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By Stephen Semeniuk

Working with junior mining companies as a valuer, my experience with regulators is that simplicity works and the sales comparison method is the way to go. What I like about Mr. Ellis?s approach (as in his Table 7) is the integration of commodity prices with the timing of acquisitions and the implication of price changes in gold on the adjustment of individual project valuations. However, the sample gold-related appraisals in Ellis?s paper are pre-2008 transactions and readers should be made aware of the current market for such projects. Before moving on, to be noted is that the effective date of the Las Brisas appraisal (Ellis?s sample Appraisal 3) is Feb. 10, 2006; given that the subject value of Transaction 9, Choco 10 in Ellis?s Table 6 was listed as $353 million for a 95% interest in the con-cession, I am uncomfortable accepting that the acquisition was suddenly worth the adjusted value of $1,205 million (as given in Ellis?s Table 7). In some cases, illiquidity discounts can apply, and commodity-wise, in the one-month period between the acquisition in January 2006 and the adjustment to the Las Brisas conditions as of February 10, 2006, the indicated gold price change was only 41 cents/g ($12.60/oz); i.e., from $17.50/g to $17.91/g ($544.40/oz to $557/oz).

Jan 1, 2012

By S. G. Malghan

W.H. Moyer In reference to the section “Excess pyrite, ” the author discusses difficulties in separating chalcopyrite from pyrite. A technique does exist, and has been used commercially, for the flotation, separation, and recovery, if desired, of both chalcopyrite and pyrite. The method was described in Mining Engineering, March 1966, “An Improved Flotation Method for Cornwall Sulfides. ” While it was not used on a porphyry copper ore, it was used on a chalcopyrite - pyrite-containing ore and might well have application. Both sulfides, chalcopyrite and pyrite, were floated simultaneously using xanthate as a collector and MIBC frother to produce a cleaned sulfide product. This concentrate was then thickened and conditioned at high solids with aeration and lime to a pH above 12.0 for a period of 1.5 to 2 hr. A supplemental suppressant, such as cyanide, could be used if necessary to sharpen the subsequent separation; however, it was not required at Cornwall. The conditioned sulfides were floated with two stages of cleaning to produce a highgrade chalcopyrite froth product and a pyrite sink product barren in copper. The separation at Cornwall gave a 28.6q. Cu grade chalcopyrite product and a pyrite analyzing 0.0670 Cu. A desliming step on the raw ore initially used at Cornwall for a period of time was later discontinued. Advantages to be found in using a strong sulfide collector, such as a xanthate, is the rapid flotation it induces and, generally, an ability to float at a coarser grind. A following regrind circuit could be used to give maximum liberation and produce a clean sulfide concentration. The “differential” separation is concerned primarily with grade and with assuring a barren pyrite for maximum recovery.

Jan 11, 1986

By Dan Plazak

Discussion by Dan Plazak The authors are to be congratulated for presenting a simple and practical tool for quick estimates of ground water flow rates. At the risk of detracting from the simplicity of the method, I would like to suggest a slight refinement to the equation as presented. As the authors noted, the Jacob-Lohman equation (1952) assumes confined-aquifer conditions. To use confined-aquifer equations in water-table aquifers, Jacob (1963) presented a correction to convert water-table aquifer drawdowns [Ahw] to equivalent confined aquifer drawdowns [Ah,: 2] [Ah Ah = Ah - w(1) Cw 2H where: Ah, = Drawdown in confined aquifer: Ahw = Drawdown in water-table aquifer and H = Original saturated thickness of aquifer] Substituting the above in equation ( I) of Hanna. Azrag and Atkinson [4nT Ah - 2H ) Qj2.25 T t(2) 1 nr2 S W] Since the above expression predicts lower pump rates, it causes the Jacob-Lohman equation to agree more closely with appropriate analytical and numerical models. Figure 1, adapted from Fig. 5 of Hanna. Azrag and Atkinson, illustrates the closer agreement of the modified equation to their numerical model, as reflected in Qm/QJL ratios closer to unity. The agreement improves particularly as drawdown [Ah] becomes large in relation to aquifer thickness H. An additional advantage of the modified equation is that Hanna, Azrag and Atkinson's correction factor C is much more stable as pit depth increases. Assuming, as did Hanna, Azrag and Atkinson, that their numerical model is an accurate predictor, it is clear that the modification incorporated in equation (2) improves the accuracy of preliminary estimates of ground water flow to a pit.[ ] References Jacob, C.E and Lohman. S. W.. 1952, "Nonsteady flow to a well of constant drawdown in an extensive aquifer." Transactions, American Geophysical Union. Vol, 33, No. 4. pp. 559-569. Jacob. C. E.. 1963, `Determining the permeability of water-table aquifers,' USGS Water-Supply Paper 1536-1, pp. 245-271

Jan 1, 1995

S.L. Barua and Y.C. Kim M. Armstrong's paper is very interesting and can be regarded as a pioneering work of applying geostatistics on washed coal. The density levels used in her study are uniformly spaced, which facilitated curve fitting through a proper experimental design. Most of the existing coal washing plants in the US, however, gather washability data at non-uniform intervals which, in turn, affects the goodness of curve fitting. We would like to emphasize the importance of having a proper experimental design, if a similar geostatistical application is to be conducted on washed coal. We, at the University of Arizona, tried the variogram on recovery (Kim, Barua, and Baafi, 1982, SME preprint 82-143) which showed a similar pattern as Fig. 4 of Armstrong's paper. In both cases, the first experimental variogram value is invariably and inexplicably higher than the second variogram value. Although Armstrong ignores the first point because of "small number of sample points," we did not do so because there were 17 sample pairs. As a consequence, we concluded that the obtained experimental variogram was unsatisfactory and that further research is warranted. In other words, this disagreement in variogram interpretation between M. Armstrong and the University of Arizona begs the question whether the experimental variogram of Fig. 4 truly reflects the spatial correlation in washed coal, assuming there is such a correlation. Finally, we wish that Armstrong could have done a similar study on clean sulfur, which is a vital interest to the coal mines in the US. Her implied remark on sulfur in the concluding section of the paper may turn out to be incorrect because our experimental variograms on clean sulfur at different density levels were even less satisfactory than those on recovery.

Jan 1, 1985

By J. W. Leonard, D. A. Newman

Discussion by N. Standish and P.N. Standish Leonard and Newman (1989) describe concepts and plans of their forthcoming project designed to increase the bulk density of broken coal at various stages of its production and handling. This is an excellent idea and something that will certainly be useful to the coal mining industry worldwide. The authors seek to achieve bulk density increase by "reducing oreliminating, through chemical additives and engineering design, particle voidage and enhance particle consolidation trickling between the point of breakage (at the face) and loading (in the shuttle car) to increase coal bulk density and productivity." In the paper, they present some supporting results of bulk density versus moisture, with or without oil addition, from the steel industry research. Unfortunately, the paper contains many misconceptions and half-truths that detract from an otherwise excellent objective of the project of increasing bulk density by reducing particle voidage-"the bottom line evaluator." Correct. Theory Definitions Coal bulk density is the ratio of the mass of a collection of discrete particles to the volume that it occupies (ASA, 1980). There can be no possible argument about this definition. Mass and volume are readily measured both in laboratory and large scale installations. So, to obtain the bulk density, one simply divides the measured mass by the measured volume. It is here that we come unstuck as everybody who has ever had to reconcile book values with the measured stock tonnages knows. Where does the problem originate? Almost invariably in a too casual a reading of the above definition and the words "a collection of discrete particles." The word collection means "a number of objects collected or gathered together, viewed as a whole" (The Oxford English Dictionary, 1989). If one knows the bulk density of a pile of coal viewed as a whole,

Jan 1, 1991

By David A. Zegeer

Since 1980, more than 40 miners have lost their lives by being trapped and suffocating in stock-piles, bins, silos, and similar facilities. Mine Safety and Health Administration (MSHA) is concerned about these accidents, most of which could be prevented if management and miners appreciated the hazard. In February 1986, the coal industry had a suffocation accident that drew national attention. A group of engineers were standing on a coal stockpile. It gave way under their feet. Five of the engineers were buried in the coal and suffocated. As with all fatal accidents, MSHA investigated the cause of this tragedy. I also asked my staff to look into the history of similar accidents. I found out it is rare to have multiple fatalities due to suffocation in material. But, over the years, there has been a long series of individual fatalities from this type of accident. Here are some examples. In Oct. 1985, a coal mine employee in Virginia entered a hopper to collect a coal sample. The feeders were started while he was standing in the hopper. He sank into the coal and suffocated. In Sept. 1985, an employee at a Texas sand company climbed onto a surge pile, to check a problem with sand flow. When other employees found him, the sand had swallowed him. Only his face and one hand were showing. Though the victim's face was clear of the sand, he suffocated before rescuers could free him. Many people do not realize you can suffocate in this kind of situation, though your head is free. Every time you exhale, the material has room to shift and press harder against your chest. Then you have less room to inhale. If this goes on long enough, it gets to the point where you cannot inhale at all. That is what happened to the victim in this accident. In May 1985, a plant operator at a Wisconsin sand and gravel operation purposely slid into the drawhole of a surge pile. He tried to break up large rocks that were clogging the drawpoint. Material piled on the side of the surge pile slid down and buried him. The operator was wearing a safety belt and line. But he was covered by 3 to 3.7 m (10 to 12 ft) of material and could not be pulled free. This incident illustrates another point. A safety line can help. But it is not enough. Workers should never put themselves in a position where loose material can fall on them from above. In Jan. 1985, a young Utah coal miner was waiting for a ride home from a co-worker whose job was at the tipple. A jam developed in the coal bin. The men could not clear the jam from the outside. So the young miner jumped into the bin. He started digging with a shovel at the toe of the coal pile. It caved in on him and he suffocated. The young miner had only three months of mining experience. He had not been trained for work around the coal bin. Since 1980, 16 coal miners and 26 nonmetal miners have lost their lives in loose-material suffocation accidents. Suffocation accidents occur in many areas around mines - bins, hoppers, stockpiles, surge piles, and silos, among others. One recent accident occurred at a surface coal mine in Illinois. The victim was walking over an area that had been drilled and blasted the week before. He fell into a void that had been crusted over with earth. He too suffocated. And many materials have been involved in suffocation accidentscoal, crushed stone, sand, salt, different kinds of ore, ammonium nitrate pellets used in blasting, and a powdery byproduct of iron ore milling called calcine. In addition to fatal accidents, there are near misses. We do not know how many. Most of them are not reported. A person who gets trapped in material, but is rescued, is usually unhurt. MSHA sometimes hears about these accidents by chance, or because they are reported in the local media. We do know that near misses are happening far more often

Jan 1, 1987

By K. J. Ma, L. Ma, J. Stankus

The paper’s intent is to evaluate pumpable foamed cementitious mine seal material by determining the shear strength of a generic foamed cement in the laboratory. The tests were conducted on three scales: small-size specimens, consisting of cylinders 4 in. in diameter or 3-in. prisms; double shear medium- size specimens, measuring 13 in. by 10 in.; and large-size specimens, measuring 6 ft by 4 ft by 4 ft. The small-specimen shear tests resulted in average peak and residual shear strengths of 128 psi and 56 psi, respectively. These values are fairly consistent with published shear strengths of foamed cementitious material. However, the other two types of shear tests did not achieve the shear strength of a typical pumpable foamed cementitous mine seal material. Based on the test results, the paper concluded that the structural analysis of cement foam seals should be based on their residual shear strengths, approximately 20 psi, determined from small laboratory-scale tests of the cement foam/concrete or rock interface, and not be based on their peak strengths.

Jan 1, 2015

By J. Grubb

"The undertakings described in the paper titled “Study on correlation of quantified coal rank with self-heating temperature” by Wang and Luo in the November 2015 issue of Mining Engineering (Wang and Luo, 2015) are mathematically sustainable and an acceptable extension of work performed by Smith and Lazzara (1987) for the U.S. Bureau of Mines. The paper is not an accurate explanation of the currently accepted leading practice for determining the propensity of a coal seam for spontaneous combustion. The paper could lead a practitioner to an oversimplified conclusion on how to determine the propensity of a coal seam to spontaneously combust.The spontaneous combustion of coal is a complex and not fully understood reaction. What is understood is that many more factors other than selfheating temperature (SHT) and coal rank are required to sufficiently evaluate the risk of spontaneous combustion in a particular coal seam. A superficial assessment can lead to an improper response to the risk. Some statements on the next steps through an appropriate literature search would have made the authors’ conclusions more credible.The SponCom 2.0 software, updated by the National Institute for Occupational Health and Safety (NIOSH) in 2011, provides a path for evaluating the susceptibility of a coal seam to spontaneous combustion. In addition to proximate and ultimate analyses to establish coal rank, and therefore an estimate of SHT, the program includes the history of heatings, friability, seam thickness, overburden thickness, rider seam presence, pyrite presence, face and butt cleat densities, geothermal source presence and other geological features in the assessment. Mining conditions including floor heave, rib sloughage and rate of advance are also assessed to conclude a propensity for spontaneous combustion."

Jun 1, 2016

By R. J. Pindred

In his paper, Davis presents an overview of risk. He also introduces the Capital Asset Processing Model (CAPM) as a foundation for selecting the appropriate discount rate for a mining project. While applying portfolio theory is more defensible than the ad hoc adjustment of discount rates, the CAPM is not a panacea.

Jan 1, 1995

By R. Bouzaiene

In the paper presented by F.W. Brackebusch on the “Basics of paste backfill systems,” paste fill was presented as a backfill process that offers many advantages over other alternatives such as hydraulic fill and rockfill. It was observed, however, that the mechanics of paste-fill transportation in pipelines is not fully understood and that the success of its implementation in practice depends on the precise control of the proportions of fines, aggregates and water in the mixture. The author emphasized the necessity for laboratory and full-scale testing to determine the suitability of a material to be used as paste fill. Although experimental testing is definitely an important operation in the design of a reliable paste fill, advances in the understanding of the fundamental laws that govern paste flow in pipelines are equally relevant. Ideally, the results of such efforts would provide predictive models for the rheology of the paste and the corresponding pumping-energy requirements. In an attempt to achieve this objective, an ongoing research program has been active at McGill University’s Department of Mining and Metallurgical Engineering.

Jan 1, 1995

By Pierre Mousset-Jones

I wish to submit this discussion item on an article published in the August issue of MINING ENGINEERING by Wala and Altman on the use of canonical diagrams for mine ventilation networks. The advent of computer graphics and plotting has enabled complex underground mine layouts to be displayed and plotted in a variety of three-dimensional formats that very closely represent the actual mine layout. This enables the ventilation engineer, who is familiar with the mine, to display the network pressures, quantities, resistances, and fan locations in a way that can be understood by everyone working at the mine. This is important if the ventilation engineer needs to discuss changes in ventilation with production personnel, or escapeway procedures with union personnel. While I can see some of the potential advantages in the canonical ventilation network diagram described by Dr. Wala, it appears to me that the fact that the diagram has no visual equivalence with the actual mine layout means that most mine personnel will find such a diagram almost incomprehensible. The authors have alluded to certain advantages with the canonical diagram. However, I find it difficult to see that these outweigh the fact that almost everything listed can be duplicated in isometric diagrams, and in a manner that is easily understandable to everyone in the mine from the miner to the mine manager. Perhaps the authors could show a clear example comparing an isometric with a canonical portrayal of a mine network showing the information that is immediately available from the canonical network that is not obvious in the isometric diagram. There is in this country a severe problem with the need for ventilation engineering and expertise to be used and accepted by mine management as a necessary part of mine planning. It seems to me that to introduce the requirement for further explanation so that management can understand a canonical network display will hinder more than help the recognition that ventilation engineering needs at so many mines in this country. However, I can see that the canonical approach does have its place in assisting the practicing mine ventilation engineer, wherever he or she exists in the organization chart for a typical mine. In this case, could the authors please indicate the availability of the computer program to compute and plot the canonical diagram for a mine ventilation network.

Jan 1, 1989

By K. Holmberg, D. Herriot

This paper by Mr Mclvor is yet another contribution to the eternal controversy surrounding the benefits, or otherwise, of high circulating loads in closed circuit grinding. Throughout the paper it is difficult to establish whether the author is referring to effects relating to the mill, classifier, or to the overall circuit. Table 1 is put forward as indicating the advantage of operating at high circulating loads. From the two (it would be interesting to see more) product sizing points provided in each case, the conclusions to be drawn are that cases one and two have very similar size analyses. The third case, however, indicates a much flatter curve, that is to say, higher fines production and a very wide spread of size ranges. If the objective of this circuit is to provide a closely sized product with minimum overground slimes to the downstream process, then it would appear to be the less "efficient" by far. No data is given for product pulp densities either. The term "efficiency" should never be used without a careful and clear definition. In this case, "efficiency" is used to express an apparent percentage reduction in operating work index (Bond) from the measured "standard" at 250% circulat¬ing load operation. The Bond calculation is itself based upon

Jan 1, 1989

By G. Bräuner

The methods of burst control treated in the authors' very informative paper have been applied in the Ruhr Coalfield, Germany, for more than 20 years and have proved to be highly effective. It might be interesting here to mention a few experiences: • The symptoms of critical stresses when drilling with small diameters ("test drilling") are essentially the same as described. However, we consider the situation safe if these symptoms occur beyond a certain distance ahead of the face. • That safety distance equals three times the seam thickness plus the daily advance. It is the total seam thickness that counts, not merely the mined portion. • Critical stresses within the safety distance are dissolved by so-called destressing measures: drilling with heavier equip¬ment, blasting, or water infusion. • To avoid bump initiation, the destressing operations should begin at the sides of the stressed zone rather than in the middle. • As to water injection, there are two different principles. If the situation is critical already, high pressures are applied for hydraulic fracturing. If the stresses still allow to drill deep holes, the fluid pressure is kept low and acts for a longer time. This is to make the coal break successively instead of violently - a long-term precaution. • When the stresses are very high, destressing is the most appropriate means. We mostly drill with 95 mm-diam (3.7 in. - diam). The techniques in detail are somewhat complex and sub¬ject to modifications. But the general idea is to accomplish burst prevention by controlling the potential burst situations.

Jan 1, 1989

By M. Albrecht

I have been following the discussions on coal degradation started by the article "Degradation process in coal slurry pipe-lines," by M.G. Ayat and B.C. Scott. I agree with Mr. Dasher's comments in January that the title of the article was perhaps misleading as the information appeared to pertain more to processing the coal than to transportation. Specifically, the authors' conclude that the cyclone had an important impact on the degradation. Dr. Ferri's comments on degradation in other physical processes was very correct, handling of coal in a turbulent manner has an impact on the size distribution of the coal. This is one reason that when I design a coal system (mine to rail car) I try to minimize the total number of transfer and processing steps consistent with achieving the desired product specifications. Another area, mentioned in passing in the article, the impact of natural causes and the mining step, is of particular interest to me. A mine in West Virginia had a variation of 50% in the -6.35 mm (-0.25 in.) raw coal depending upon the direction of mining. Since it was a single unit mine, this significantly impacted the final product. Two articles on mining's impact on size distribution are: • Stefanko, R., Chopra, 1K., and Ramani, R.V., 1974, "Mining influence on size consistency and washability characteristics of coal," SME Fall Meeting, September; and • Albrecht, M., and Olsen, D., 1980, "Mining methods impact on coal preparation plant design," 1st International Mine Planning and Development Symposium, September. Both articles presented information on variation of size distribution and cleaning plant performance on a coal depending on the mining method used. Both articles considered a single coal seam at one mining operation. Stefanko's article considered continuous and conventional mining of coal in Pennsylvania. Albrecht's article considered surface, underground continuous, and underground hydraulic mining of a coal in British Columbia. Both concluded that mining methods do impact coal characteristics. I agree with Mr. Ayat's comments in reply to Dr. Ferri's discussion, "... degradation is a multi-faceted phenomenon involving many natural and man-made factors." As efforts are being made to produce a higher quality (lower sulfur) coal, this topic becomes even more important

Jan 1, 1989

Dunn has written an excellent paper providing a wealth of advice on what is often the most costly consumable in the mineral processing plant. As he pointed out, grinding media selection and quality control can play a major role in successful plant operation. Owing to the economic importance of this topic, there are some important factors related to media sizing that I believe merit further comment. Dunn made reference to a slightly simplified version of the Azzaroni (1981) equation and pointed out its successful use for the Henderson SAG and Climax ball mills. Although both formulas are empirically based on successful plant experience, a notable discrepancy between them is that Bond has factored the work index by the ore specific gravity. This is invariably so, as the Bond work index is derived from the mass grindability (in grams per revolution) measured in a test mill with a fixed volume of solids load (Bond, 1961). A material having the same structural integrity but double the density of another will have double the grindability and just over half the work index. Thus, for the general case, specific gravity should be incorporated into any media sizing formulation that uses work index as the criterion for resistance to breakage.

Jan 1, 1990

By G. Le Bel

The paper proposes three methods to determine optimum lifetime of a mining project1. The conclusion is that the three methods arrive at the same optimum production rate (by coincidence). It is also stated that: "The stochastic and option pricing models, in addition to providing a more realistic optimum mine life than the traditional NPV (net present value) calculation, showed that the example project has the potential for substantially higher return. By explicitly considering a project's potential risks, the stochastic or option pricing techniques can more accurately describe its probable investment performance." This discussion is intended to demonstrate that: • It is no coincidence that the three methods arrive at the same optimum production rate. • The stochastic and option pricing models do not provide a more realistic optimum mine life than the traditional NPV calculation. • The potential for a substantially higher return is misrepresented. First, a review of investment decision theory in mining is needed. Net present value is the preferred performance measure because of the flexibility offered when comparing various alternatives. However, it presents the following disadvantages: • The discount rate selection is not straightforward. • The NPV obtained represents a deterministic characterization of the variables included in the cash flow calculations, i.e. a single-point estimate.

Jan 1, 1993

By P. Malozemoff

I have no objection to using mathematics in the value appraisal of properties. The factor that is not emphasized in the paper is that very few important properties have a lifetime that reflects the initial ore reserves of the property divided by the rate of production adopted. In the majority of cases, major properties mine out several times the initial reserves. How much more will be mined out is not known, and only exceptionally gifted geologists can predict that additional reserves will be found. There is the further factor of metals price, which, over the long term, is not predictable. Another factor vitiating making calculations on initially established ore reserves is the speculative nature of exploration success. In my judgment, the life of a property, in addition to an established ore reserve, should include ample time for exploration to find additional reserves. The likelihood of additional reserves is based on the geological promise of the general area where the mine is established. This is especially important because of various risk factors associated with small orebodies, the economic life of which is about five to seven years.

Jan 1, 1993

By S. Serata

Background Results of the above study in the August 1991 issue of Mining Engineering offer valuable lessons in the solution of cutter-roof problems. The original study plan was initiated by the discussion authors to solve the problems using the "stress control method" of mining (Serata 1976,1982; Serata, Carr and Martin, 1984; Serata and Gardner, 1986; Serata, Gardner and Preston, 1986; Serata, Gardner and Shrinivasan,1986; Serata and Kikuchi, 1986; Serata, Preston and Galagoda, 1987) However, the plan and the planner were changed to the arrangement reported in the paper. The change was considered reasonable at the time due to the mine engineers' uncertainties about the stress control method. Consequently, the basic principle of the study was shifted from the original stress control method to the method described in the paper, which will be called the "yield pillar method" for the purposes of this discussion. The paper convinces the reader that the yield pillar method fails to solve the cutter-roof problems. This doesn't mean that the stress control method also fails. Actually the contrary is true, as discussed below. Limitation of the yield pillar method The paper illustrates clearly how poorly the yield pillar method performs in solving the problem. The reason for this failure is the lack of the protective stress envelope needed to stabilize the cutter roof. Unfortunately, the protective envelope cannot be formed properly without utilizing the stress control method of mining. Changing the pillar size does not make much difference in the roof stability.

Jan 1, 1992

By E. J. Garrison

A.C. Noble presents a clear and concise summary of the factors pertinent to a competent reserve evaluation of a mineral occurrence. However, I believe that he confuses the difference between a resource and a reserve. By definition, a reserve estimate includes all known economic, legal, mining, metallurgical and environmental factors impacting the recovery of the product(s) found in the deposit. A resource estimate, however, is an estimate of what is there - what is available for exploitation and should not include economic and other considerations. Only after defining the resource, its extent, location, physical and chemical properties, can an intelligent selection process begin of the extraction and processing technologies most appropriate to the circumstances encountered. The early and almost certainly inappropriate application of these often cookbook choices has probably lead many companies to walk away from good properties. (How many mines have been found in others rejects?) An example of the process is the early selection of metallurgical samples, after encouraging values are found in a few holes or surface samples, in an attempt to determine the economics of the prospect. Unfortunately, the sampler does not know how representative of the occurrence the sample is because he does not know what the deposit looks like. As a consequence, the property becomes saddled with a bad metallurgical sample that is often inappropriately used in evaluating the deposit which it does not represent.

Jan 1, 1994

By D. G. Krige

I read the paper “The jackknife approach” with interest. There appears to be a widespread misunderstanding of the kriging variance as a measure of the error variance of the kriged grade estimate of an ore block. Virtually all mineral deposits display skew frequency distributions and are subject to the well-known proportional effect. Thus, it is usual to find in an ore body, when subdivided a really into subpopulations, that the variances of the grades (or the total sills of the variograms) for the subpopulations will be correlated with the corresponding mean grades (usually linearly with the squares of the mean grades). Thus, the variogram will be stationary.

Jan 1, 1996