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By S. P. Singh

Drifts are the nerve centre of underground mining. They provide essential communications between primary openings and production areas. In view of their strategic importance in the overall safety, economy and productivity of mines, an empirical approach for blast design during drifting has been proposed in this paper. The approach was developed with the objective of minimum blast damage and provides computerized blasthole patterns for drifting operations. The drift blasting practices and problems in Canadian mines have also been reviewed in this paper.

Jan 1, 1991

By R. R. Klimpel

The process of flotation is an extremely important solids separation technique employed in a wide variety of industrial process applications. Froth flotation is practiced in the coal industry to recover fines produced during the mining and cleaning of coal. A recent review article by Aplan (1) covers the processing variables encountered in the flotation of coal and the variety of chemicals and operating conditions which affect recovery. Numerous engineering and scientific studies have been performed with the goal of providing semifundamental laws of flotation behavior for improved screening of flotation chemicals, better equipment scale-up factors, etc. Often the results of such work are expressed in the form of mathematical models (e.g. 2, 3, 4, 5, 6) which are potentially useful for evaluating (simulating) on paper such alternatives as possibles process operating or configuration changes, the effects of ore and reagent changes, etc. It appears to be commonly accepted that a complete scientific/engineering characterization of flotation is immensely complicated and not sufficiently developed at the present to accurately predict (or understand) many important industrial facets of flotation. Thus, one is often reduced to detailed experimental characterization in both the laboratory and plant in order to fully evaluate flotation changes or alternatives. It is the purpose of this paper to describe a method of batch laboratory data characterization that is useful in evaluating flotation chemical reagents including froth flotation of coal. The approach consists of running a series of time-recovery curves in a batch float cell followed by the computer curve fitting of this data to a series of mathematical models based on semi-physical reasoning. The unknowns or parameters of these models are then optimally chosen so as to best fit the experimental data. It has been found in both coal and mineral systems that changes in reagent type, concentration, system pH, grind, etc. cause very consistent and logical changes to

Jan 1, 1979

By E. C. Watkins

The fiber glass industry is an industry that. hopefully is still in its infancy and far from a mature business. Fiber glass, as will be discussed in this paper, deals with continuous strand glass as used in the textile laminate and reinforced plastics industries. More specifically, we are speaking of glass used for printed circuit boards, tapes, paper, cloth, pipes, tanks, boats, recreational equipment, roofing shingles, and automotive applications. Compared to the flat, container and decorative glass industries, we are very young. The early patents on fiber glass applications were issued in the 1930's. The type of glass most commonly used in the world and that discussed in this paper is designated as "E" glass. This is a boro-silicate type glass having specific chemical and electrical resistance and low thermal expansion properties.

Jan 1, 1986

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

Ground water inflow into a mine and the re¬sultant impact on district hydrology can be pre¬dicted on a site specific basis once the ground water flow systems at the site are sufficiently well under¬stood so that a valid conceptual model can be de¬veloped. Ground water flow systems are controlled by the hydrogeologic properties of the materials at the, site, including hydraulic continuity along flow paths, and boundary conditions. Preliminary delin¬eation of hydrostratigraphic units for conceptual models frequently can be based on surface water data; specifically, spring locations and the gain-loss reaches of streams. Quantitative predictions of mine water inflow and its impact on regional hydrology require col¬lection and analysis of numerous types of data. Hydraulic conductivity must be defined quantita¬tively both parallel and perpendicular to bedding as well as areally for each hydrostratigraphic unit. The characteristics of the Davis shale in case his¬tory number 2 illustrate this concept very well. The storage characteristics also must be determined for each hydrostratigraphic unit and the potential or hydrostatic head should be quantified vertically and areally for each unit. The potentiometric measure¬ments in the Bonneterre Formation as discussed in case history number 2 illustrate this point. The data collection and data analysis techniques constitute the weak link in the analysis of mine inflow and district impact. The technology for char¬acterizing a mine site in the field has lagged behind the technology for mathematical simulation; we at¬tribute this development to the fact that computer modeling is relatively inexpensive compared to field data collection and analysis. On-site tests provide the best means of quan¬tifying hydraulic conductivity parallel to bedding and the storage characteristics of the unit. Pump tests with multiple observation wells provide the best information from on site tests. Hydrogeologie boundaries can be located from such multiple well tests. Multiport piezometers provide a means for determining the vertical potential in a single borehole. Borehole geophysics can facilitate the def¬inition of hydrostratigraphic units. Vertical hy¬draulic conductivity is difficult but not impossible to quantify with existing field test methods; labo¬ratory tests of cores fail to give a value represent¬ative of the field situation. The storage characteristics of a confined (artesian) hydro¬stratigraphic unit can be difficult to quantify when pumping changes the confined unit to an uncon¬fined (water table) case after depressurization and dewatering. Transmissivity and storage coefficient can be approximated to some extent for given field con¬ditions by computer modeling if enough is known about hydrogeologic boundaries. Modeling can be used in a trial-and-error approach to match transmissivity and storage coefficients to existing water level data corresponding to estimated discharge rates. A solution can also be obtained using a tech¬nique referred to as the inverse method. Unfortu¬nately, such a solution is seldom unique. Fracture flow models exist but practicality usu¬ally necessitates reversion to a basic Darcy flow model. Data requirements for fracture flow models could be difficult to fulfill where fracture aperture, spacing, orientation, and length are required. Ob¬taining such data at depth prior to mining would be impossible for any mine simulation in fractured media. We are of the opinion that most fractured media simulations can be accomplished using con¬ventional Darcy flow models by viewing the field problem on a macroscopic scale. The real challenge is to delineate hydrostratigraphic units properly in the field. The question of whether a broken-up fault zone is an aquifer or a barrier boundary is seldom simple. The fractures usually can be treated as an equivalent porous medium in either case. Only a few techniques exist for predicting un¬derground mine water inflow. These techniques fre¬quently are designed to evaluate shaft or tunnel inflow under steady-state head conditions in a sin¬gle layer rather than in multiple hydrostratigraphic units. Few computer programs exist that have the

Jan 1, 1986

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's control on the left front side and the opposite standard car has the opera¬tor'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 Andrew J. Gaber

Beginning in 1961 the Coal Research Bureau of the West Virginia University School of Mines began a program of research to investigate the potential of the magnetic separation process to remove sulfur of the pyritic form from pulverized coal. Using a so-called isodynamic magnetic separator (1) it was demonstrated that significant quantities of pyritic sulfur could be removed from coal. Pulverized coal samples representing several different coal seams were individually screened into sizes in the 48 mesh by 200 mesh (297 to 74 microns) range, and untreated and steam-air treated samples of each were processed by the magnetic separator. The results of the investigations to magnetically remove pyrite from coal were published in two separate reports issued by the Coal Research Bureau (2,3), The isodynamic magnetic separator was designed about 30 years ago for use in geological and mineralogical laboratories (1,4,5) to isolate one or more minerals exhibiting only small differences in their magnetic propertiese Specifically, to accomplish this, the air gap of the unit was designed to provide a nearly constant magnetic gradient both laterally and longitudinally in the separating zone. It is reported that mineral concentrates of 99.5 percent purity can-be produced on closely-sized fractions after several cleanings through the unit. The isodynamic magnetic separator, therefore, was not expected to process coal containing sizes smaller than 200 mesh, or to demonstrate commercial capability to process pulverized coal. The primary purpose of the Coal Research Bureau in using the unit was that of determining if the magnetic process could remove pyrite from coal, and thereby encourage further research directed toward development of acceptable full-scale processing equipment.

Jan 1, 1969

By A. E. Koski

This paper describes the implementation of specialized software in mine planning at the Cleveland-Cliffs' managed Empire Mine on Michigan's Marquette Iron Range. The application puts together years of mining experience on the eastern end of the Marquette Range beginning in the early 1960's. The Empire Mine, in its thirty-fifth year of production, uses the software to model a complex and highly variable low grade magnetite ore body, capturing the economics of mining and processing to the final product. With special attention given to geostatistical methods, the software is used for all geologic modeling, ore reserve estimation, ore control, pit design and short and long range planning.

Jan 1, 1999

By P. Grotjohann

In processing heavy mineral sands, the low feed concentration and respective high value of the desired minerals require high capacity production and sharp classification/separation techniques to reduce the specific production costs. The allflux® separator, a new process system introduced to the heavy mineral industry, meets these criteria in an economical way. Its two-stage process, a combination of rising current and fluidized bed technology, allows the treatment of extremely high massflows in a single unit. Using concepts that have proven successful in processing concrete sands for many years, testwork in coal, iron ore, and heavy minerals looks very promising.

Jan 1, 1998

By William G. Doepken

The extensive use of rubber-tired support vehicles as a complement to a fleet of load-haul-dump and drill jumbo units has proved invaluable in maintaining scheduled multilevel/multiheading development drifting advances at the Henderson Mine. Support equipment, whether self-powered or towed, releases primary drill-blast-muck equipment from the supportive functions of ground control, utilities installation, steel and concrete construction and supply handling, thus providing both balance and increased flexibility to an underground, rubber-tired mining system. Experience has indicated that increases in man productivity and improved safety and working conditions coupled with decreases in overall costs and mining cycle times are realized with the inclusion of support vehicles in the mining fleet.

Jan 1, 1976

By D. Kelley

Due to pricing volatility and lack of availability of metals, the use of non-metallic materials of construction in metal extraction and refining processes has increased significantly in recent years Even though costs and availability of metals has improved recently, these factors are not expected to continue into the future as metal supplies once again fall short of global demand. One of the most common non-metallic materials of construction used in metal extraction and refining processes today is Fiberglass- Reinforced Plastic (FRP). These processes often require materials of construction that can resist acids and high concentrations of acid chloride salts. As shown in Table 1, FRP based on epoxy vinyl ester resin has the same chemical resistance to acid and better resistance to acid chloride salts than alloy C-276 and is superior to 2205 stainless steel in both cases.

Jan 1, 2010

By W. M. Cross

The use of fillers and reinforcements for plastics has increased considerably over the past few decades. For example, in the 1980's the average growth rate of filler use was 5.8%. [Katz and Milewski, 1987] Similar filler use growth rates occurred in the 1970's due largely to the increase in price of petrochemical starting materials for plastics. While filler use is still sometimes predicated on the cost benefits associated with their addition to the plastic, other mechanical and performance advantages associated with filler addition are being realized, and account for most of the increased usage. Plastic filler types are many and varied, and include both inorganic minerals/metals and organic compounds, with the former type being, by far, the more common. Common mineral fillers include oxides, including both natural and synthetic silica; carbonates, particularly calcium carbonate; clays such as kaolin; and silicates like mica and talc. It has been recognized that a number of factors influence filler performance, including particle size, shape, surface area, and hardness to name a few. In addition, the interaction between the filler and the plastic is of paramount importance as this interaction will determine the wettability, bonding and the adhesion between the two materials, and ultimately, the mechanical properties of the composite. Froth flotation has long been recognized as a viable process to separate different mineral types. [Leja, 1982] Prior to flotation the minerals are liberated from the ore by comminution. The mineral particle size range produced by this comminution process is typically between 300 and 10 microns. This size range is also typical of mineral fillers for plastics. The size similarity between mineral fillers and minerals concentrated by froth flotation is only one of many similarities that can be found.

Jan 1, 1997

By G. F. Raymond

Discussion by G.F. Raymond Knudsen's study presents two curious conclusions: • The kriging of blasthole assays can systematically overstate mill head grades by as much as 21% as a result of unbiased sample variance. • No estimation method is able to reduce this bias to a small margin. The study is based on a simulation using real data and what are presumed to be actual variogram parameters from a real deposit. Although I have no doubt that the first conclusion (21% overestimation) is correct for the author's simulation, I do not believe this represents a realistic mining situation. Over the past 15 years I have done extensive comparisons between exploration drill-hole assays, blasthole assays and mill head grades on seven major open-pit mines, including some very erratic gold deposits. Commonly, nugget effects on blasthole variograms were 10% - 20% higher than on exploration variograms. And in one extreme case, the difference was 50%. Even in the extreme case, ordinary kriging on blastholes agreed well with the mill head grade over the long term. In Knudsen's simulation, the blasthole nugget effect is assumed to be 200% higher than exploration data's. He supports this by variogram plots from each. My guess is that the apparent, large difference between these variograms results from a failure to account for the proportional effect (blasthole assays are likely from a higher-grade area). A simple check would be a comparison of the variance of exploration samples nearest blastholes. As for a nearly conditionally-unbiased estimator of a large random error, the arithmetic mean of all of the data certainly qualifies, provided there is an even data spacing. As a corollary, so does simple kriging, which would include, in this case, a large weighting to the arithmetic mean. Similarly, using a large number of samples with ordinary kriging or indicator kriging would significantly reduce the bias in the case of a large nugget effect for the variogram.[ ] Reply by H. Peter Knudsen Raymond questions two conclusions in the paper. First, he wonders whether a 21 % overestimation represents a realistic mining situation. I agree that 21 % is high, but overestimation in the range of 10% to 15% is certainly common in my experience. Furthermore, several years ago I consulted on a gold mine that was experiencing a 45% overestimation due, predominantly, to poor blasthole samples: In further questioning of the 21 % value, Raymond wonders whether the nugget effect of the blastholes is really so much larger than the nugget values of the exploration data. It is consistently larger throughout the deposit. In my experience with six Nevada gold mines, the high nugget value is not unusual. In fact, for some reason, nugget values for blasthole samples are typically about 0.0005 (opt squared). I am of the opinion that this high nugget effect observed at many gold mines is predominantly due to the inherent inadequacies of the blasthole sample and subsequent sample preparation. The second conclusion Raymond questions is the inability of the estimators tested to reduce the conditional bias. In fact, the conditional bias is extreme with the polygon estimator and greatly reduced by ordinary kriging. However, it was not eliminated. Raymond suggests using simple kriging, or perhaps a larger number of samples, to reduce the conditional bias. The technique of simple kriging may be less affected by the random errors in the data, but I did not test the technique. Using a larger amount of data presupposes that too few samples were used initially. In ordinary kriging and indicator kriging, the screen effect comes into play and ensures that samples beyond the second screen are given zero weight. Hence, increasing the sample size does not change the estimates nor the conditional bias. The main point of my paper is that the random unbiased errors (a fact of life in blasthole samples) cause a conditional bias in our estimates. The mechanics by which the conditional bias is introduced are nicely explained by Springett. My paper simply shows that the bias is also present when working with linear estimators, such as ordinary kriging, and even with nonlinear estimators, such as indicator kriging.[ ]

Jan 1, 1993

By Martin P. Schroeder

"Challenges have emerged as subway systems age and as new tunnels are bored in congested corridors or in high-risk environments. To help public transportation agencies address tunneling challenges, the American Public Transportation Association has convened numerous independent industry peer reviews composed of top experts in their fields to come together and provide recommendations and solutions. This paper provides case histories of peer reviews that include topics such as boring through toxic soils, boring underneath existing tunnels, retrofit of tunnel ventilation systems, and mitigation of water intrusion. Peer reviews are applicable to all tunnel projects worldwide. The paper will highlight the conditions of these challenges, peer review recommendations, steps taken to reach agreement, and the successes that resulted, which led to major changes, new legislation and reduction of risk. INTRODUCTION Public transit agencies face many challenges in maintaining and expanding safe transit operations, especially as systems age, ridership expands, and budgets remain flat. The development of new industry standards and federal safety regulations places further demands on transit systems, with compliance often taxing in-house resources. Studies of the state-of-good-repair across US public transit systems indicate a current backlog of $84 billion in maintenance projects, which will require an investment of $18.2 billion per year to eliminate the backlog by 2030[1]. Systems not in a state-of-good-repair may not operate efficiently, may not be available to meet the growing ridership demands, and may not meet required safety goals or reduction of risk. Innovation is another clear need transit agencies have for improving system performance and lowering operating costs.To meet these challenges, transit agencies often turn to industry peer reviews for guidance and recommendations as part of their overall strategy to maintain and grow the safe operation of their systems. As a service to its members, the American Public Transportation System (APTA) has developed an industry peer review program designed to give agencies opportunities to learn from each other and help them address challenges in operations, engineering design, and construction of their systems. Over the past 20 years, APTA has performed hundreds of peer reviews covering all aspects of transit system operations from fare collection to agency organization, project delivery methods to safety culture, and engineering design including tunneling and tunnel fire ventilation. As rail transit systems expand to meet increased ridership, it is often necessary to bore new tunnels for alignments in congested areas where surface rights-of-way are lacking. Such proposed alignments can pose significant challenges for transit agencies, as they contend with various tunneling problems."

Jan 1, 2016

By J. A. Cremeens, D. D. Overton, J. C. Varnier

"Slope failures that occurred in 2008 and 2009 at the Pikeview Quarry in El Paso County, Colorado, USA were investigated using three-dimensional discrete element modeling methods. Large scale block sliding failures occurred along weak bedding planes that dip steeply into an excavation located at the toe of the slope. Several mining sequences at the toe of the slope were simulated to evaluate the effects of various degrees of extraction and backfill on slope stability. Slope stability back analysis was conducted using pre-failure and post-failure conditions to determine the joint shear strength properties. Laboratory testing was also conducted to determine intact rock properties. Three-dimensional discrete element analysis was uniquely applicable to this evaluation because it allowed representation of shear stress of the basal sliding surface and it afforded evaluation of several scenarios of extraction and backfill at the toe of the slope. INTRODUCTION The Pikeview Quarry is an aggregate quarry located approximately five miles northwest of Colorado Springs in El Paso County, Colorado, USA. The quarry has been operated by several owners since the early 1900s. The Pikeview Quarry has been a major source of construction aggregate for the Colorado Springs area for many years (MSHA 1986). In the 1950s, the site was mined by Peter Kiewit and Sons to provide limestone for the construction of the nearby Air Force Academy. A large scale slope failure occurred on December 2, 2008, which stopped mining activity at the Pikeview Quarry. A second, related slope failure occurred on September 13, 2009. The second slope failure was located immediately adjacent to the initial December 2008 failure. The objective of this publication is to discuss the three-dimensional slope stability evaluation. The causal relationships of slope instability are not within the scope of this publication. However, omitting the precursors to slope failure would limit understanding of the nature of the slope stability evaluation. The regional and local geology and the history of slope instability at the Pikeview Quarry aid in understanding these precursors. Two main failure mechanisms were hypothesized during the failure investigation to explain the triggering of the 2008 and 2009 large slope failures at the Pikeview Quarry. Both mechanisms are given equal credence in this publication. Quarry excavation was a key factor in the slope instability for both failure mechanisms. The objective of the three-dimensional model was to evaluate the effects of extraction and backfill at the toe of the slope with respect to slope stability. GEOLOGY Regional and local geology are both important factors in slope instability at the Pikeview Quarry. Geologic structure at the quarry is related to regional mountain building and basin development associated with the Laramide Orogeny and the Rocky Mountains. Lithologies at the Pikeview Quarry are also related to regional trends. Localized geologic features, primarily related to weathering and erosion, have also been major contributing factors to slope instability at the Pikeview Quarry."

Jan 1, 2017

By R. D. Cook, J. L. Johnson

"Measuring mud bed level in a sedimentation device is an important control variable. Methods used to date for this purpose have had limited success providing continuous, consistent measurements at a reasonable price. This paper describes a new instrument for detecting the mud bed / liquor interface as well as bed density profiles. The state of the bed is measured directly with sensors located in the thickener. Data from bench-scale and full-scale tests show that the bed level is well defined and easily detected. In addition, full-scale testing revealed unprecedented information about thickener beds. Variation in bed density gave new insight into thickener bed dynamics, providing design and control opportunities never before possible. This novel instrument can significantly improve thickener control by providing reliable and consistent bed level measurements. It is also capable of measuring bed density profiles that can potentially identify significant operating problems early enough to take corrective action before experiencing costly shut downs. INTRODUCTION Numerous studies and papers have been devoted to the subject of thickener design and operation to achieve specified throughput and discharge densities. Mathematical models have been developed that present methods for sizing thickeners and predicting performance under varying process conditions. The compression effect of deeper beds on dewatering performance is a recognized phenomenon (Wilhelm and Naide 1981). This is one variable of many that are inter-related and influence thickener performance. For example in the steady-state, continuous operation of a thickener, if the solids feed rate increases and the bed level remains the same, the underflow discharge rate must increase with the result of solids concentration being lower. The interactions of these parameters are at the heart of thickener models. One such model (Burger and Narvaez 2007), predicts solids concentration profiles increasing downward and discontinuities between settling zones. This model uses sediment height (bed level) as a parameter that influences capacity calculations to achieve the specified solids concentration of the underflow. Knowing the bed depth is critical to maintaining stable thickener operation. Implicit in these modeling strategies are some basic assumptions such as homogeneous distribution of solids in the different settling zones, plug flow of solids through the settling zones and steady state operation of the thickener. The accuracy of these assumptions has not been verified because there has been little visibility inside the bed of full size operating thickeners. Testing of the instrument presented in this paper demonstrates that it provides a degree of visibility of the dynamics in a thickener not previously possible and indicates that the basic assumptions of the modeling strategies may not be completely accurate."

Jan 1, 2015

By Robert R. French

Commercial grade gypsum beds up to 16 feet thick is present in southwestern Indiana. The major deposits are found at a depth of 350 feet to more than 500 feet below the surface. The ore body averages about 75 to 95 percent gypsum. Two companies are currently exploiting a deposit which underlies an area of several square miles near Shoals, in Martin County. Another deposit has been located in southwestern Owen County but has not been developed. The Indiana Geological Survey is currently drilling in Greene County and is considering further exploratory work in Perry end Dubois Counties.

Jan 1, 1965

By R. C. Milici

Historical production data and reserve decline rates can be used both to predict future coal production rates and to estimate when the reserve will be depleted. Reserve estimates inferred herein from historical production data and decline rates are commonly lower than estimates made from geologic data. The latter include significant amounts of lower quality (deeper, thinner, higher sulfur) coal which may be mined when the better deposits are depleted.

Jan 1, 1997

By Richard C. Marshall

The heightened concern over the environment which has developed in recent years has caused activities on National Forest lands to come under more intense public scrutiny. This has been especially true for activities oriented toward the production of commodities. Cattle grazing, mining, timber harvesting, and oil and gas leasing and drilling have all been increasingly subject to appeal and litigation. As a result, environmental analyses being conducted by the Forest Service have become more comprehensive, and measures to reduce or eliminate adverse impacts on resources have been given more attention. The costs of such environmental mitigation measures can be significant even when viewed in the context of the millions of dollars a company may spend, for example, to develop a mine. Using the Montanore Project in northwestern Montana as a case study, this paper examines the costs of alleviating some of the detrimental effects which may occur as a result of a mining operation. Although many of the general issues that have been raised about Montanore are common to other mineral projects, the scale of the project, as well as the fact that the ore body is within a wilderness area, the site is in grizzly bear habitat, and the surface waters are considered to be pristine, all combine to make Montanore particularly challenging from an environ- mental standpoint. Much of the information which follows is contained in the most recent version of the Draft Environmental Impact Statement (DEIS) which is scheduled to be published in several weeks. Based on agency and public comments, some of the information presented in the paper will change before the Final Environmental Impact Statement is issued in 1991.

Jan 1, 1991

By Peter V. Avotins

During last year's AIME meeting in New Orleans, Deepak Malhotra and I were discussing some of the topics that would be presented in this Symposium. It seemed appropriate that along with papers on specific reagent and flocculant types, we should address research management aspects of new product development. This development is a long, arduous task that requires a significant technical and fiscal commitment by the chemical company and potential users. A flow chart depicting the research process is shown in Figure 1. The sequence of events are plotted on an undefined time line. Our experience has shown that it takes 3-5 years to complete a typical reagent development project. A new project starts with an idea of how to do something better. The ideas range from relatively simple product modifications to opportunities created by technological change. An example of this is the use of flocculants in tailings operations to close the water recycle loop. The change was brought about by the need to conserve water and comply with environmental statutes. The result was the development and wide-spread use of synthetic polymers to help in solid-liquid separation. More recently the ups and downs of oil prices have led to development of coal water slurries as an alternative to fuel oil. The slurries contain high levels of chemicals (up to 3%) to insure proper stability and burn characteristics. There are technical changes under way right now that may lead to ideas for other new products. One exciting source of new technology is the ability to manipulate genetic material of microbes to select for specific capabilities. This has evolved by leaps and bounds over the past 5 years. Biologically assisted leaching and removal of wastewater impurities are two possible end uses. I'm sure that there are many others. The ideas can originate anywhere and are often half-baked at first, but to lead to new products they must be refined to provide project definition and eventually a research plan. A chemical ,development project cannot be started without approval to commit the company's resources. The internal selling job is often very difficult. The company must see clear financial benefits in order to launch efforts that may not pay off for many years. It is critical to project the total market, profitability, eventual market share, and timing of introduction of the new product. There is much uncertainty in this and the research manager is often skating on thin ice to retain his credibility. Once we obtain support from upper management, the research team is assembled to start the project. The teams generally consist of organic chemists, research metallurgists, and technicians. They have access to analytical scientists and technicians for surface and analytical chemistry. The projects are often augmented by help from university staff and consultants. This is particularly true where there is a need to develop fundamental understanding of the technology in addition to synthesis of chemicals. The project now proceeds along two lines. First, there is an emphasis on understanding the prior art. Extensive literature and patent searches are undertaken to provide research leads and to keep from "reinventing the wheel" and infringing on someone else's patent. The scientists and technologists are qualified to conduct computer-assisted library searches. This process usually uncovers 80-90% of the relevant information. The remainder is very much harder to get and requires the help of professional library search staff with scientific backgrounds. Along with literature searching, the team begins synthesis of new compounds. These are screened by use of techniques such as microflotation or filtration in a Buchner funnel. The scientists are looking for the direction of response as they modify structural features . Inevitably, a large number of candidates reach the next stage, a laboratory bench test. At this point it is possible to set up a research plan and use project management techniques to monitor progress. The bench tests are conducted in larger equipment, usually requiring 100-1000 grams of test substrate. Typical laboratory flotation tests are run in Denver or Wemco cells.

Jan 1, 1986

The mining industry is one of the three pillars of the Namibian economic development, besides tourism and the fishing industry. Generating 20 % of the GDP and more than 50 % of the export income, half of this amount derives from the diamond mining industry. In 2008, the annual production of 2.2 mio carat achieved an income higher than 1.44 billion Namibian Dollars (more than 750 mio Euro). After the global recession in early 2009, which also caused a significant downcrease in production to only 0.93 mio Carat, the diamond mining industry easily restored itself and the production figures for 2010 are expected to be even higher than those achieved in 2008. Compared to other diamond producing countries, Namibia is mainly producing diamonds of highest quality (jewellery quality), with less than 10 % of the mined diamonds not falling into this category. This high quality stands also for the worldwide highest prize achieved per carat. With 370 $US per carat it is more than four times higher than that of the average South African stone. In 1908 the finding of the first diamonds in the Namib desert close to the coastal town of Lüderitz was reported by a Mr Stauch, railway inspector for the then German colonial railway company. This event immediately triggered a Diamond Rush and subsequently the set-up of the restricted diamond mining area ?Sperrgebiet? through the German colonial administration. In 1913, only five years after the first mining operations had started, the steadily increasing production allowed for more than 20 % of the diamond production worldwide. After the First World War, the new South African government concentrated all mining activities in one organization, CDM (Consolidated Diamond Mines of South West Africa) which in following decades conducted an extensive and detailed exploration of the whole Sperrgebiet and started numerous new diamond mining operations.

Jan 1, 2011