Search Documents

By Brandon Barnes, Michael Haggerty

"Around Minneapolis and Saint Paul, Minnesota, the local geology includes the St. Peter Sandstone, which is easily minable and has facilitated construction of interceptor tunnels for storm water and sanitary tunnels since the 1800s. The elevation of the geologic units allowed gravity-driven flows from the downtown areas all the way to the Mississippi River. Through many years of service, some of these tunnels have experienced degradation and require maintenance. Additionally, increased stormwater surface runoff can result in internal pressurization of the tunnels, accelerating degradation. Because structural failure or collapse of any of these interceptor tunnels would create a significant repair challenge, the city of Minneapolis has been inspecting and designating tunnels in its system for periodic review and repair. One of these, the St. Mary’s/Hiawatha tunnel, was rehabilitated using a grouting program. A hydrologic model was also used to improve understanding of pressurization events during some storms, leading to construction of a near-surface pressure dissipation chamber to provide increased hydraulic capacity. BACKGROUND The St. Mary’s/Hiawatha tunnel in Minneapolis has three segments; the first two were constructed in 1942 and the third was added in 1987 (Figure 1). Segment 1 is approximately 4,700 feet in length and outlets to the Mississippi River. Segment 2 intersects Segment 1 approximately 3,500 feet upstream of the river outlet and continues for approximately 2,500 feet in a southeasterly direction. The third and newest segment is approximately 500 feet long. There are four different cross sections of tunnel along the entire system.Originally, tunnel construction often consisted of mining the St. Peter sandstone with a pressurized water lance and constructing wood lagging support (Payne, 1967). An underdrain was generally installed below the invert during construction; this resulted in localized dewatering and provided groundwater control during construction. Non-reinforced concrete liner sections were cast in place. Often void space existed between the tunnel liner and the native sandstone bedrock. These voids, particularly in the crown, prevent direct contact between the liner and the bedrock."

Jan 1, 2016

By J. E. Urosek

The Mine safety and Health Administration (MSHA) formed the Mine Ventilation Bleeder and Gob Training Committee in 1993. This committee was charged with developing a training course designed to enhance the knowledge and understanding of practical bleeder ventilation systems. The course will probably include a discussion on the theory behind bleeder system design, an in-depth look at the strengths and weaknesses of the basic bleeder system types, case studies, and a review of the systems currently in use. A database on active underground bituminous coal mines, gob areas, and the bleeder systems presently used was developed from questionnaires distributed to nine Coal Mine Safety and Health (CMS&H) districts. Representative systems were selected for thorough investigation to develop case studies. This paper introduces some of the material to be included in the course, including a description of basic bleeder system designs, a review of a portion of the database, and facts about the gob areas and bleeder systems investigated for the development of case studies.

Jan 1, 1995

By B. M. Davis

Frequently ore reserve studies utilize samples from a series of drilling or other types of sampling campaigns. Sometimes there is a mix of sample types, say for instance, channel samples from underground workings and drill core from surface drilling. Since drilling, sampling, logging, and sample preparation protocols tend to change over time, there is always a question whether the samples give similar results that vary only by chance alone. The problem of comparing the various campaigns is complicated by the presence of spatial relationships among the samples. Geostatistical tests take the spatial correlation into account, but these tests may be complicated and cumbersome to implement. A comparison method based on familiar statistical graphics is presented which accounts for the spatial relationships. It allows comparisons to be made quickly by using simple statistical graphs such as scatterplots and boxplots.

Jan 1, 1996

By Remo Grandori

"INTRODUCTIONDouble Shield TBMs were introduced in the market in the early seventies with three main targets:• Increase the TBM performances in term of advance per day thanks to the ability to erect a segmental lining simultaneously with the excavation phase.• Be able to advance through unstable and disturbed rocks under the protection of the shields• Provide a finished tunnel by installing a precast lining as the TBM advance.In 2003 SELI developed the design of the Double Shield Universal (DSU) TBM as an evolution of the original Double Shield TBM design.This new TBM design was targeted to improve: the TBM capability to advance through squeezing/converging rock formations under high cover, the telescopic joint design (allowing by such to operate the machine in double shield mode in very weak rock), the capability of the TBM to investigate and treat the ground around and ahead of the tunnel face.In the same lapse of time Single Shield TBMs, originally utilized for soft ground tunnelling, extended progressively their range of application to mix rocks and finally to hard rock tunnelling with increasing performances.This also thanks to the substantial reduction in the installation time of the segmental lining ring that contributed to minimize the TBM stand by time between two consecutive boring strokes.Today the choice between the two types of TBMs needs to be carefully evaluated project by project since one type of TBM may “prevail” in certain geologies and diameters and vice versa.THE VISHNUGAD PIPALKOTI TUNNELThe Vishnugad Pipalkoti Tunnel is part of the 444 MW Vishnugad Pipalkoti Hydro Electric Project (VPHEP) on Alaknanda river in Uttarakhand (India).The Owner is the Indian electric company “Tehry Hydro Development Corporation” (THDC) and the World Bank finances the project.The Headrace Tunnel total length is 13.4 Km and a section of about 12,3km will be excavated by TBM by SELI Overseas (Italy) in subcontract to HCC (India) the General Contractor for the project. The TBM will be delivered ex workshop in March 2016 and it is foreseen to start boring in the second half of 2016.Figure 1 below shows the location of the project and the following Table 1 summarizes the geometrical characteristics of the TBM tunnel."

Jan 1, 2016

By E. D. Attanasi, M. W. Bultman, J. H. DeYoung

The 1987 recovery in the US mining industry permitted modest increases in exploration expenditures that year (Fig. 1). Favorable market conditions apparently allowed the gains to be sus¬tained in 1988. There was great uncer¬tainty early in the year as to whether world and US economic growth would be a casualty of the October 1987 worldwide stock market crash. By nearly all measures, the performance of the US minerals industry was strong¬. A weak US dollar and the strong growth of worldwide capital spending resulted in in¬creased US exports of con¬sumer and producer durable items as well as strong domes¬tic demand for metals. In fact, 1988 US steel industry exports reached levels not attained since the mid-1970s. It was re¬ported that steel industry labor productivity is now roughly twice the 1981 level. Among metals with signifi¬cant average nominal price increases in 1988 over 1987 levels were copper (43%), alu¬minum (52%0), and zinc (44%). Even though domestic copper production has grown since 1984 by 30% in mining and 24% in refining, em¬ployment in mining and milling is 4% below the 1984 level. Domestic unit costs are reported in the range of $1.10 to $1.33/kg (50 to 60 cents per lb). The US base metals industry is proba¬bly in its best cash position in many years. However, there has not been a large increase in exploration expendi¬tures. This is due mostly to the percep¬tion that worldwide capacity can easily be increased. Industry reports do show that "low cost" copper producers have been buying reserves from weaker com¬panies. The 1987 Society of Economic Ge¬ologists (SEG) survey of exploration activity indicated that 95 cents out of every exploration dollar spent by US and Canadian firms was for base and precious metals (Economic Geology, January 1989). Press reports, though, make it clear that most of that spending was targeted for precious metals. US claim staking data (Fig. 2) rein¬forces this suspicion. The apparent lag in new claim filings in Alaska and phe¬nomenal growth in filings in the lower 48 states suggests that exploration is directed at relatively low risk targets. In the conterminous 48, more than 40% of the new claims were made in Nevada, indicating an emphasis on gold (although even Nevada had some base metals exploration). Even with a price decline during 1988, the difference between the gold price and average global "cash" production cost of $7.30/g ($227 per oz) still leaves room for an attractive return (Northern Miner, 1988). Exploration plans and results Metals Economics Group (Halifax) indicated in its 1988 exploration survey that exploration expenditures by US and South African firms would be at about 1987 levels. It also reported that Canadian and Australian firms were ex¬pected to reduce their exploration spend¬ing significantly. Firms in both coun¬tries were caught in changes in tax laws. In Canada, the Mining Ex¬ploration Depletion Allow¬ance (MEDA) once permitted generous tax write-offs to in¬dividuals investing in miner¬als exploration. It was to be replaced in January 1989 with the Canadian Exploration In¬centives Program (CEIP). Although structured and ad¬ministered differently than MEDA, CEIP can provide an equivalent or possibly greater financial incentive for min¬eral exploration, for junior mining companies. In 1987, it is estimated that junior com¬panies accounted for two¬-thirds of the off-property (or grass roots) exploration ex¬penditures in Canada. Australia will revoke the tax exemption for gold in 1990. Australian operators, therfore, were expected to concentrate efforts on speeding up pro¬duction during 1988 and 1989. In 1988, Chile and Mali announced special efforts to attract gold explora¬tion by foreign companies. The Cana¬dian government continues to express concern about the write-off (and drawdown) of base metal reserves. Falconbridge did announce a new initiative to identify additional base metal reserves in the Sudbury district. Discoveries Gold deposits dominated discoveries

Jan 1, 1989

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By Hans Smit

The face of mining as we know it is set to change in the very near future as mining companies all over the globe realize the implications of accessing the vast mineral wealth contained on, and under, our ocean floors. With some land based mineral deposits being fully exploited across the world, deep sea mining is set to revolutionize global approaches to the search for precious metals and other sought after minerals. [ ] Previously, the financial implications of deep sea mining outweighed the benefits of access to the mineral rich deposits identified all around the world. The process of marine diamond mining off the west coast of southern Africa started in the late 1950?s by a Texan, Sammy Collins, and subsequently continued by De Beers to the present day, has proven to be an excellent proving ground for the development of technology and mining systems that ensure a sustainable and financially viable business. [ ]

Jan 1, 2008

By S. Bhattacharaya, S. Aich, B. K. Nandi

"At present scenario, there has been an increasing demand in the usage of blend coal in power sector; regardless of how, specific knowledge on burning and ignition characteristics and their interactive effects is stingy. In this work a detailed investigation was therefore taken up to study in detail the combustion characteristics of selected coal samples and their blends. Certain parameters were believed to have significant effects on combustion characteristics of individual coal and their blend. Major parameters identified for this purpose included proximate analysis, combustion conditions, heating rate, reaction regions, blending ratios, ignition condition, reactivity (or reaction rate) and burnout temperatures. With the help of these parameters, thermal features and reaction kinetics of the individual and blended samples were investigated through a qualitative and quantitative thermo-gravimetric analysis of two different individual coals and their blend carried out in an oxygen atmosphere from ambient to 800 0C at a constant heating rate of 10 K/min. A significant difference in terms of rate of mass loss with temperature correlated from Arrhenius expression along with coal-coal blend interactions and deviation from their expected behaviour, have been observed between the kinetics of individual and blend samples, the latter showing better ignition performance and comprehensive combustion phase. It was also found that particles having the lowest percentages of volatile matter and ash could possibly be blended with particles having high ash and volatile matter to meet the required burning condition which can be effective in terms of cost savings, improved fuel economy, reduced undesirable exhaust emissions and improved safety of coal burning equipment. Experimental results further indicated that individual samples and coal blend samples have different combustion features in terms of residue left, peak temperatures and activation energies. Thermodynamic activation parameters such as activation energy, pre-exponential factor and order of reaction were computed from the thermal data using the modified form of non-isothermal methods. INTRODUCTION Coal has been the most significant source of energy for India and would remain so at least for the next few decades. About 60-65% of the electricity produced in India is from coal fired power plants. Indian coal is of drift origin and of high ash content. Ash content of the coal burnt in power plants in India is in the range of 30-45%, most of it being Run-of-Mine coal. Most of the power plants in India source their coal from number of mines, most common being multiple sourcing. Usually coal from each source [1] is burnt separately. For various reasons mostly associated with logistics, burning of blended coal is rarely practiced. Blending of coal is practised mainly for improving the quality of available coal and meets the shortage of coal supply. Blending technique is generally practised in power industries to reduce the cost of power generation along with mitigating combustion performance or slagging problem ensuring consistency in as fired coal quality and also for environmental compliance. Blending good quality imported coal can improve coal quality and helps in preserving indigenous coal resources for longer period [2]. Use of blended coal shall reduce load on all boiler components (Mill, CHP, Boiler, AHP) & reduce O&M cost and increase the availability. The main purpose of blending is to get a blended as fired coal having ash, volatile matter and gross calorific value close to design coal and the two coals are to be blended in such ratio proportions that resultant coal blend have desired ash, volatile matter and gross calorific value [3]. For a coal to be blended certain parameters have to be considered to get compatibility with respect to combustibility ( Burn ability) of coal, grindability ( HGI), ash fusion behavior, swelling characteristics of coal, need to be checked"

Jan 1, 2017

By David R. Kalasky

One of the primary areas of computer application in mine ventilation has been the analysis of mine ventilation networks. A variety of network analyzers that are aimed toward the solution of flows in a given network has been developed. These programs typically use the Hardy Cross or other iterative algorithm for solution of any network, even those that are complex. The computer software has become more efficient and sophisticated over time and has thus been widely adopted for ventilation network analysis. One of the shortcomings of these programs is that they have been developed with the idea of analyzing the ventilation requirements at a single point in the lifetime of a mine. This provides the ventilation engineer with an important tool for analysis but falls somewhat short of providing a convenient means to economically com- pare alternative methods for ventilating a mine. The program described in this paper provides such a tool. The principal concepts that were necessary for the completion of a computer program that analyzes ventilation alternatives is the capability of efficiently determining the ventilation requirements at a variety of times during the lifetime of the ventilation system and some method of analyzing the costs of the alternatives on a realistic comparative basis. To accomplish this task, the authors began using Penn State's mine ventilation simulator designated as PSU/MVS, Didyk et al. (1977a, 1977b) and added the necessary additional software to accommodate the need for multi-period, time-dependent analysis of the net- work. The software was then tested on a coal mine ventilation problem with a 23-year life. The problem points out the value of such a program for ventilation analysis.

Jan 1, 1986

By Wayne Taylor, Martyn Dawborn

Simulation modeling of portions of two platinum mines using General Purpose Simulation Software (GPSS/H™) is presented in this paper. Both mines, located in the Bushveld Igneous Complex PIC) of South Africa, exploit Merensky and UG2 reef horizons, which are processed separately and therefore handled individually from the stope face to surface, requiring duplication of storage arrangements and batch operation of transport facilities. The first simulation assignment concerns the upgrading of a shaft system operating exclusively on Merensky ore, to hoist UG2 ore and increase production. The second assignment comprises a simulation of rock flow from the stope face to surface by way of tramming, inclined conveying and vertical hoisting. Development ends and stope boxholes were modeled down to individual hoppers (cars) demonstrating the versatility of the GPSS/H™ language.

Jan 1, 2002

By Jean-Luc Cornet

As modern conveyors are increasing in size and complexity, their control requirements are becoming more and more critical. In addition to the safety and equipment protection requirements, control systems are now called to implement intelligent drive controls derived from dynamic analysis modelling, as well as to provide fast and accurate diagnostics while keeping operator controls simple. These requirements led to the development of a new and innovative approach to control system development. This paper describes the control philosophy used by Conveyor Dynamics, Inc. (CDI) and some of the key features of our approach to the architecture of control systems.

Jan 1, 1993

By Mike O’Brien, Rob Croll

The Evolution of AngloGold -Incorporated in 1944 under the name of Vaal Reefs Exploration and Mining Company Limited -Name changed to AngloGold Limited in 1998 as the vehicle consolidating the gold interests of Anglo American Corporation -Diversification (geographic and orebody) through the sale of low-margin, deep level, short-life assets and the acquisition of higher margin, open-pit and shallow underground operations

Jan 1, 2003

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By B. R. Satio, P. G. West-Sells, D. B. Dreisinger, B. J. Y. Leong, G. A. Kordosky

Electrolytic techniques have historically been used to remove arsenic, antimony and bismuth from copper refining electrolytes. However, both solvent extraction and ion exchange techniques have been heavily researched and in a number of cases commercially introduced. In this study a large number of solvent extraction reagents and ion exchange resins were screened to determine the selectivity of the reagents for impurity removal. Solvating extractants were found to efficiently remove arsenic from electrolyte. The extent of arsenic extraction increased in the following order: TBP < DBBP < DPPP < Cyanex 925 < Cyanex 923. Antimony and bismuth extraction from electrolyte were demonstrated with the reagent 2EHAP04. The ion exchange resins UR-3300 and Duolite C-467 were found to strongly extract antimony and bismuth from electrolyte. C-467 appeared to be superior to UR-3300.

Jan 1, 1993

By F. Romero, L. Yacher, M. Orchard

Specific software tools for Statistical Multivariate Analysis combined with real time process data were applied to Concentrator Plant at Compañía Minera Doña Inés de Collahuasi SCM in Chile, basically to two Mill Lines, each one with one SAG mill (32’x15’8MW)and one Ball mill (22’x36’9.7MW). The main objective was to empower the operator actions using process patterns references instead of process variable set-points. Pattern clustering and the analysis of pattern deviations regarding normal or desired pattern, allows for improvements in process control and the further implementation of soft sensors for those variables not included in the initial model.

Jan 1, 2003

By Jordan Hoover, Erika Moonin, Chuck George

"In response to a severe regional drought and in preparation for the further decline of water levels in Southern Nevada’s Lake Mead, the Southern Nevada Water Authority (SNWA) built a new intake system to ensure water supply to Las Vegas and its surrounding communities. In 2015, design began on the Low Lake Level Pumping Station (L3PS) that will move the water from this lower intake, ensuring operability of the water delivery system despite these declining lake levels. The L3PS Project includes a 159-meter-deep shaft, a large underground chamber, and 34-each 152-meter-deep, 2.4-meter-diameter drilled well shafts to be constructed in highly faulted and fractured ground. The Construction Manager at Risk (CMAR) delivery method was chosen to facilitate early contractor involvement in the design process to address the significant challenges and risks anticipated on this US$650 million Project. This paper describes the CMAR selection process; the CMAR involvement during design and benefits resulting from that involvement; and the unique collaborative effort among the owner, engineer, program manager and CMAR used to evaluate and select the best design approach, preferred underground construction methods, and coordinated geotechnical baseline report. PROJECT BACKGROUND AND NEED Seven of the United States’ western states rely heavily on the Colorado River and its tributaries for their water supply. The primary source of water for the Colorado River is winter snow that falls in the mountains ringing the Colorado River Basin. Flows in the Colorado River vary significantly from year to year, depending on the amount of snow and rain deposited in the basin. Beginning in the year 2000 and continuing to the present, snowfall has been well below documented historical averages and Colorado River flows have diminished."

Jan 1, 2016

By P. James

"INTRODUCTION Dissolved salts build up in leach solutions over time as leach solutions mature. Iron can be particularly problematic for hydrometallurgical copper production as it can compete with copper for separation during solvent extraction (SX), leading to iron bleed through and reduced electrowinning (EW) efficiency. In its ferric form (Fe+3) it precipitates as a hydroxide at fairly low pH and can promote clogging and disrupted flow within leach pads as a result of localized pH deviations – such as might result from weather events. The solids formed can cause relatively irreversible degradation to pad productivity. The precipitation pH point for solids formation falls as the ferric (Fe+3) concentration increases, making the issue more prevalent as leach circuits for ore bodies containing appreciable iron get older. It can also be more of a concern when leaching lower grade ore like tailings where lower copper content may increase iron/copper ratios which may increase iron accumulation rates. The precipitation pH point for solids formation falls as the ferric (Fe+3) concentration increases, making the issue more prevalent as leach circuits for ore bodies containing appreciable iron get older. It can also be more of a concern when leaching lower grade ore like tailings where lower copper content may increase iron/copper ratios which may increase iron accumulation rates. Control of ferric (Fe+3) levels could be used to lower ferric (Fe+3) levels and keep the ferric hydroxide [Fe(OH)3] precipitation pH sufficiently high to avoid the associated solids formation and potential leach pad occlusion. Control of ferric (Fe+3) and ferrous (Fe+2) tenors and ratios may also provide benefit for improving sulfide ore leaching or bioleaching performance and effectiveness. Traditionally options for such control were rather limited with bleeding contaminants off or chemical dosing treatments being prominent methods. Now Blue Planet Strategies introduces a new option. Blue Planet Strategies (BPS) focuses on improved resource utilization targeting electrolytic treatment and waste conversion into resources and has now demonstrated that its proprietary DeMet™ (Dynamic Electrolytic Metal Effluent Treatment) platform is highly effective at efficient low-cost conversion and control of iron content in leach streams in addition to targeted dilute metal recovery.[1-6] This powerful new capability provides a new treatment opportunity to control leach solution ferric (Fe+3) and ferrous (Fe+2) tenors to enhance leach operation production."

Jan 1, 2015

By A. M. Galperin

Back analysis is a very important procedure for proper understanding of physical processes. Back analysis can be used to establish the role of different factors, define parameters to be calculated and formulate the behavior of natural or physical processes based on field experimental results. It is evident that, in order to obtain the necessary data for back analysis, field observations are required. For the Michailovsky region (Kursk Magnetic Anomaly, KMA) the long-term strength characteristics of the clay covering the quarry slopes were established with the help of back landslide calculations. The obtained data allowed quantification of the shearing resistance parameters at the presence or absence of intermittent dips of discharge in the pit bank massif structures. Based on the results of investigation for material composition, strength and deformation of finely dispersed (clay-chalk) deposits in the Kursk Magnetic Anomaly including measuring of the settlements of filled structures, complex probing and laboratory determination of water-physical and mechanical properties of "man-made" soils, engineering-geological zoning of filled structures was carried out.

Jan 1, 2005

By Barry N. Doyle, Tarcisio M. Carvalho, Alan K. Haines, Marcos P. de Melo, Edward J. da Silva

A metallurgical complex to recover gold from highly refractory arsenical ore was recently commissioned by Sao Bento Mineraco. The plant, with a nominal capacity of 20 000 tonnes per month, is situated in the Minas Gerais province of Brazil. The process consists of run-of-mine milling, flotation, pressure oxidation of the concentrate, carbon-in-leach, elution and electrowinning, and effluent disposal. The paper provides an account of the investigations that resulted in the decision to install the pressure oxidation circuit, process and engineering design and plant start up. A comparison is made between the results achieved during the first two years of production and the design parameters.

Jan 1, 1989

By Larry Taylor, Thomas Zdeb, Brian Jacobs, Dennis Jensen

Subway tunnels proposed for the Los Angeles County Metropolitan Transportation Authority (LAMTA) East Side Extension are to be excavated below the groundwater through an area of Los Angeles, California with dissolved hydrogen sulfide concentrations in excess of 200 mg/L. To address unique construction and worker safety issues, bench scale and field pilot test studies were conducted to evaluate various mitigation alternatives including in-situ oxidation by chemical injection, slurry pH adjustment to suppress H2S evolution, and chemical treatment to remove the H2S from the slurry and muck material. The results provided the basis for detailed hydrogen sulfide control specifications for the contractor.

Jan 1, 1999