sufficient molten bath is obtained and slagging completed, the metallic concentrate is charged. Materials are charged through the water-cooled, variable position additive chute mounted on the TBRC

Jan 10, 1980

By B. K. Mishra, B. Das, L. Panda, D. S. Rao

"The dissolution behavior of lateritic nickel ore containing 1.2% Ni, 40.6% Fe, 9.5% SiO2 and 9.59% MgO by sulfuric acid was investigated. The ore was characterized adequately by different techniques such as chemical analysis, scanning electron microscope (SEM), X-ray diffraction (XRD), field emission scanning electron microscope (FESEM), thermal analysis and Fourier transform infrared (FTIR) spectroscopic measurements. The characterization results did not reveal any distinct mineral phase for nickel; it was associated with mineral phases like goethite, hematite, chromite and serpentine in various proportions. The beneficiation studies by hydrocyclone, magnetic separation and flotation did not enrich the nickel value. Reduction roasting followed by magnetic separation indicated only 2%Ni with 28.9% yield. Consequently, the recovery of nickel values from the lateritic ore by acid leaching under different conditions was studied. The kinetics of leaching obey the first order rate equations. The recovery of nickel was found to be influenced by temperature. It was observed that at an acid concentration of 2M, leaching temperature 363 K, time 240 min and solid-to-liquid ratio of 1:10, it is possible to leach out more than 95% Ni.IntroductionThe demand for nickel is ever increasing. Nickel is used to impart specific properties to steel, titanium and aluminium, and also is used in petrochemicals, catalysts, batteries, etc. It has been estimated that nickel oxide ores account for about 70% of the world nickel reserves. However, most nickel is sourced from sulfide ores. Nickel is extracted from sulfide ore by size reduction, followed by flotation to obtain a nickel-rich concentrate. The concentrate is then pyrometallurgically processed by smelting to produce a nickel-rich matte after discarding an iron-rich slag. Due to the increasing demand for nickel in stainless steel and other alloys, nickel is also extracted from many secondary sources. The recovery of nickel from oxide ores is now gaining momentum, and more exploitation of this ore is being attempted by advanced as well as environmentally feasible, cost-effective techniques. Dalvi et al. (2004) has reported that, due to the increase in demand, the world nickel production from laterites is expected to increase every year from the current level of production. In recent years, investigations of the use of low-grade ores, spent catalysts, clays, etc., have intensified to meet the market demand and also due to a shortage of high-grade nickel bearing ores (Al-Mansi and Abdel, 2002; Baba et al., 2009; Ismail et al., 2011)."

Jan 1, 2015

By W. G. Pariseau

Equivalent elastic properties of discontinuously jointed rock masses are computed using a non-representative volume element(NRVE) approach within a three dimensional finite element program. Joint persistence, defined as the percentage of a joint plane that is actually occupied by joint material, is used in the NRVE computations as a step towards greater realism in stability analysis of mine openings. Persistence p is also the probability that a segment of a joint plane is occupied by joint material. Computations reveal a threshold phenomenon in jointed rock mass moduli similar to that observed in percolation theory. A rapid decrease in moduli occurs above a critical value pc(about 0.6). A detailed analysis of the well-known solution for stress about a circular opening in isotropic, elastic ground illustrates the effects of joint set orientation, spacing and persistence. Although stable “on average”, joint segment failures reveal a threat to stability and point towards rational design of reinforcement or support of discontinuously jointed rock masses.

Jan 1, 2003

By David Morrison

On the Corramandel Peninsula of New Zealand, the Waihi Gold Mine has undergone a major redevelopment. The upgrade included the construction of two very large overland belt conveyors, both rated for a capacity of 6000tph. The first of these conveyors (CVR2) traverses a total of 1600m over undulating terrain, including a significant section running through tunnels. A large hill is climbed, just after the mid-section of the conveyor run, before dropping back down prior the discharge point. The second of these conveyors (CVR8) traverses 1100m along an alignment with a horizontal curve, a number of vertical curves and a significant lift before the discharge point. The CVR8 conveyor also included a shuttling head pulley arrangement. The mine they service is inside the town of Waihi. The conveyor alignment runs through the town and out to the metallurgical process site on the outskirts of the municipal area. This paper documents the engineering process that took place in the design of these conveyors. The Martha Hill mine in Waihi, New Zealand is unique. The mine is placed in the township. Consequently there are quite stringent environmental limitations on the mine&apos;s operation. In [Figure 1] the proximity of the mine to the town is shown. In the top-centre of the photograph the alignment of the existing smaller overland conveyor can be seen leading from the mine to a tunnel through the hill in the background.

Jan 1, 2000

By D. Craig Camp

Microtunneling has successfully completed both horizontal and vertical curves. This paper covers the historical development of curved microtunneling and a description of one microtunneling project. The completed microtunnel is used for electrical transmission in Japan. The crossing had several concerns including right of way, social disruption, obstructions, the cost of intermediate shafts, and political restrains. In order for this drive to meet all of the imposed restrictions, four horizontal curves and one vertical curve were planned and executed in accordance with the design. Iseki Poly-Tech of Tokyo, Japan designed and built the microtunneling machine to meet the project requirements. Curved microtunneling has been successfully used on fourteen projects in Japan. The reason for a curve has been that the project could not be rerouted, short straight drives could not be used to negotiate the curve, or the project’s social and environmental conditions required a curved tunnel. The following table (Table 1) includes summary information on 14 curved microtunnling projects that have been successfully completed in Japan. The first column is a reference number. Columns two through four include basic project information. The fifth column is the total project distance completed by microtunneling. The sixth column is the project’s microtunneling duration. The seventh column is the machine type and size. All of the machines listed are manufactured by Iseki Poly-Tech of Tokyo, Japan. The size is the internal concrete pipe diameter; all of the projects listed were completed with Japanese standard concrete pipe. The eighth column contains a general description of the soil. The soil type and blow counts are approximately equal to ASTM D 2487 standards. The ninth column lists the radius of the curve and the span distance or the distance microtunneled that included the curve. The radius of a vertical curve is denoted with a “VR”, a span with both a left and right hand curve is denoted with a “S”. Other significant first events are also noted in this column. The first curved microtunneling project was completed in 1982 using a conventional microtunneling machine with standard Japanese concrete pipe and a very long radius curve while maintaining grade for the gravity sewer. The second curved microtunneling project included multiple radii curves, all in the same direction, on a single drive. In 1989, on the fifth curved microtunneling project, a “S” curve was successfully negotiated. On the seventh curved microtunneling project, completed in1995, a tight radius curve was negotiated. In 1996, the eighth project included a vertical curve. The thirteenth curved microtunneling project, completed in 1997, was the first project to utilize a microtunneling machine specifically designed to negotiate curves. The fourteenth curved microtunneling project was the accumulation of all

Jan 1, 2001

By Stephen LaBrooy, David Muir

Gold ores in the geologically more recent rocks are often associated with copper minerals which pose problems of high reagent consumption with conventional cyanidation practice. Many examples exist in northern Australia, Indonesia, Papua­New Guinea, Chile, and Brazil where, despite various treatment options (LaBrooy, 1992), the processing of oxide and transition ores is uneconomic either for copper or gold recovery. Hence, there is a need for a reagent which is selective for gold without leaching copper minerals. The ammonia-cyanide system appears to meet this criterion (Muir et al, 1989; LaBrooy et al, 1991; Costello et al, 1992; Hayes and Corrans, 1992) but it has proven difficult to optimise and apply to a variety of ores due to differences in the mineralogy and reactivity of ores and a lack of understanding of how the system works. Recent work on the mechanism, however, (Muir et al, 1993) has thrown a new light on the problem which should allow better process control. Ammonia­cyanide mixtures behave neither like free cyanide nor free ammonia and too much of either reagent in the mixture is detrimental. The key to the process is understanding how ammonia enhances gold extraction yet inhibits copper extraction in an essentially cyanide environment.

Jan 1, 1993