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By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger's and Kick's laws (Bond, 1952). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed. The work input is given by the following equation: [111W = 1 Owi( FP FF(1)] where W is the The energy input to the mill, (kWh/st); F is the Feed size based on 80% passed, (µm); P is the Product size based on 80% passed, (µm); and W; is the Work Index, (kWh/st) Methods to determine the Bond Work Index Various methods have been proposed by Bond to determine the Work Index. There are some problems with the Standard Bond Grindability Test: •Requiring a special mill. •Requiring a relatively large amount of sample. •Being relatively time consuming and tedious. [ ] During the recent years, numerous attempts have been made to install faster and easier methods to determine the Work Index. A complete description of these methods is presented by Yap et al. (1982). Determining the Bond Work Index by a new size ball mill During the recent years, researchers at the University of Tehran determined the Bond Work Index, using the Anaconda simplified methods by its Denver 305- x 127-mm (12 x 5-in.) laboratory wet ball mill. The equation used to calculate the Work Index for the above mill containing 16.12 kg (35.5 lbs) of steel balls is as follow: [1 _ 1-1W;=0.54x [P [)(2)] In addition, an attempt has been made to determine the Bond Work Index similar to standard Bond ball mill procedure but by a different ball mill, smaller and in consequent portable and less laborious than the Bond ball mill. Method description Just as the standard Bond grindability test, this method is a locked-cycle dry grinding and screening process that is carried out until steady state condition is reached. Test equipment The test is conducted in a special ball mill, called the New Size Ball Mill (NSBM). The mill is scaled down with a coefficient of two-thirds from the standard Bond ball mill. It measures 200 x 200 mm (8- x 8-in.) inside (instead of 300 x 300 mm or 12 x 12 in. in Bond ball mill). Just like the Bond

Jan 1, 1995

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

It is imperative that the hydrogeologic frame- work and the ground water flow systems of a Potential mine area be investigated prior to initiation of mining. Such investigations are necessary so that potential mine water inflow problems can be anticipated and controlled. Obvious difficulties can arise if large inflows of ground water occur in underground workings. Costs of water removal and disposal, if not foreseen and included in economic planning, can severely limit or eliminate profit from a mining operation. Safety problems associated with large inflows are an additional potential problem. The inflow of water into a mine necessarily alters the associated ground water system. The inflow also may impact nearby surface water regimes, depending upon the interaction of the surface water and ground water systems. The inflow of water into a mine inevitably causes a decline in the ground water levels in the vicinity of the mine. This decline may be restricted to a small area or it may extend a considerable distance, depending upon the hydrogeologic properties of the associated hydrostratigraphic units. A decrease in surface flow will result if the area of ground water level decline intersects a stream or spring that is hydraulically connected to the ground water system. Mine inflow problems historically have been evaluated after-the-fact. Premining hydrologic investigations only recently have been initiated. Brown's (1977)* report on anticipated hydrologic impacts of oil shale mining in the Piceance Basin of Colorado is an example of premining hydrologic planning. Observed difficulties inherent in designing a data collection network and then utilizing the data to predict mine inflow have led us to the preparation of this manuscript. This manuscript reflects a two-phase effort. The first phase describes and develops analytical and numerical techniques for mine inflow prediction. Emphasis in this section is placed on the selection and application of data collection techniques and data analysis procedures. The second phase applies the prediction technique(s) developed during phase 1 to two existing mines. The objective of phase 1 efforts is to evaluate the hydrologic variables associated with surface water systems and ground water systems that can affect the inflow of water into an underground mine. Subsequently these variables are incorporated into predictive techniques for estimating underground mine water inflow and its effect on associated ground water and surface water systems. The specific objectives of phases 1 and 2 are to: 1) Determine the hydrologic parameters required for predicting underground mine water in- flow and the potential effects of the inflow on surface and ground water systems. The hydrologic parameters represent both premining and past initiation conditions. 2) Outline methods for obtaining the necessary input data for the predictive techniques in pre- mining and post initiation periods. 3) Present existing numerical techniques for predicting underground mine water inflow including assumptions Ad required input. 4) Evaluate existing numerical techniques previously tabulated based on their applicability to mine water inflow prediction. 5) Attempt to develop analytical techniques for predicting underground mine water inflow. 6) Develop, through modification of existing models, a general mathematical model as a numerical technique for predicting underground mine water inflow and the resultant effects on surface and ground water systems. 7) Apply the predictive techniques to two existing mines selected for this purpose. 8) Compare the predicted inflow rates to measured inflow rates at both mines. 9) Compare predicted impact on district hydrology to observed impacts where possible. 10) Evaluate the predictive technique.

Jan 1, 1986

By Jay S. Silhanek, Vernon E. Andrews

[Introduction] Section 122 of the Clean Air Act Amendments of 1977, Public Law 9595, directed the Administrator of the Environmental Protection Agency, to review all relevant information and determine whether emissions of radioactive pollutants into ambient air will cause or contribute to air pollution which may reasonably be anticipated to endanger public health. In August 1979, EPA published a preliminary report on assessments of public health impact resulting from emissions of radionuclides into air from a broad spectrum of major source categories (1). A section of this document presented a theoretical analysis of the radioactive airborne emissions from several non-uranium mines including iron, copper, zinc, clay, limestone, fluorspar, and phosphate. Since 1978 EPA's Las Vegas Laboratory has been gathering field data on actual radionuclide emissions from these mines to support the earlier theoretical analysis. The purpose of this paper is to present the results of those field measurements in comparison with the assumed values for the theoretical analysis. Data were presented in EPA's preliminary report on both gaseous radon-222 and particulate uranium-238 emissions from non-uranium mines. Most of the emissions data were rough estimates based upon the uranium content of the ore. Preliminary data from a few field studies were also used. Although both the theoretical and field studies by EPA were concerned with radon-222 and radioactive particulate emissions, this paper will concentrate only on the radon effluents. Radon from mines seems to have the greatest potential for possible health effects due to radiation. Data on particulate emissions are presented in other technical reports (2-5). [Theoretical Study] The theoretical emission data are presented in Table I. The estimates for the iron, zinc, and clay mines are based upon preliminary measurements taken by the Las Vegas laboratory. The estimates for the other mines were made assuming that 20% of the total radon-222 in the ore was released during the mining process and the radon release rate from the exposed ore surface was a factor of two greater than that resulting from mining. [ ] Dose assessments for radon from these mines were carried out using model facilities located on generic sites (1). The AIRDOS-II computer code (for estimating dispersion of airborne pollutants) with some minor modifications was used to determine working level exposures to a maximally exposed individual and a regional population. A conversion factor for risk/WL-year(O.021 fatal cancers/person-WL-y) was developed primarily from information on uranium miner exposures (6). This computer code and conversion factor were then applied to the assumed emission rate of radon-222 from each type of mine to develop estimates of working levels, lifetime risks and number of fatal cancers. A summary of this data is presented in table II. [ ] The highest estimated number of cancer deaths per year of operation is 0.05 cancer deaths or 1.0 death due to radon during the 24 year life of a phosphorous mine. For comparison, EPA estimates that underground uranium mines would cause .03 cancer deaths per year of operation in a regional population(1). [Field Study] As a followup to the theoretical study, EPA's Las Vegas laboratory began making measurements in 1978 of actual emissions of radon-222 and radioactive particulates from mine vents at several facilities for each of the industries listed in Table I and II. They were intended to verify general information and radon emissions from EPA's theoretical estimates.

Jan 1, 1981

By Carrie Smith

This fall, the Society for Mining, Metallurgy & Exploration (SME) will begin a new chapter by welcoming its new executive director and chief executive officer (CEO). Melissa Russell, a successful nonprofit executive with more than 30 years of experience in global association management, officially stepped into the role on Sept. 29, 2025. She succeeds longtime CEO Dave Kanagy, CAE, who is retiring in March 2026 after more than two decades of transformative leadership. “I am honored to begin this new chapter with SME and inspired by the vital role this community plays in shaping our world,” said Russell. “I look forward to listening, learning and working alongside this remarkable group of professionals to advance innovation, champion sustainability and create a brilliant future together.” Russell brings a deep portfolio of leadership experience in technical and industrial associations to SME. Most recently, she served as executive director of the Institute of Electrical and Electronics Engineers (IEEE) Computer Society, the largest of the IEEE’s 39 societies, with more than 50,000 members across 150 countries. During her tenure, she strengthened global engagement, advanced financial sustainability and launched strategic humanitarian initiatives.

Oct 1, 2025

By T. C. Page, P. C. Gibson, M. A. Miller, J. D. Sell

[The Yarnell gold deposit in the Weaver mining district, Yavapai County, Arizona, occurs in a structurally controlled, hydrothermally altered zone in a 1700 Ma granodioritic intrusive. Both potassic and sericitically altered rocks around and above the low-angle, northeast striking Yarnell fault are known to host economic gold mineralization. A wider envelope of weakly-propylitic altered rock also occurs in this area. During mineralization, strong sericitization accompanied several stages of quartz (±adularia) veining, stockwork formation, localized silicification and potassic replacement, in association with deposition of specularite, pyrite (now oxidized) and gold. The footwall of the fault is also sericitically altered but poorly mineralized. Mineralization along the Yarnell fault continues both northeast and southwest from the main deposit, although the thickness of the zone and associated alter avian envelope diminish away from the orebody A sample of undo formed illite taken from the Yarnell fault zone was K/Ar dated at 69 + 1.6 Ma. This date reflects a minimum age for both mineralization and latest movement on the fault. Gold mineralization is accompanied by modest increases in Ag, As. Cu and Mo content. Both CO2- and H2O- rich fluid inclusions were studied and suggest mesothermal pressures and temperatures of formation for the deposit. Ninety-six reverse circulation holes and four diamond drill holes outline a bulk-minable mineral reserve of'4.1 million tons at 0.051 oz of Au per ton. An additional 2.7 million tons at 0.017 oz of Au per ton occurs in a low-grade zone above and subparallel to the main mineralized zone. Total calculated mineral inventory stands at 6.8 million tons at a grade of 0.038 oz of Au per ton, with a waste-to-ore ratio of 1.45:1. Column leach tests indicate that cyanide heap leach gold recoveries should exceed 70%.]

Jan 1, 1994

By W. Hopkin

INTRODUCTION Necessary science and technology are not well established for the design of responsible treatment and disposal of arsenical residues from refractory gold ores. Specifically, there is uncertainty about the precise composition and behaviour of oxidised arsenic species in these wastes, and an analysis of possible chemical behaviour suggests that potentially difficult control and disposal problems remain to be so 1 ved. In tackling these the undoubted toxicity of arsenic in excess and its bad press should not blind one to the realities of actual behaviour in the environment, and that some adverse behaviour inevitably has to be to 1 erated as a cost in return for the recovery of an operation's gold for society. The problem is to minimise any such cost. Each refractory ore will be different. Many will contain so little arsenic that it will cause only trivial problems in disposal, but there remain others with significant arsenic content, say in the order of 1% or more, to cause concern. As already implied, concern can be over emphasised but equally underestimated. The balance which must be struck in the design of responsible disposal in the widest sense requires objective and practical assessment of a complex range of chemical, biochemical, engineering and economic design factors, all of which should involve the process design team in constructive dialogue with the 1 oca l regulatory authorities. In British terms, the general objective should be the Best Practicable Environmental Option. Arsenic is ubiquitous in waters, rocks and soils and all living things. Its compounds undergo natural geochemical and bi ochemi cal cycles including powerful adsorption and precipitation processes which return to nature all leakage from the inevitable disturbance of mining and extraction operations. Within limits, all life is tolerant to its presence, and adaptation and detoxication processes occur to deal with local excesses.

Jan 1, 1991

By Satid Therdkiattikul, Donald A. Dahlstrom

Many gold ores are known to be refractory so that conventional cyanide processing will result in serious losses of gold and silver values. A significant percentage of these refractory ores contain aresenopyrite and the gold and silver is usually encapsulated within this matrix as well as the pyrite matrix. Normally the values can be recovered as flotation concentrates because of the sulfide compounds with recoveries of over 90% and at very good grades. In order to free the gold and silver, the sulfide matrix must be destroyed. This can presently be achieved by (1) pressure­temperature leaching (2) roasting and (3) bacterial leaching. All of them have advantages and disadvantages as well as an obviously increased cost for the each processing method. However, the increased gold and silver recovery should and must more than offset this cost. It is believed that gold recovery should be well over 90% from the concentrate so that overall recovery from the ore should be in the 90-95% region. All three processes must consider pollution control of both sulfur and arsenic to prevent contamination of air, water or land through improper waste disposal. This paper will concern only bacterial leaching and its necessary control of effluents and solids. Sulfur does not appear to be a problem as the element is oxidized to sulfate and is precipitated with an acceptable solubility. Other important improvements that must be controlled are arsenic and iron which are present as significant values (1-12 g/1 of arsenic, 20-40 g/1 of iron). In addition much smaller values of base metals, aluminum and magnesium can be present as well as other elements depending on the ore and these must be low enough after treatment to meet EPA regulations.

Jan 1, 1993

By Henry N. McCarl

From the earliest use of crushed and broken stone as landfill for eroded gullies in trails and roads, to the mixing of sand and gravel with cement to make concrete, men and women have found mineral aggregates to be useful materials in satisfying human wants. Crushed stone, sand and gravel, volcanic cinders, pumice, manufactured lightweight aggregates, slag, cinders, expanded perlite and vermiculite, and a great array of recycled materials such as grog (broken or crushed brick) and fly ash have served as mineral aggregates. These materials have provided bulk and strength in portland cement concrete, bituminous concrete, road base, fill, concrete block, and plaster and stucco finishes. Mineral aggregates provide many special characteristics such as compressive and tensile strength, surface textures, weight and density, thermal and acoustical insulation, abrasion resistance, and impermeability to various concrete products and mixes. While most materials used as mineral aggregates are consumed on the basis of their low unit values, the large volumes extracted in the United States each year make them one of the most valuable mineral resources in current production. In 1960, US production of construction aggregate was just over a billion tons per year (Meyer and Zelnick, 1991). Two decades of accelerated interstate highway, residential, and commercial construction brought the total annual production of mineral aggregates to roughly 1.4 Gtpy by 1970 and just over 1.6 Gtpy by the late 1970s. General economic and related construction slowdowns in the early 1980s saw annual US production of aggregates fall to below 1.1 Gt in 1982, and then rise to just over 1.7 Gtpy in 1990, a new peak in mineral aggregate production and use. Per capita annual consumption of mineral aggregates was approximately 5.68 t/capita in 1960, 6.89 t/capita in 1970, rose to a high of 7.25 t/capita in 1979, and has since settled to just under 7 tpy for every man, woman, and child counted in the United States Census of 1990. The value of these materials came to approximately $9 billion in 1990. During the period 1960-1990, roughly half of all mineral aggregates were used in the construction of residential, commercial, and industrial structures, and the remainder was consumed in highways, bridges, and other transportation facilities such as railway roadbeds, airports, and water-related projects (Meyer and Zelnick, 1991). During the period 1988-1992, public works (highways, waterway projects, road maintenance, etc.) represented closer to 60% of annual mineral aggregate demand with residential and non-residential construction approximately 40%. The growing market share of public works projects has been mostly due to the dramatic slowdown in residential, commercial, and industrial construction during the 1988-1992 period.

Jan 1, 1994

By J. A. Dos Santos

High angle conveyor (HAC®) systems in open pit mining offer many advantages over the traditional, truck only, haulage systems, including: (1) superior energy efficiency, (2) less dependency on petroleum products, (3) less sensitivity to inflation, (4) less labor, (5) less excavation for the amount of ore recovered, and (6) less ramp construction and maintenance costs. Previous studies had revealed the technical and economic feasibility of high angle conveyors. This paper presents the Sandwich Belt HAC® as the vital link to optimization of in-pit conveying systems. HAC development to date is discussed, beginning with studies in 1979, to the first large scale prototype in 1983, to the first commercial unit in 1984, into "the latest commercial installations. Discussion continues with the HAC systems to suit the various open pit mining operations and the corresponding characteristics of the system components, particularly the high angle conveyor. The range of HAC systems from the low capacity (up to 1,814 t/h), low lift (10.7 to 36.6M), highly mobile to the moderate capacity (1,814 to 3,629 t/h), moderate lift (36.6 to 76.2M), intermittently mobile to the high capacity (greater than 3,629 t/h), high lift (greater than 76.2M), semi-mobile, are discussed along with some possible combinations of these.

Jan 1, 1989

By Charles A. Ramsey, Kathleen S. Smith, Philip L. Hageman

We developed a statistically based strategy for sampling the surficial material of mine-waste dumps for use in screening and prioritizing historic dumps on abandoned mine lands. This sampling strategy entails the collection of a representative composite sample from individual dumps and allows for regional or watershed-based assessments. One 30-inment dump-composite sample collected using this sampling strategy contains as much information, relative to average value, as 30 individual grab samples at 1/30 of the analytical cost.

Jan 1, 2000

By Mark S. Stagg, Michael J. Nutting, Douglas R. Barlet

Single-row bench blasts at a reduced scale in the field were used to study the effects of delay time on fragmentation. Nine shots with a spacing of 21 in. and a burden of 15 in. (S/B = 1.4) and nine shots with a spacing of 30 in. and a burden of 15 in. (S/B = 2.0) were completely screened to determine particle size distributions. Delay intervals between holes ranged from 0 to 45 msec, corresponding to effective delay intervals ranging from 0 to 36 msec per foot of burden. Each shot was instrumented with strain and pressure gages to measure in situ blast dynamics and to evaluate interactions between blastholes. Finer overall fragmentation was produced for shots with delay intervals between 1 and 17 msec per foot of burden. Only for shots within this optimum delay range was it observed that the strains induced by stress waves constructively interacted with strains induced by gas pres- sure from an earlier detonated hole. Coarsest fragmentation resulted at delay intervals less than 1 msec/ft, where stress waves from each hole were observed to interact destructively, and at delay intervals greater than 24 msec/ft, where no inter- action between holes was observed, indicating a condition that can be considered as the firing of single hole shots independently.

Jan 1, 1986

By H. Köse, C. Tatar, C. Helvaci, F. Simsir

The Çayirhan coal basin is located 124 km (77 miles) west of Ankara,Turkey. With approximately 520 Mt (570 million st) of lignite reserves, the Neogene Çayirhan basin is one of the important coal districts of the central Anatolia region. The Çoraklar formation consists of crossbedded sandstone, channel fill conglomerates, mudstones, carbonaceous shales and volcanoclastic pebblestones, which are fluvial in origin. The lower and upper lignite seams occur, respectively, in the lower and the uppermost parts of the Çoraklar formation. The lignite deposits are in the Çayirhan and Koyunagili portions of the Beypazari Neogene basin. The Çayirhan coalfield consists of lower and upper lignite seams. The lower lignite seam does not outcrop at the surface. Its lateral discontinuity and relatively low sulfur content, plus sedimentologic and stratigraphic features of the associated sediments, indicate a fluvial origin. Thus, the lateral spreading and geometry of the lower lignite deposits were controlled by the channel, flood plain and well drained marshes of a fluvial system. The mining method employed at Çayirhan is the fullymechanized, retreat, working longwall with caving system. Due to ventilation, haulage and tectonic conditions, panel lengths from 0.6 to 1.7km (0.4 to 1.1 miles) are chosen, and 220-m (720ft) face lengths are used. The annual coal production is 2.2 Mt (2.4 million st).

Jan 1, 1994

By Cyrus W. Field

Chemical, mineralogical, and petrographic data have been obtained for suites of plutonic rock samples from the Guichon Creek Batholith of British Columbia and the Mineral-Iron Mountain-Sturgill Peak- Cuddy Mountain-Peck Mountain areas of western Idaho. Intrusions at these two geographically separate localities represent composite plutons ranging from early gabbro through quartz diorite to late grano- diorite and more silicic-alkalic phases. They are early Mesozoic (200 m.y.) in age, and contain porphyry-type mineralization in close spatial and temporal association with late-stage intrusions of porphyry, aplite, pegmatite, and (or) breccia. Country rocks are eugeo-synclinal assemblages of metasedimentary and metavolcanic greenstones of Permian, Triassic, and Jurassic age. The petrochemical evolution of these intrusions, and the effects of subsequent hydrothermal alteration and metallization, are illustrated by standard ternary diagrams of the major oxide and mineralogical data. Samples plotted on an AFM diagram suggest a calc-alkaline magma trend. However, their distributions on NKC and Qz-Kf-Pf diagram show well-defined trondhjemitic (Na20-rich) and tonalitic (quartz-rich) variants to this trend, which are similar in part to those reported for mineralized island arc plutons of the northern Caribbean and southwest Pacific, but which contrast markedly to those for continental plutons of the Bingham District, Boulder Batholith, districts of southern Arizona, and elsewhere.

Jan 1, 1975

By S. Hicks

This technical note describes an alternative procedure that can be used in linear programming calculations where the compliance limit of lbs SO2/million Btu for the blending coal is subject to change. Dr. Mateer in his technical note titled "Coal blending for sulfur compliance" (ME, September 1994, p. 1103) defined an "excess sulfur" parameter that provides a valid constraint equation for a linear program optimization. The coals used by Dr. Mateer in his examples are shown in Table 1. To create the parameters for the constraint equations to blend to an SO2 content of 1.2 lbs SO2/million Btu, the excess sulfur for Coal A was defined as Sulfur excess% = 0.5 - 0.6*13,000/10,000 = -0.28 This equation was used to create the constraint equation -0.28A + 0.44B + 1.36C[<] 0.0 Another way to create the parameters for the constraint equation is given below. In these equations: •W is the total weight; •A, B and C are the fractional proportions of each potential blend stocks; •Sa, Sb and Sc represent the percent sulfur content of the blend stocks; •Ha, Hb and Hc represent the calorific content of the blend stocks in Btu/lb; and •H and S represent calorific and sulfur contents of the composite blend. The weight of blend to produce one million Btu is [ ]

Jan 1, 1998

By Rodney H. Sergent

After being forgotten or ignored for nearly 100 years, the ability of bromine chemistry to extract gold from ore has been rediscovered. While bromine chemistry is similar to chlorine, there are some important differences. Bromine in its elemental form is a dense, fuming, red liquid. Elemental chlorine is a gas and requires a substantial amount of equipment to handle, store and deliver to solution. Bromine has an advantage in its ability to rapidly and completely oxidize heavy metals to water soluble halide salts. The first patent for use of bromine chemistry for gold extraction was issued in 1881 (1). By 1895, work done in Colorado concluded well roasted ores required approximately 1-1/2 lbs/ton of elemental bromine for superior extraction of metals (2)). At that time, bromine was deemed to be cost effective vs. chlorine even though it was more expensive on a pound basis than chlorine. The advent of cyanide technology virtually eliminated both chlorine and bromine from consideration for leaching of ores. Cyanide was so inexpensive that it made deposits requiring unique technologies unattractive for mining. Only since the deregulation of gold price has bromine/chlorine technology been considered for those ores not amenable to cyanide. While chlorine has found its way into the pretreatment steps, bromine is just beginning to realize its niche in the marketplace.

Jan 1, 1998

By P. Ollivier

A comparative study on potential refractory ore processing techniques was launched at the BRGM to enable exploitation of numerous small deposits of arsenopyrite and antimony sulphide ores.

Jan 1, 1998

By Luther J. Russell

Mineral development in South Dakota is aggressively encouraged by both state statute and county ordinances. Yet, an increasingly conservative philosophy is taking hold, resulting in expanded regulatory controls and concomitant increases in direct and indirect costs to permit and operate a mining operation. This paper examines the hidden costs associated with protracted permitting timeframes, land buy-outs, attorney's fees, etc., associated with the increasing emphasis for stiff controls on mineral development in South Dakota.

Jan 1, 1998

By Vytas Svalbonas

Sophisticated computer analysis methods such as finite elements, finite differences, and numerical integration have become tools in the mineral processing industry for design of large mill shells. While use of such methods allows the analyst-designer to take new freedoms with his designed structure, there are also pitfalls and dangers which must be recognized for the proper use of such sophisticated tools. One of the worst of these is treating any computer answers as "gospel truth". The present paper discusses the knowledge required by the analyst-designer for proper use of computer analysis, as well as explaining to the layman how to identify questionable areas in such structural analyses.

Jan 1, 1979

By R. H. Olson

Pigments are used in two segments of papermaking, as the major component in paper coating formulations and, in small percentages, as fillers in the paper web. Historically, the dominant use has been in coating, however, recent developments have fostered a multifaceted effort to substantially increase the pigment component in the paper furnish. Factors responsible for these developments together with their relationship to recent and forecast changes in papermaking technology are reviewed. Finally, some thoughts are presented on the probable timing and extent of these changes as well as their expected impact on the future demand for the common paper pigments.

Jan 1, 1985

By Vincent R. Bovino

Introduction Only a person fresh from an information vacuum could not know the status of this nation&apos;s economy. The inability of the US to compete in the world marketplace is manifested in many industries, from shoes to computer chips, from cars to copper. What industry will next fall prey to world competition? Perhaps the question should be, what insustry will next fall prey to declining productivity? American industry is being forced into an era where the focus must be on productivity, quality, cost control, and competition in the world marketplace. It is approaching a change that could rival the American Industrial Revolution. The future of the US is bright only if it can increase productivity by trying different and new ways to manage its businesses and people. Unlike the easy successes during the 60s and 70s, success in the decades ahead will depend on an ability to compete in the world market. Any operation can achieve unlimited productivity by developing the resources within its employees. As an industry, mining has harvested the earth&apos;s resources. As an industry in trouble, it now has to learn to tap human resources to release the energy and creativity in each person. Recently, I worked with a mining operation to reopen a mine and implement a different management approach. The extensive changes resulted in a 50% reduction in cash cost of producing silver (from $0.40 to $0.19/g or from $12.50 to $6 per oz). Thus, an unprofitable mine became an attractive business venture without technological changes or significant capital infusion. The operation is union-free, in spite of being in a highly unionized area. The mine implemented a real productivity improvement process called "gain-sharing." Employees now recognize their individual success is tied to the success of the business. They recognize the day of the free lunch is over for mining employees. This paper will explore three areas related to productivity: • some of the unspoken "whys" of the mining industry&apos;s paralysis and how management has caused unproductiveness in mining operations; • major components of a productivity improvement process; and • characteristics common to companies with successful productivity plans. Unspoken whys When things go wrong, everyone tends to point a finger at someone, anyone. This industry has blamed every conceivable person, corporate group, government agency, foreign government, and private group for its failures. The fault finger will not work anymore. Nobody is listening. It is a characteristic of winners in management to assume responsibility for less than acceptable performance. Excellent people tend to blame others less for problems. Everyone involved in mining should shoulder part of the blame for the industry&apos;s illness. Is the problem low prices or our high operating costs? To regain our status in the world marketplace and increase our productivity, we need to painfully and honestly talk about our mistakes.

Jan 3, 1986