Search Documents

By J. Leinart

As is readily apparent, the reliability of the early-stage costs estimated for any mining project are critical to the viability of that project. An estimate too high might leave a potentially profitable deposit undeveloped or worse yet, available to a competitor. One too low might waste investment dollars in a project not yet economically ripe. There are two aspects critical to any estimate: 1. The cost data used to represent the values of the wages, salaries, supplies, and equipment. 2. The design used to determine the engineering parameters upon which the above estimated values are based. Given that all of the information contained in Mining Cost Service is provided by suppliers, equipment manufacturers and mine operators, one wonders why there is any question as to the reliability of cost estimates developed using this information. Given that it is all actual (as opposed to statistically derived) data, where do the problems with estimate reliability lie and what is the impact of the data on that reliability? The following presents a series of analyses designed to assess the reliability of the cost estimating process. Both the actual cost data and the engineering parameters upon which they are based are examined in an attempt to quantify and understand the importance of each. While several sources of cost information are available (Means, Richardson, Primedia?s Equipment Watch, and individual suppliers and manufacturers), the following rely solely on data contained in InfoMine USA, Inc?s, Mining Cost Service and in its Sherpa Mine Cost Estimating Software.

Feb 27, 2013

By C. A. Rowland

The rising costs for and the possibility for limited availability of electrical energy are emphasizing the continual conflict faced in the selection of comminution circuits used in preparing ores for concentration, namely; capital cost vs. operating cost. Frequently, the circuit with the lowest capital cost is not the most efficient in the use of electrical power and wear resistant materials. The simplest circuit in terms of flow, pieces of equipment and capital cost is the single stage autogenous or partial autogenous circuit used to grinding primary crusher discharge or run of mine ore to the desired product size. (Fig. 1) From this circuit, complexity increases to multi-stage crushing circuits preparing feed for multi-stage grinding circuits. Fig. 2 showing a three stage rod mill ball mill arrangement is an example of a complex grinding circuit. Stage concentration further complicates the flow. Multi-stage circuits being the most expensive to install, often give the greatest operating flexibility and offer the best possibilities for the most efficient use of available power and wear resistant materials. Within available grinding circuits, other comparisons such as high speed vs. low speed mills, grate discharge vs. overflow discharge, metal liners vs. rubber liners, high charge vs. low charge, etc. all add to the conflict

Jan 1, 1977

Recent advances in GPS, machine guidance technology and telemetry allow an unprecedented level of control of surface mining operations. GPS technology can now locate the bucket of a shovel or the blade of a dozer within 50mm (2 inches).Some of the outcomes possible are: precise selective mining – even when the ore is visually indistinguishable from waste: tracking grade and tonnage of every ore parcel through the mine: real time productivity monitoring: detailed post accident safety analysis: tracking the movements of every piece of equipment. This paper will consider the contribution that GPS based machine control systems can make to ensuring that surface mining projects deliver the outcomes promised in the feasibility studies. Case studies of installations are presented.

Jan 1, 2004

By B. J. Arnold

Prior laboratory work has demonstrated -that much of the fine ash-forming minerals entering a flotation froth are there as a result of water carry-over from the bulk contents of the flotation cell (Arnold and Aplan, 1986a). Using a bank of four 300 cu ft WEMCO flotation cells cleaning a well-liberated Lower Kittanning seam coal, test results show that the sulfur and ash content of the froth is directly related to the water content of the froth (or inversely related to the percent of solids in the froth). Reagent changes which result in a watery froth produce a higher ash content in the froth. A direct comparison of laboratory and plant results show that the standard laboratory flotation test produces a product higher in sulfur and ash than that produced in the plant. Caution should be exercised when extrapolating laboratory test results to the plant situation. A method of estimating coal flotation product quality in the plant based on laboratory tests is given.

Jan 1, 1990

By G. T. Lineberry, L. Adler

Due to safety-imposed constraints on clearances for mobile bulk handling equipment, it is necessary to optimally select a unit based on speed and size. The unit cannot be both large and fast. Since output is proportional to both size (or capacity) and speed, the combination must be carefully examined to maximize output. A simple graphical procedure and a calculus-based solution are presented to accomplish this. While the procedure applies to all mobile bulk handlers, as an example of the solution technique, attention is focused on shuttle cars. Background Major variables To fully define a bulk handling problem, the conditions, equipment, and activities must be characterized. The conditions and activities, when lumped together, are termed "encumbered space," since they "act to impede motion, as with a burden" (Lineberry, 1979). The insertion of mobile bulk handling equipment into what is already a difficult set of conditions and activities compounds the state of encumbrance, as limited space is competed for by many normal and life support systems. Long (1972) recognized the effects of design criteria, including road width, on haulage vehicle speed and, therefore, on production. Velocity -clearance curves Velocity-clearance curves are either available or can be constructed in the field. A preliminary set of curves for average conditions and activities is shown in Fig. 1, where clearances are only for the side of the vehicle (Lineberry and Adler, 1984). These curves set the safe speed limits for a vehicle operating at given clearances, under given conditions and activities. When feasible, velocity-clearance curves should be constructed through field measurement and observation. Curve construction should also consider haulway roughness, distances, and operator skill, as well as loading and dumping characteristics. It should be noted that haulage capacity determination is a total systems problem, not one just dependent on velocity and clearances. Production requirements and company capabilities also influence capacity determination.

Jan 10, 1985

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

By T. V. Ramey

Heap leaching in certain parts of the world, such as Chile's Atacama Desert, brings unique design problems in the form of possible large differential settlements of the foundation soils due to the dissolution of salts in northern Chile the soils often contain very high concentrations of soluble salts (nitrates, potassium, magnesium, and common sodium). When a leak in the liner occurs, the escaping solution dissolves these salts and triggers differential settlement. If the settlement is great enough, the geo-membrane tears, the leakage increases, triggering more differential settlement. The nitrate plants have experienced many occurrences of this phenomenon where a brine pond will undergo a sudden and complete loss of the impounded liquors. The authors present a simplified technique to quantify the differential settlement resulting from dissolution and provide a method which can be used to select a geo-membrane with suitable flexibility under axi-symmetric (multi-axial) load conditions to tolerate the predicted settlement. (This paper was previously published for Randol Gold Forum 1996 [Ref. 6]).

Jan 1, 1997

By A. Fiszer

One of the major problems with deep Polish hard coal mines are deformations of close-to-wall galleries since they are necessary and must he kept open for emergency situations and are largely responsible for work disturbances. In order to ensure safe and effective working operations, it is necessary to often rebuild galleries which, apart from the inevitable stoppage of operations, is greatly disadvantageous for the operation of a mine. The largest problems with maintaining the galleries open occur in Lower Silesian Coal Field mines where sometimes it is necessary to reconstruct the galleries several times [(Fig. 1)]. In light of what has been said above, the aim of this article is to present the reader with methods and ways of determining deformations occurring in close-to-wall galleries. Squeezing or convergence of headings is only a portion of the entire process of ground movement caused by ground disturbance due to workings. The term "squeezing of headings" is to be understood as a shortening of the distance between a point set on the roof of a heading and the corresponding one on the floor. It is very important to know the rules governing the process of squeezing in a particular mining system since it affords possibilities to define changes of the process and its parameters at any moment for a given system. This is of utmost importance because convergence of a heading is one of the most significant manifestations of ground pressure which has already been proved by studies performed by Everling (1). On the basis of those studies it has been observed that there exists an almost linear relation between the rate of convergence and loading [(Fig. 2)]. This, in turn, means that the deformations of the already exploited portion of heading are proportional to the pressure changes.

Jan 1, 1982

By C. L. Sainsbury

For 10 years, the author has been engaged in a program of geologic mapping of the Seward Peninsula, Alaska, for the U.S. Geological Survey. Initially, two 1:63,360-scale quadrangles were napped by traditional methods of traverses by foot and by tracked vehicle, supplemented by reconnaissance from a light aircraft owned and piloted by the author. Five summers were required to map those two quadrangles, which lie north of the tree lice in typical Arctic climate. In 1967, the project was reoriented toward the reconnaissance capping of four 1:250,000-scale quadrangles, with helicopter support to be used for cart of each of four summers amounting to about 45 hours time per support year. The light aircraft was use-,' to support the field operations. Over the year,, the aircraft was used in progressively more geologic tasks, culminating in 1970 with a successful field test of rudimentary devices made to collect soil, rock, and sediment samples while the aircraft remained airborne. Additionally, an attempt was made j- 1970 to prepare reconnaissance geologic maps entirely by the use of the light aircraft combined with foot traverses [fr??], the landed aircraft. Several such maps, prepared at the rate of about one 1:63,360-scale quadrangle per day, were checked from a helicopter by an impartial observer end were found to be sufficiently accurate for reconnaissance mapping when reduced to 1:250.000 scale.

Jan 1, 1971

By W. E. L. Minter

The most important witwatersrand gold placers occur on extensive unconformity paleosurfaces that were produced during protracted periods of fluvial degradation. The erosional etches on what was generally unlithified substrate indicate that low sinuosity erosion patterns prevailed. The transport systems were dominantly braided fluvial with a minor debris-flow component. The placer sediments were non-cohesive sands and gravels which were evidently transported by fluctuating, yet perennial, discharges across extensive, low relief, bra1dplains, not far from littoral environments. Progress made in understanding heavy-mineral transport processes now emphasizes the importance that selective entrainment probably had in producing concentrations. Cognizance that these characteristics are not unique to the Witwatersrand should be taken into account when designing exploration models.

Jan 1, 1988

By James M. Barker

The occurrence of borates and zeolites at Mesa del Alamo is well documented by McAnulty and Hoffer (1) and Lefond and Barker (2). Further investigation indicates the presence of priceite and meyerhofferite in addition to the howlite (1) and colemanite (2). Analcime occurs in addition to the clinoptilolite and phillipsite (2). The Mesa del Alamo area is about 200 !a south of Tucson, Arizona, and 13 km southeast of Magdalena, Sonora, Mexico, just south of the unpaved road to Cucurpe (Figure 1). The geology and mineralogy of northcentral Sonora is known only on a regional basis except for small areas mapped in detail during copper exploration. The study area is within the Basin and Range province of North America from which borates and zeolites are now produced in California. The borates and zeolites at Mesa del Alamo occur in geological environments similar to those of the Mojave and Death Valley regions of California (3, 4). Thus, a detailed exploration pro- gram in the Magdalena area could result in the discovery of economic borate and/or zeolite deposits.

Jan 1, 1979

By Lynn Huntsman

INTRODUCTION The Sunnyside mines of Kaiser Steel Corp. are lo¬cated on the west flank of the Book Cliff Range, a tongue of the Wasatch mountain range in Utah. The mines are 193.12 air km (120 air miles) southeast of Salt Lake City and 40.23 km (25 miles) east of the town of Price. The area has a semiarid climate with an an¬nual rainfall of 304.8 to 381 mm (12 to 15 in.) and a temperature range of -23.3 to 32.2°C (-10 to 90°F). The base of the Book Cliff Range is approximately 1981 m (6500 ft) in elevation with some mountain peaks in excess of 3048 m (10,000 ft) (Fig. 1). The Sunnyside mines were opened in the late 1890s and have been operated continuously since that time. Kaiser Steel leased the No. 2 mine in 1942 and pur¬chased the entire property in 1950. There are three mines adjacent to one another, each with its separate access, ventilation, and haulage. The three mines en¬compass an area extending 12.87 km (8 miles) along the outcrop and dipping northeasterly 2.41 to 3.22 km (11/z to 2 miles) into the Book Cliffs. During 1975, only the No. I and No. 3 mines were operating to sup¬ply 908 184.7 t/a (1,000,000+ stpy) of clean coal to the Kaiser Steel mill at Fontana, CA. Sunnyside coal is blended with coal from the Kaiser Steel mine near Raton, NM, and with purchased coal from noncompany sources to form the mix for the coke ovens. At that time, Sunnyside operated three longwall units and five to seven continuous miner units. Typical production for that time came from one miner unit pillaring an area not conducive to longwall mining; one or two miner units on construction or modernization projects, and the remaining units used to develop future longwall panels. Longwall mining has been a part of the Sunnyside scene since 1961 when the first panel was set up in the lower Sunnyside seam of the No. 3 mine. Through 1975, 27 longwall panels were mined (Fig. 2). Of the more than 47 173 604 t (52,000,000 st) of clean coal that were taken from the property through 1975, 6 350 292.9 t (7,000,000 st) were mined by longwall methods. Coal was brought to the surface in 4.5-t (5-st) mine cars and crushed to -152.4 mm (-6 in.) before entering the preparation plant. Coarse coal is cleaned by Baum jigs and fine coal (-28 mesh) is cleaned in flotation cells. The coal is reduced to 50.8 x 0 mm (2 x 0 in.) and belted to an aboveground storage pile with a ca¬pacity of approximately 113 398.1 t (125,000 st). The coal is gravity fed through two 1.52-m (5-ft) diam culverts directly into railroad cars, which pass beneath the storage pile through an arched excavation. Eighty-four 90.7-t (100-st) cars make up a unit train which makes the round trip to the steel mill in four days. Additional trains are scheduled in as demands dictate. Loading rates in excess of 13 607.8 t/h (15,000 stph) have been attained. Geology The coal seams of this area are of Cretaceous age and are classed as a medium volatile bituminous coal. coal, with a sulfur content less than 1 % and 6 to (20-mile) zone of the Book Cliffs, and this area is owned or under lease by either Kaiser Steel or US Steel. Coal in adjacent areas is noncoking but an excellent steam coal, with a sulfur content of less than 1 % and 6 to 7% ash. Sunnyside coal contains 40% volatile, 53% carbon, less than 1 % sulfur, and 51/2 to 61/2 % ash after washing. There are two minable seams throughout most of the Sunnyside property known locally as the upper and lower Sunnyside. Separation between the seams varies from less than 300 mm (1 ft) near the southern boundary adjacent to the US Steel Columbia mine, to over 12.19 m (40 ft) in areas of the No. 3 mine in the center of the Sunnyside property. The upper seam varies in height from 1.2 to 1.98 m (4 to 61/2 ft) while the lower seam varies in height from 1.52 to 4.28 m (5 to 14 ft) in areas of No. I mine where the two seams apparently come together. The majority of early work was con¬centrated in the upper seam which has a stronger roof and a lower sulfur content. When Kaiser Steel began mining the No. 3 mine in the early 1950s, it mined in the lower seam beneath the old workings with conven¬tional and continuous mining equipment. Numerous bounces and pressure points created by unextracted pillars and barriers in the upper seam made this a low¬production, high-cost mine and brought about the ad¬vent of the first longwall system. The first unit, while small by today's standards, successfully mined panels that had been all but given up and saved the mine from closure. There were areas that could not even be mined by longwall because of the erratic extraction in the upper seam. The system was successful enough to allow operations to gain access to and develop virgin areas of the upper seam. Development of longwall mining in the No. 3 mine has been confined to the upper seam since late 1964. With the almost complete extraction afforded by the longwall system, it is felt the lower seam can be successfully worked in future years (Fig. 3). In most of No. 1 mine, only the lower seam (or lower seam and upper seam combined) is minable, except in the northwest portion where a rock split has again separated the seam into two 1.83 to 2.13-m (6 to 7-ft) thick beds. In this area, the upper seam or upper split is being mined. The separation between seams at the present workings in this area varies from 0.609 to 6.1 m (2 to 20 ft). It is hoped that most of the lower split can be mined using shield supports and/or what¬ever new technology is developed in the coming years. The coal seams in this area dip to the northeast at an average grade of 10%. The dip near the outcrop reaches 22% and gradually decreases downdip. Over¬burden ranges from a few meters at the outcrop to 762 m (2500 ft) over some of the deeper workings. In the next several years, some overburden will be in excess of 914 m (3000 ft) as the lower mine workings extend ever deeper under the Book Cliffs. To cope with the

Jan 1, 1982

By Robert Stefanko

The most perplexing problem that has confronted mining engineers since the dawn of mining has been strata control. For the purposes of this paper, strata control will be defined as control of ground in the immediate area surrounding the, mine opening. To prevent movement of the surrounding strata into underground mine openings, miners have traditionally relied upon some type of artificial supports placed within the openings and located in a manner determined by trial and error from previous experience. Without the means to measure changes of stress in the supports and within the mine structure itself, an evaluation of the efficiency of the supports and mining system was impossible. Strata control thus remained an art with very little application of science. During recent years a revolution has occurred in mining. The modern mining engineer is contented no longer-with trial and error methods but is utilizing all the technological information at his command to obtain a solution to this age old problem. However, serious problems have arisen due to the lack of certain basic knowledge in three important areas: 1, The original stress distribution within the earth strata prior to mining. 2, Redistribution and concentration of the original stresses with mining activity. 3. The properties of rock that are significant in failure and its prevention.

Jan 1, 1960

By M. J. Blattman

Upon completion of an underground mine design, the planning engineer is faced with a surprisingly complex problem, “What development and mining sequence should be implemented so as to maximize the return on investment?” The problem relates to the scheduling of development headings to access stope production areas in such a way that the most money is made as quickly as possible for as long as possible. The total dollar value of the stopes, the average grade, the total tonnage, nor even the profit margin do not accurately give an indication of the “proper” mining sequence on their own. This paper will outline a method being developed by the author that can create a sequence of events that maximizes the net present value (NPV) of a given development and production design using information already being generated during the design process.

Jan 1, 2003

By W. A. Haley

Wheel loaders were first introduced shortly after World War II as small machines designed for handling loose materials in various industrial and processing applications. It was not until the early 1950?s, however, that machine size permitted widespread loading of loose or stockpiled material into on-highway type trucks. An ever increasing number of machines continue to be used for this purpose today. In fact, it is by far the major use for wheel loaders in mining and quarrying. Many of these loaders were working at mines or quarries handling in-process or finished material and some were doing shovel clean-up. It was quite natural then, when the primary pit loading shovel was down, to attempt to maintain some production by replacing it temporarily with the wheel loaders. Most of these attempts were not very successful, however, until the really large wheel loaders, 4- to 6-yd. size were introduced in the early 1960's. Since that time these machines have proven they can be productive and economical primary pit loading tools, a job that was once the exclusive domain of the power shovel. To be competitive with the shovel however, conditions must exist or be made to exist, that will be favorable for the wheel loader.

Jan 1, 1967

By E. Costa Silva

Among the methods of mineral processing, gravity concentration methods are, probably, those which have lower costs, higher capacities and lower operational complexity. In order to be really effective, these methods require a feed size distribution in which the particles are liberated. Further, there must be a significant difference (usually higher than 1-1.5) among the real densities of the particles. However, in most cases, the traditional gravity concentrators (jigs, shaking tables, spirals and dense medium hydrocyclones) lose efficiency (sharpness of cut and recovery) when the mean diameter of particles are lower than 0.05mm and the difference among densities are smaller than 1. Most gravity concentrators developed in the last 20 years are based on the use of centrifugal forces that create large "apparent gravitational gradient" seeking to overcome the limitations of particles size distribution. This paper presents the experience of the CVRD Research Center in evaluation and development of new process based on using this new generation of gravity concentrators.

Jan 1, 1999

By B. Seymour

The National Institute for Occupational Safety and Health (NIOSH) is conducting research to develop safe practices for the use of shotcrete as ground support in underground mines, particularly mines operating in weak host rock. As part of this research, tests were conducted with a commercial poly-fiber reinforced shotcrete mix to develop a practical means of measuring shotcrete adhesion strength. Full-scale tests were conducted in a test frame equipped with concrete panels having three distinct surface roughness profiles. Adhesion test fixtures either embedded or epoxied in the shotcrete were overcored into the underlying concrete, and direct tensile tests were conducted to determine the bond strength of the shotcrete cores to the substrate after a selected shotcrete curing interval (1, 3, 7, 14, 28, or 90 days). Measured adhesion strengths typically ranged from 0.5 to 2.0 MPa depending on the curing age of the shotcrete samples. This paper presents the results of these tests, describes the portable test system components, and addresses some of the design issues encountered in the development of a rugged and reliable method for determining shotcrete adhesion strength in underground mines.

Jan 1, 2010

By Dennis S. Kostick

Soda ash is produced in over 39 countries by the Solvay ammonia-soda process or by mining natural sodium carbonate deposits. The total production of soda ash in 1980 is estimated at slightly over 29 million metric tons. Synthetic production accounted for over 75% of this amount with the balance of production originating from natural sources (either trona or sodium carbonate-rich brines) found primarily in the United States, Mexico, Kenya, and Chad.

Jan 1, 1981

By Alfred L. Anderson

Thorite and rare earth deposits in the Lemhi Pass area of southeastern Lemhi County, Idaho, are contained along complex shear and fracture zones and in reopened gold-quartz veins and copper lodes in quartzitic and phyllitic rocks of the Pre-cambrian belt series. Some of the shear and fracture zones are several tons of feet across and comprise broad zones of irregularly mineralized rock. These lodes contain a notable concentration of thorium and rare earth elements along with considerable amounts of barium, .potassium, calcium, iron, aluminum, phosphorus, sulfur, and silicon. The principal minerals are thorite, locally allanite, monazite, xenotime (?), apatite, specular hematite, barite, potash feldspar, occasionally calcite and quartz. Exept for some of 'the specular hematite, barite, reddish feldspar, calcite and quartz, the minerals are not distinguishable without the microscope. The thorite and locally allanite were deposited first, then the monazite and other phosphates, which were in turn followed successively by barite, potash, feldspar, calcite, and quartz. In exceptional cases, deposition of thorite and specular hematite was repeated after the feldspar. Some of the deposits contain several percent thoria, but the average os generally under one percent. The rare earth content is comparable with that of the thoria. The deposits appear to be of hydrothermal origin and are probably related genetically to late Cretaceous or early Tertiary igneous activity, manifested locally by a few dioritie and lamprophyric dikes.

Jan 1, 1959

By W. M. Colwell

When it comes to the analysis of gold bearing ore, there are literally dozens of variables that contribute to the final "true value" of a given sample. Initial sample selection, splitting size and method, inherent nugget gold, pulverization method, and analytical method variations are but a few. However, it seems that the inherent nugget gold particle is targeted as the primary reason for intra and inter-laboratory variations in assay. The following study will show that this is not always the case. A series of tests were undertaken for the Aurora Partnership-Aurora Project to solve inter-laboratory variations in the analyses of blasthole cuttings. Initial concern was centered on the probability that free gold occurrence was resulting in erratic results for the mine's cyanide shake assays and outside laboratory fire assays. Comparison of fire assay and cyanide shake assay was sporadically consistent with previously demonstrated extraction percentage for the cyanide method. The goal was to delineate the factors that ensure consistency in analyses; implement, where needed, Good Laboratory Practices; and reestablish acceptable analytical method correlation.

Jan 1, 1993