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By M. R. Yalamanchili, J. D. Miller

Discussion by M.R. Yalamanchili and J.D. Miller The authors, M. S. Celik et al., should be recognized for their efforts to describe the flotation behavior of boron minerals. In the case of borax and other soluble salt minerals, analysis of the flotation chemistry has been difficult because of the high ionic strengths associated with these soluble salt systems. However, considerable progress has been made in this area, and recently a surface charge/collector colloid adsorption model was proposed by Miller and his coworkers to explain the collector adsorption phenomena observed in soluble salt flotation systems (Milleret al, 1992; Yalamanchili et al., 1993; Miller and Yalamanchili, 1994; Yalamanchili and Miller, 1994a: Yalamanchili and Miller, 1994b). In this work, the sign of the surface charge of alkali halides in their saturated brines was established on the basis of nonequilibrium electrophoretic mobility measurements by laser-Doppler electrophoresis (Miller et al., 1992). Generally, these results are what would be expected from the simplified lattice-ionhydration theory. This electrokinetic information coupled with the stability and prevalence of collector colloids in such soluble salt flotation systems indicates that the selective flotation of alkali halides is due to the adsorption of oppositely charged collector colloids by heterocoagulation. Experimental flotation/bubble attachment results for 21 different alkali halides (Yalamanchili et al., 1993; Yalamanchili and Miller, I994b) confirmed that the flotation response of soluble salt minerals with weak electrolyte collectors can best be explained by the adsorption of oppositely charged collector colloids rather than by the adsorption collector ions and/or neutral molecular dipoles as originally suggested by many researchers (Fuerstenau and Fuerstenau, 1957; Schubert, 1967; Roman et al., 1968). In addition, the flotation of certain alkali oxyanions (Pizarro et al., 1993) and double salts such as schoenite and kainite can be explained by the same collector colloid adsorption mechanism (Miller and Yalamanchili, 1994). The borax flotation results reported by Celik et al. need to be examined in terms of the above mentioned surface charge/ collector colloid adsorption model. Unfortunately, the authors seem to be unaware of this recent work that nicely describes soluble salt flotation with weak electrolyte type collectors such as amines and carboxylates. In view of our past work, the flotation characteristics of borax were of particular interest, and, in this regard, the results of dodecyl amine flotation of borax reported by Celik et al. have been examined in further detail in the light of experimental results from our laboratory. In our research, a vacuum flotation technique was used to study the flotation response of borax (Na2B407.10H20), which has a solubility of 39 g/L at 25 °C) with dodecyl amine hydrochloride as collector. These chemicals were purchased from Eastman Kodak and used as received. Saturated solutions of borax at desired pH values were prepared by continuously stirring the salt solutions over a period of about 10 hrs. It should be mentioned that the conditioning time to achieve equilibrium is an important variable and can significantly change the flotation response of some soluble salts (Yalamanchili et al., 1993). Collector was added to the saturated borax solutions containing about one gram of 100x 150 mesh borax particles, and conditioning was done for about 20 minutes prior to flotation. The borax flotation recoveries from saturated brine are presented in Fig. 1 as a function of collector addition at the natural pH of 9.3, as reported both by Celik et al. and as measured in our laboratory. In addition, the region of precipitation for the dodecyl amine hydroborate is included in Fig. 1. It can be seen in Fig. 1 that the flotation response curves are separated by about one order of magnitude in R12NH3CI collector addition. The flotation results of Celik et al. show that the maximum borax recoveries can be obtained below the solubility limit of the dodecyl amine hydroborate collector. However, in our experiments borax flotation seems to occur only after the precipitation of the dodecyl amine hydroborate collector as might be expected from the collector colloid adsorption model (Yalamanchili et al., 1993) if borax were negatively charged. Further analysis by nonequilibrium and equilibrium electrophoretic mobility measurements for borax indicates that borax is negatively charged at the natural pH of 9.3, as discussed below. The reliability of the nonequilibrium electrophoretic measurements has been demonstrated previously for alkali halides and alkali oxyanions (Miller et al., 1992; Miller and Yalamanchili, 1994). The equilibrium and nonequlibrium electrophoretic measurements for borax were found to be consistent and are presented in Table 1. These results provide clear evidence that borax carries a negative surface charge in its saturated brine (pH 9.3), and the sign of the surface charge of borax reverses and becomes positive if the pH is reduced to 8.6. The equilibrium between borax and its saturated brine can be described by the following reaction: [2Na2B407.1OH2O-4Na++B407=+HB4O7 +OH+19H20] It appears that the oxyanions of the borax lattice provide

Jan 1, 1995

By E. J. Garrison

Discussion 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. Reply by A.C. Noble E.J. Garrison asserts that I confuse the difference between a resource and an ore reserve. In fact, this paper was motivated by the observation that resources and reserves are almost universally confused throughout the industry. While Garrison is concerned that too-early consideration of engineering, metallurgical and economic factors will result in dropping properties too soon, some issues that must be considered include: •The purpose of exploration is to discover ore reserves, not to discover resources. •Ore reserves are always smaller than resources. • Decisions on exploration expenditures must be made based on the expectation of the quantity of ore reserves that will ultimately be defined. Noble's Fig. 2 can be used to give guidance as to when it is appropriate to begin reserve calculations. When exploration enters the "flat" portion of the curve, representative samples can be selected for metallurgical studies and appropriate mining methods chosen. It is at this stage that we know enough about the deposit to make intelligent choices of representative samples and recognize constraints on mining caused by the deposits shape, environment and physical properties. Once the appropriate technologies have been chosen, then their cost can be estimated with reasonable accuracy. Early entry of economic and process factors gives the false appearance of reserve status to the resource estimation. It would be more informative and less subject to abuse and misunderstanding if resource calculations were made at a number of cutoff grades. The cutoff grades should be chosen to cover the range of grades found in practice in existing operations of similar type. This would emphasize the open nature of resource estimates and encourage creative evaluation of mining and process technologies. It does not preclude the use of economic, environmental and recovery technology factors in deciding the advisability of continuing exploration expenditures, but would highlight their speculative nature. It is thus hoped that decisions on the future of the deposit would not be made on the presumed suitability of a given technology but on the actually known reality. I agree that resource estimates should be reported as a range of cutoff grades. Since reporting at a cutoff implies application of underlying economics (even by analogy to similar properties), however, it should be clearly stated that resources do not consider economic, mining or cost factors, and that ore reserves will be substantially less. The dilemma of the explorationist is that until the continuity and limits of ore-grade mineralization have been established, it is difficult to make quantitative estimates of either resources or ore reserves. Further industry discussion is needed regarding standards for disclosing early exploration information in such a way as to fairly represent the potential of an exploration property.

Jan 1, 1995

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The authors claim that their proposed jack-knife method, which is a rekriging of each block by eliminating, in turn, one sample from m the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" .of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper con¬sist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data value-based measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text.

Jan 1, 1997

By D. G. Krige

Discussion 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 station. When this feature is evident, the relative variograms for the subpopulations (i.e., with variogram values divided by the relevant mean squared) will be similar, or they could even be accepted at an averaged model for the whole ore body with stationarity. Using this approach, the kriging variance will reflect the relative error variance that is to be multiplied by the kriged block grade estimate squared to yield the usual kriging variance for purposes of confidential limits. The objection that the kriging variance does not take into account the local grades used for a block valuation then falls away. In the example given by the authors, the global mean grade is not recorded. But the global variance, taken at the value of the total sill, is 0.005, whereas the variance of the 11 values used in the example is 0.000152. The two variances differ by a factor of nearly six. If the relative variogram approach is applicable in this case, the need for a jackknife technique will fall away. Also note that, for skew distributions, the confidence limits for ore-block estimates will be skew as well, and symmetrical normal limits will not apply. RepIy by G.S. Adisoma and M.G. Hester There are a few interesting ways to utilize the block-by-block kriging variance that the authors have encountered. One of them is to classify a particular block into a proven-probable reserve or a possible-resource category, using a somewhat subjectively determined kriging variance value. Here, the kriging variance still represents the drilling density, i.e., data configuration more than the data values. Professor Krige mentions another ingenious approach through the use of relative variograms. This will yield a block kriging variance that takes into account local (block) grade, as he correctly points out, for purposes of confidence limits. However, we are more interested in quantifying the uncertainties associated with grade estimation in a much larger scale than a block. e.g., for a mining phase or the minable reserve. The jackknife, through the block kriging shortcut, provides a simple solution to obtain a single estimation variance for a general shape. The authors' main reference discusses different ways to handle skewed distributions typical in mineral resource estimation, e.g., the weighted jackknife approach or data transformation. Also, as with any technique, we can always improve our estimate by subdividing into smaller populations that are geologically and statistically separate domains. As we have mentioned; the jackknife is but a tool in our collection of tools to solve estimation problems. The jackknife's appeals are its simplicity and versatility. It is handy when no other solution is available, or when the available solution is not practical. It is used with other methods, not to replace them altogether. ?

Jan 1, 1997

By R. J. Sweigard, R. V. Ramani

G.F. Learning The paper by R.V. Ramani and R.J. Sweigard is a wonderful description of the labyrinthine web that has been spun about the mining industry by energetic bureaucrats and politicians over the past 50 years. The remedy for the problem, however, is not more of the same, but less. That may be difficult for the industry to achieve, for it is not a technical solution but a political one. And the current fervor for more detailed planning at all levels of government and private enterprise has become deeply ingrained. The authors recommend the provision of more information about mining and mineral resources to "macro" (i.e., government) land use planners. They apparently overlook, however, the already strong tendency on the part of most government land use planners to consider themselves omniscient. Thus, giving them more information about the technical problems of mining will only make them want to get more and more involved in the "micro" (private, site specific) mine development and production plans of the individual mining firm. In fact, this has already happened at all levels of jurisdiction from municipal to federal government. Examples are legion. The most effective way to ameliorate the adverse impacts of government land use planning on existing and potential mining operations is to: (1) introduce greater flexibility in the definition of land use zones by local and state governments; (2) adopt realistic and relevant ambient environmental performance standards in governing relationships between mineral land uses and concurrent or subsequent nonmining land uses; (3) allow greater leeway for economic considerations in land use decisions in contrast to the explicit legalistic approach now in vogue; (4) recognize that all minerals are not the same and that sand and gravel mining should not be treated the same as underground metal mining, coal stripping, oil field production, or in situ leaching; and (5) eliminate the notion that mining operators should be responsible for determining in detail the use of land by subsequent owners of mined land. This last bit of conventional ethic really makes no more sense than requiring the builders of every shopping center or government office complex to provide detailed plans for the use of that land when its use for shopping or government is ended. Did the builder of Ebbetts Field plan for Brooklyn after the Dodgers went to Los Angeles? Should the developer of the Bingham Pit plan for suburban Salt Lake City after the copper mining goes to Chile? The nation's mining industry must address these questions before further bankrupting itself to provide more data to planners and spending thousands of dollars per acre to create land that when reclaimed is worth only a few hundred dollars per acre. Reply by R.V. Ramani and R.J. Sweigard We thank Mr. Leaming for his valuable contribution. His views on the problems of land use planning and mineral resources are most welcome additions to our paper. As the title indicates, our paper was more concerned with the impacts of land use planning on mineral resource conservation than with the details of the planning process. On the whole, his five recommendations would be helpful for mineral resource conservation. However, we would suggest that the argument he presents for his final recommendation does not address the differences between mining as a land use and commercial or institutional uses. We believe that this difference is the crux of the issue. We share Mr. Leaming's desire to ameliorate the adverse impacts of land use planning. Possibly the most detrimental impact is the loss of mineral resources. Any development, whether mineral or community, that does not give proper consideration to other resources can result in permanent loss or sterilization of resources. With proper planning, some of these losses can be avoided. As our paper indicated, one factor that limits the consideration of mineral resources, and ultimately leads to their sterilization, is the generally inadequate levels of resource characterization and understanding of the unique nature of mineral resources and mining operations. The last point raised by Mr. Leaming is also important. In terms of reclamation and land use planning in mining districts, we certainly do not advocate spending more than what the results are worth. The main thrust of the paper was to explore the avenues for conserving the mineral resources so that, at some appropriate time, the issue of mining and reclamation can still be addressed.

Jan 4, 1985

By J. L. Weeks

Discussion by J.L. Weeks Discussion I wish to comment on Kissell and Sacks's paper "Inaccuracy of area sampling for measuring the dust exposure of mining machine operators in coal mines." It is well established in industrial hygiene that area sampling is not an accurate measure of personal exposure. To measure personal exposure, one must use a personal sampler. Kissell and Sacks, using data from coal mining, reached the same conclusion. This is the right answer but it is the wrong question. This issue arises because the machine-mounted respirable dust continuous monitor (MMRDCM) is being considered to monitor coal-miners' exposure to respirable dust. This device may be mounted on a continuous mining machine, and it would take a modified area sample. I refer to it as "modified" because a mining machine is not really a fixed location, as is the typical area sample. The concern with whether this machine would provide an accurate estimate of the machine-operator's exposure, however, overlooks important features of the MMRDCM. Specifically, it displays dust concentration as it occurs and at the place where it is needed, not 10 days later, as is the case with the current system. Used in this way, it allows for timely intervention to control exposure to dust and it leads to greater first-hand knowledge of specific conditions and mining practices that result in excess exposure. If one knows the dust level is excessive at the time the dust is being generated, it is possible to use known and effective methods of dust control to reduce exposure. By measuring again, one can determine whether control methods have worked. This is a generic procedure followed in any industrial process: measure, adjust and measure again. It leads to more effective and more efficient dust control and is preferred to taking a sample, waiting 10 days for the results and then trying to apply that information to a situation that no longer exists. The MMRDCM also records the dust level automatically, and it runs continuously, making it harder to tamper with. Dust monitoring in underground coal mines has been corrupted by some operators having submitted fraudulent samples (and being convicted on criminal charges because of it), weakening the credibility of the current system. The MMRDCM is, in part, a response to this corruption, and an effort to focus attention on dust levels and dust controls rather than on the perpetrators of fraud. Use of the MMRDCM is an evolutionary change from the current system. With the current system, one takes a personal sample of exposure of the "designated occupation" (DO, defined at 30 CFR 70.2 (f)), the occupation regularly exposed to the highest level of dust on a section. Typically, on sections using continuous mining machines, this is the mining-machine operator. The rationale behind DO sampling rests on some of the findings that Kissell and Sacks report. They reviewed several studies that showed that dust concentration varies with location on a mining section and that there is a concentration gradient with the highest concentration occurring closest to the face. The logic behind DO sampling strategy is that if the dust concentration at the place where it is highest is below the exposure limit, then concentration elsewhere on the section is also below the limit. In this way, one will have achieved the objective or keeping dust concentration below the exposure limit for all miners on the section by measuring the exposure of only one miner. The

Jan 1, 2003

By G. Chi, M. C. Fuerstenau, R. C. Bradt, A. Ghosh

Technical Papers, Mining Engineering Vol. 49, No. 12, pp. 70-75, Discussion by McIvor, R. and Fink, R. P. We must challenge what we consider to be dubious statements in the above-mentioned article. Much of what is reported was published elsewhere over a year ago (i.e., in their reference: Chi, G., Fuerstenau, M.C., Brandt, R.C. and Ghosh, A., 1996, "Improved comminution efficiency through controlled blasting during mining," International Journal of Min¬eral Processing, Vol. 47, p. 93). There is nothing unusual or significant about a 10% or more difference in grindability between two materials of apparently similar origin and other attributes (such as chemical assays, hardness and specific gravity). Consider also that: • the materials came from different locations in the pit; • only two sets of four blastholes were considered per blast: • spatial orientation with respect to the face and other blastholes was also altered; • only two pairs of data were collected; • the impact of sampling error and bias, • both of which could be significant in this case; and • normal experimental error in parameter measurement and calculation. It is clearly inappropriate to draw any conclusions with regard to cause-and-effect relationships without first presenting a thorough error analysis. A substantial number of randomly selected test and control areas might provide meaningful statistical evidence of the claimed effect of the variables being studied. Even that appears doubtful, however, because Figures 3 and 4 show about as much variation in crushing energy for the two "regular" conditions vs. regular and altered conditions. On a separate, but also important issue, broad experience with autogenous/semiautogenous milling has shown that smaller or weaker feed material results in reduced tonnage. ? Reply Thank you for the opportunity to respond to the comments of Messrs. Mclvor ant Fink. The points raised are addressed in the order of presentation. As the commenters noted, there are parallels between this article and a previous manuscript. The reason for these complementary articles is that the audiences for the two publications are different, although there is obviously some overlap. We believe that the different audiences each profit from the technical aspects of the two manuscripts. We are not in agreement that a grindability difference of 10% is insignificant. The laboratory work is corroborated by data obtained at plant scale, which confirms that an increase in throughput of the same magnitude (between 5% and 10%) is achieved when the blasting pattern is modified, i.e., reduced from 9 to 8 m (29.5 to 26.2 ft). This increase in throughput translates into substantial power savings for ore grinding. No differences in appearance in the rock were noted between the two adjacent areas. Rock in both areas belonged to the same rock mass/block, which was quite homogeneous and without geological features in between that could create drastic differences in comminution characteristics. Given the resources, we would have sampled from all of the holes. Additional data would certainly have been desirable, as would have been additional funding to expand the study. Their comment, "spatial orientation with respect to the face and other blastholes was also altered," is not clear. Figure 1 does not show the initiation sequence. Without the initiation sequence, we do not understand how the commenters concluded that spatial orientation with respect to the face and other blastholes was altered. The commenters may be trying to make the point that the four-hole configurations with respect to the original bench face are not exactly the same. In fragmenting the

Jan 1, 1999

By R. F. Kovar, R. N. Torbin, R. E. Gundersen

I have to express my .surprise and must send a note of warning concerning the use of polyurethane foam as described in a paper by R.N. Torbin and R.F. Kovar, entitled "Innovative Shaft Lining Method," published in your February 1989 edition. In September, 1986, 177 miners lost their lives when they were exposed to fumes downwind from a fire in a tunnel lined with polyurethane foam. Whilst polyurethane exhibits reasonable fire retardent properties in small-scale tests, mass use in a tubular configuration, exacerbated by the furnace effect of cross-radiation, can lead to a rapid, uncontrollable conflagration. I do not wish these comments to detract from the engineering achievements of the authors but rather to sound a loud warning as to the application of the system. Any shaft designed for human access- be it for transport or shaft work - and any intake airway, should not consider the use of polyurethane foam. It only needs a slip with a cutting torch or an electric fault to initiate a disaster. Many countries have banned or restricted the use of polyurethane foam in the workings of a mine and maybe alternative foams (e.g. phenolics) can be used or developed

Jan 1, 1991

By G. Sanders

Discussion by G. Sanders The studies reported on in this paper were initiated to draw attention to the severe contamination problem in the Section 30 drilling program at Chimney Creek. The lithologic-subset sampling study reached a different conclusion from that presented in your paper, and I wish to comment on your subsequent analysis of the data and your conclusions. Request for more complete data In the section on subsampling, you mention that the subordinate lithologies were separated and sampled, yet only the dominantlithology gold value is plotted in Fig. 4. In a contamination study, the reader is interested in the assay values for the individual subsets. Please include a table of the subsample assay data in your reply. Also, please indicate which analytical methods were used to arrive at the gold values in the subsampling study. Turning barren rock into low-grade ore Figure 5 is very revealing and typical of all of the cross sections in Section 30. Note the long strings of low-grade mineralization spread out for hundreds of feet below the ore zones. There were some very high gold values found in certain contaminated fractions during the subset sampling. The conclusion, here, was that the distinctive, strongly-mineralized dolomite layer was probably loose and crumbly and continued to disintegrate during drilling. This caused salting of the unmineralized rock samples below. Missing the high-grade part of the ore body In your statistical analysis, you directly compare the reverse circulation assays to the diamond drill assays in Section 30. Two points argue against a direct comparison and suggest the differences are greater than the 3 % that you report. First, any core loss in a gold zone most likely means that the true gold values are greater. The drillers lost significant amounts of the clay-rich, Section-30 gold mineralization. Also, the initiated salt-mud system, an attempt to improve the core recovery, met with little success. Second, the practice of not sampling core geologically, but instead sampling on even 5-foot intervals, adds a deliberate dilution to the core assay values by including a portion of nonmineralized rock in the first and last samples of each high- grade intercept. The result is often a pair of low-grade assay values on either side of a high-grade gold zone. In reality, a high-grade gold zone has a very sharp assay wall that is often bounded by barren rock. This sampling method may make the diamond-drill core assays more like the reverse circulation values and may help explain the statistical similarities you found. However, it does not represent the true gold values in the high-grade parts of the deposit. You cannot deny that, by careful geological sampling of the drill core, higher and sharper assay values will be obtained. The low core recovery and the diamond-drill-core sampling method used act together to lower the diamond-core assay values. The 3% difference you found between the reverse-circulation and diamond-core assay values could be much larger when you consider what the true diamond-drill core values would be with optimum core recovery and a geologic sampling method for the core. Should statistics have been applied here? The statement "... that reverse circulation holes have overestimated the values of some ore zones and underestimated the values of others" (p. 345) is not correct. The subsampling confirmed what the cross sections hinted at in Section 30. Namely, beneath the high-grade zones, the reverse circulation holes created, by contamination, large intercepts of low-grade ore in regions of barren rock. Because the low-grade material was not there to begin with, this is not a process of overestimating low-grade mineralization. The next statement that "the average result is similar to that of the diamond drill holes" may apply to the data set numerically, but it is not true when viewed spatially on cross sections. Adjacent reverse circulation and diamond drill holes are almost impossible to correlate, high-grade zone values vary widely and many low-grade intercepts make no geologic sense. The subset sampling and cross sections presented in the first part of the paper show that the reverse circulation portion of the data set has some serious problems, as highlighted above, and should not have been dealt with statistically at all. Conclusion Each ore body is different, and each drilling method presents unique sampling problems. In this case, the diamond drill is the

Jan 1, 1994

M.S. Azun I have many objections to the content of the author's paper. Before discussing it, however, I would like to repeat the property of semivariogram function. Second order stationary properties of regionalized variables (ReV's) such as semivariogram function y(h) are perfectly known in geostatistics. Also, the kriging equations in the language of mathematical statistics using second order stationary properties are well understood. However, the way to use the sample (estimated) semivariogram function in any one of the kriging procedures is vague. The sample semivariogram function is given as follows: [ ] where N is the total number of samples, Xi is the sample value at the i - th location, Xi+h is the sample value at the i + h - th location, and h is the distance among the samples. An estimation variance of sample semivariogram function of first lag is smaller than that of higher order lag. The theoretical semivariogram function reaches the variance of samples asymptotically. But this is not easily observable because of the larger variation involved in the estimate of semivariogram function. In general, an estimation procedure is done for h = 0, 1, 2, up to the greatest integer less than N/2, even though sample semivariogram function can be computable through N-1. After estimating semivariogram function, the critical question of how to model sample semivariogram function arises. As seen in the above equation, sample semivariogram function is discrete and can be smoothed by the model being selected. Therefore, modeling of sample semivariogram function is the most important step in geostatistics. It not only smoothes a discrete function but also affects the results of the kriging procedure. When the only aim is to model the semivariogram function, which is the basic point of the author's paper, one can employ any fitting techniques, such as curve fitting, or any arbitrary functions, which are called submodels in the paper. The term "arbitrary function" is used rather than "submodel" because there is no basic understanding of developing them. The author suggests that the sum of those submodels can also be used for the modeling of sample semivariogram function. The combination of any arbitrary functions brings many problems instead of giving an insight of the domain structure considered. The author used two arbitrary functions and the nugget effect in response to sample semivariogram function (Fig. 10). For the same example, he stated that the parameters involved in the mixed arbitrary function model can be accepted when the discrepancy between sample semivariogram function and the model is small visually. For verifying the fitting behavior of any selected model, one should not be contented with the visual satisfactory. Some statistical measure such as goodness of fit has to be used. The author's practice is no more than an exercise in curve fitting without any fundamental understanding or conceptualization of the underlying physical mechanism. Furthermore, the selection of any model is not an easy task if the purpose is the search for the "best" response to the observed second order properties of ReV's. I suggest that the Markovian model (Azun, 1983), on the basis of a theoretical understanding of underlying mechanism, which gives more information about the occurrence of regionalized variables, is used to respond all properties of ReV's. There are a lot of problems for modeling of one-dimensional sample semivariogram function. Thus, it is not appropriate to go to higher order dimensional sample semivariogram function modeling. In the meantime, I would recommend that one can connect the values of standardized sample semivariogram function rather than simple values of semivariogram function in the two-dimensional estimation. The standardized values can be computed in dividing the semivariogram function value by the number of sample pairs involved in each lag regardless of the directions. In conclusion, geostatistics is an interdisciplinary area in mining that uses the principles of mathematical statistics. Thus, it should not violate any probabilistic and statistical rules. When Matheron was developing the theory of geostatistical study in the early years of geostatistics, many mining people had a reservation accepting the geostatistical tools. However, this does not mean that we, the geostatisticians, might try to convince those people using some "strange" tools or rules as some authors implied (Baafi and Kim, 1984). Instead, we have to develop and explain the geostatistical tools staying only in the framework of statistical concepts and properties. References Azun, M.S., 1983, "Stochastic Process Modeling of Spatially Distributed Geostatistical Data," Columbia University, Ph.D. Thesis. Baafi, E.Y., and Kim, Y.C., 1984, "Discussion - Comparison of Different Ore Reserve Estimation Methods Using Conditional Simulation," Mining Engineering, Vol. 36, No. 3, p. 280. Reply by J.M. Rendu The interactive method proposed by Rendu allows practitioners to develop semivariogram models that take into account not only the numerical information obtained by sampling, but also highly significant additional information that often cannot be quantified. The geology of the deposit - including hypotheses concerning its genesis, sampling methods, assaying methods, and mathematical methods used to calculate the semivariograms - all have an influence on the numerical results obtained and on how these results should be interpreted. If all the information concerning the spatial distribution of values in a mineral deposit was contained in the sample values, it could be argued that statistical techniques alone would produce optimum models. However, this is rarely, if ever, the case. Methods that allow the user to take into account his experience and his geologic understanding of the deposit should not be rejected for the sake of theoretical statistical purity.

Jan 11, 1985

By B. Beatty

Discussion by B. Beatty This is a response to M.J. Hrebar's paper, which tracks in detail the highly productive stripping capabilities of modern large bulldozers at two separate installations. It proves the old adage "economies of scale" and quantifies the advantage of being able to move large loads of dirt over comparatively short haul distances by sliding rather than carrying the load. This discussion examines whether a bulldozer is the most efficient way of using this geotechnical principle. Case 1 - Study eastern US: Contour coal stripping application The energy consumption for the Case 1 mining method (eastern US contour stripping operation) shows the D11 averaged 117 L (31 gal) of diesel fuel per scheduled hour, and produced 333.3 bank m3 (436 bank cu yd) per scheduled hour (0.351 L/m3). Assuming 1 L (0.26 gal) of diesel fuel contains 10.4 kW-hr of energy, the unit expenditure of fuel energy for Case 1 Study was 3.66 kWh/m3. Another way of shifting this material is by using an overburden slusher (OS), as shown in Fig. 1. The OS as shown in Fig. I. The OS is electrically powered and consists of a manned low wall winch (LW ), a remote controlled (radio telemetry) highwall winch (HW) and a hoe. The winches operate to pull the hoe back and forth across the pit. The hoe carries load in the forward direction ahead of a rear mounted blade and returns empty. A spreadsheet (Lotus 123) simulation package, BOSSIM, models the OS operation on a load by load basis over a wide range of pit configurations. The geometry shown in Fig. 2 of the paper was modeled using BOSSI M and the results are summarized in Table1. Table 1 shows a LW with 150 kW (200 hp) and a HW of 120 kW (160 hp) installed power, towing a 15-m3 (19-cu yd) hoe, would accomplish the Case 1 stripping task at about the same scheduled stripping rate (D11=333 m3/hr or 433 cu yd/ hour, OS=342 m3/hr or 447 cu yd/ hour). It was assumed the winches in this equipment have an electromechanical efficiency of 90%. This gives an average electrical energy expenditure rate of 151 kW-hr, and a unit energy expenditure level of 0.44 kWhr/m-3(0.57 cu yd/kWh)-an improvement of 8291%- over the D11 alternative. The energy comparison illustrates the importance of using more efficient electric power if possible, ahead of the internal combustion engine, and delivering power from a fixed ground position rather than through a moving track system. Another important energy observation is the comparative action of the dozer blade and the OS hoe. The former implement continuously digs and spills as it works the load forward. The OS hoe digs the load until full, the blade then lifts due to the contained load pressure and a shear plane forms under the mass, enabling it to slide forward in a force reduced - transporting mode. Economic comparison Table 2 operating cost per scheduled hour and unit cost comparison follows a format used in Tables 2 and 3 of Hrebar's paper and compares Case 1 Study with the OS alternative. Drill and blast costs are kept constant in both cases, although when larger OS units are considered, it is expected these costs will reduce because bigger hoe sizes will permit larger rocks to be transported efficiently.

Jan 1, 1995

L.R. Bower In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the µ factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the µ of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure ; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves. Supports to this design of both 450- and 280-t (496-and 309-st) rating have been successfully used in the United Kingdom for several years, negotiating many faulted areas without one single reported need for repair or maintenance. This includes supports left unattended during the year-long strike, proving the reliability of the system.

Jan 8, 1986

By J. B. Gwiazda

S. Budirsky J.B. Gwiazda's article deals with an interesting problem that has not been studied thoroughly up to now. Gwiazda has proposed a technical solution that eliminates horizontal load imposed by the supports on the roof. That load is due to a discrepancy between the trajectories of the canopy and the overlying strata. It is true that the improved lemniscate guidance system proposed by Gwiazda could eliminate the horizontal load but, on the other hand, it would represent a higher cost of the supports, and it would complicate the already complicated mechanism of powered supports. That is why we must be convinced that the proposed improvement is actually needed. From that point of view, the following questions should be elucidated. As shown in Fig. 1, the point 0 travels along the curve a, which produces horizontal load on the roof in the direction of the coal face. As stated by Gwiazda, "such a high load is capable of destroying the roof above the support, causing rock debris to be scattered around the face." The results of our measurements carried out in coal mines of the Czechoslovak part of the Upper Silesian Coal Basin show that the convergence in the modern type of powered supports does not exceed several millimeters (not more than 10) per hour. Similar conclusions were drawn from field measurements performed in the Polish part of the Upper Silesian Coal Basin (see S. Romanowicz and H. Szopka in Proceedings, Scientific and Technical Symposium Simmex'85, Katowice, Poland, pp. 73-83). The small vertical closure of the supports also induces a small horizontal displacement of the canopy. In addition to this, the transmission of forces between the roof and the canopy is influenced by the layer of debris on it, i.e., the canopy slides without imposing a greater horizontal load on the roof. This was deduced from field measurements of Voest-Alpine F 4/4500 powered supports equipped with rams instead of the common front lemniscate links (denoted as 1 in Fig. 5). In these supports, the pressurizing of the space of piston rod (1a in Fig. 5) causes the rams to close, which induces horizontal load on the roof in a direction toward the coal face. After the setting of a support unit, the pressure in the piston rod space was found to drop, frequently to zero, because the small horizontal sliding between the canopy and the debris or inside the debris caused the horizontal reaction from the roof to fall to zero, although the debris were pressurized by vertical setting load. As a consequence, there was no horizontal load imposed by the canopy on the roof. A rise of the horizontal load was recorded only after the passage of the shearer and during lowering of the adjacent units. Thorough information on the results of the aforementioned field measurement is given in the paper, "Analysis of the performance of shield powered supports installed in a thick seam," published in Mechanizacja i automatyzacja gornictwa, 1982, No. 12, pp. 38-46. An English translation of the paper is available from the author. We have deduced from our observations that horizontal load imposed by the canopy on the roof in the direction toward the coal face is useful for roof control because it limits the displacement of the immediate roof toward the gob particularly during the rise working. We came to the conclusion that horizontal load on the roof toward the coal face should be induced on purpose. Could Mr. Gwiazda prove with the results of field measurements his contrary opinion? Has the small horizontal displacement of the canopy toward the face actually had an adverse effect on the roof? As shown in Fig. 2, during the setting of a support unit, the canopy imposes a horizontal load on the roof in the direction of the gob. In this case, I share Gwiazda's opinion that the horizontal load could have an adverse effect on the roof. Nevertheless, two important circumstances were neglected - the horizontal compliance of the supports and the interaction of adjacent units. The question of the compliance was analyzed precisely by I. Krumnacker in Gluckauf-Forschungshefte, 1984, No. 5, pp. 219-223. He found that due to the clearance between the hinge pins and the eyes of the sheild and due to the elasticity of the steel structure,

Jan 1, 1987

By J. B. Gwiazda

S.S. Peng The highlights of this paper are : • The shield supports with lemniscate links will exert a tangential force on the roof strata, causing unnecessary and harmful roof movements, thus damaging the roof structure. The principal source of the tangential force is the horizontal movement of the point O, the pivotal joint between the canopy and the caving shield. When the support height is changed during setting, the point O will move along a curve upward and rearward toward the gob. As a result, it produces a vertical movement, causing the roof to move up a distance of ?d, and a horizontal movement toward the gob, creating vertical fractures in the roof. • The major purpose for employing the lemniscate-link system is to protect the hydraulic legs against the action of the bending moments induced by the tangential force that causes deformation or destruction of the legs. • Based on these two points, the author suggested that, in order to eliminate the harmful effects of the lemniscate-link system and strengthen its advantageous features, a shield support with a new design of that mechanism has been developed. With the new design, the conventional rigid lemniscate links are replaced by hydraulic rams. When the pressure inside the rams reaches a certain level, the ram will yield.

Jan 12, 1986

By G. W. Heunisch

G.A. Carmody and L.B. Anderson We wish to comment on the paper by G.W. Heunisch (MINING ENGINEERING, January 1987), which raised several important points regarding treatment of acidic waste waters. This paper confirms, in part, tests carried out by Allyn Corp. Allyn is the manufacturer of the proprietary product Alka-Ment, one of the materials that Heunisch tested. The first point is the extreme sensitivity of the sedimentation rate on the nature of the treatment material used. This apparently results from differences in nucleation of the metal hydroxides, and is consistently higher for Alka-Ment. This decreased sludge volume is also observed when Alka-Ment is combined with flocculants, such as organic polymers (see Table 1). Our second point is to emphasize that Heunisch's work shows how dependent the sludge volume is on pH. It is common practice to overtreat AMD by raising pH to 9 or 10. This practice, in addition to being wasteful of material, actually decreases the density of the resulting sludge, probably by creating charged particles that do not readily coalesce. When treating with Alka-Ment, we recommend that the treatment be monitored by testing both pH and dissolved iron. For both tests, it is important to allow adequate time to elapse between treatment and measurement, so that dissolution of alkali, aerobic oxidation of iron, and precipitation of the metal hydroxides are nearly complete. Our tests have indicated that from 30 minutes to one hour is commonly required. Only in this way can a minimum pH be attained, with adequate treatment, denser sludge, and lower treatment rates achieved. Thirdly, Heunisch correctly recognizes that ASTM titration of Alka-Ment does not reflect the true treatment value available in a unit of Alka-Ment. Our work at Allyn indicates this is a most important consideration when evaluating various materials for the treatment of acidic waste waters. While we were unaware of Mr. Heunisch's work, we welcome comparisons between Alka-Ment and any competitive products readily available for field use. Allyn has hundreds of hours of laboratory and field data we are ready to share with interested parties treating acidic waste water.

Jan 9, 1987

By L. B. Anderson, G. A. Carmody

We wish to comment on the paper by G.W. Heunisch (MINING ENGINEERING, January 1987), which raised several important points regarding treatment of acidic waste waters. This paper confirms, in part, tests carried out by Allyn Corp. Allyn is the manufacturer of the proprietary product Alka-Ment, one of the materials that Heunisch tested. The first point is the extreme sensitivity of the sedimentation rate on the nature of the treatment material used. This apparently results from differences in nucleation of the metal hydroxides, and is consistently higher for Alka-Ment. This decreased sludge volume is also observed when Alka-Ment is combined with flocculants, such as organic polymers (see Table 1). Our second point is to emphasize that Heunisch's work shows how dependent the sludge volume is on pH. It is common practice to overtreat AMD by raising pH to 9 or 10. This practice, in addition to being wasteful of material, actually decreases the density of the resulting sludge, probably by creating charged particles that do not readily coalesce. When treating with Alka-Ment, we recommend that the treatment be monitored by testing both pH and dissolved iron. For both tests, it is important to allow adequate time to elapse between treatment and measurement, so that dissolution of alkali, aerobic oxidation of iron, and precipitation of the metal hydroxides are nearly complete. Our tests have indicated that from 30 minutes to one hour is commonly required. Only in this way can a minimum pH be attained, with adequate treatment, denser sludge, and lower treatment rates achieved. Thirdly, Heunisch correctly recognizes that ASTM titration of Alka-Ment does not reflect the true treatment value available in a unit of Alka-Ment. Our work at Allyn indicates this is a most important consideration when evaluating various materials for the treatment of acidic waste waters. While we were unaware of Mr. Heunisch's work, we welcome comparisons between Alka-Ment and any competitive products readily available for field use. Allyn has hundreds of hours of laboratory and field data we are ready to share with interested parties treating acidic waste water.

Jan 1, 1988

By R. C. Kirkman

Your interesting and well-written paper brings up the issue of the purpose or use of a mineral property appraisal, specifically the value of these evaluations when the quality and quantity of the mineral resource to be exploited have not been thoroughly investigated. I am not a practicing professional in regards to valuation of limestone investment opportunities. I simply wish to make a few generic observations about the appraisal process in order to stimulate meaningful thinking by appraisal professionals whose reports may be reviewed by professional investors and "laypersons" alike. It is my contention that the analysis given in your paper is very thorough, considering the data that was made available to you. But I would not like to be the person responsible for investment decision making without having a feel for the risk involved. Therefore, the reviewer must classify this type of analysis as relatively "subjective" because they have not received an indication of the inherent risk involved in exploiting a mineral resource. This risk can best be quantified by using probabilistic methods in conjunction with the quality and quantity of the mineral to be produced. Most appraisals passively ignore this problem. Specifically, you have accurately identified the various costs and values entering the economic appraisal process. These can usually be quantified with a great deal more certainty than the capability of the resource to provide 25 years of quality material. Also, in Cases A (West Virginia)and B (Alabama), you are fortunate enough to have previous operations data. What if there had never been production from these properties? How does a potential mineral property investor evaluate the resource capability to meet various projected financial outcomes without spending a good deal of up-front capital to drill and delineate enough quality resources for a 20-plus year operation. These questions pose a substantial risk to any potential mineral investor. Only probabilistic methods should be used to quantify this type of problem, short of drilling and metallurgical testing. I might be considered a "layperson," not being familiar

Jan 1, 1991

By C. W. Hunt

I wish to register strong disagreement with the choices of drilling sites chosen by the Continental Scientific Drilling Committee, while endorsing the stated objectives. Of special significance is the lack of any site involving the metals most needed today in the US, while 9 to 13 km of drilling are allocated to gold and silver districts and 11 to 17 km to base metal (mainly copper) districts. Why is nothing proposed for platinum group or chromite? Is this because the committee did not know of such a site? If so, I offer such a site in California (see Fig. 1). The mineralization shown in Table 1 occurs in a large peridotite body in western Tehama County, CA. Petrography has demonstrated unequivocal evidence of migrated PGE deposited in peridotite and chromite of hypabyssal nature. Active diapirism (due probably to serpentine buoyancy) above a reactivated mantle to surface conduit marks the position of the conduit. Young explosion breccia cemented loosely by volcanic glass and tuff demonstrates the significance of the conduit as the prime source of the various emissions. Deep drilling could result in major metals discoveries as well as shedding light on the grand questions of platinum group mineralization. What role water has played in this occurrence is not yet certain; but chlorine and asphaltite have been detected in and adjacent to native PGM, suggesting a major role for water. The explosion breccia suggests that this excellent site could overlie entrapped geothermal steam.

Jan 1, 1988

By C. Meyer, J. J. Eidel

C. W. Hunt I wish to register strong disagreement with the choices of drilling sites chosen by the Continental Scientific Drilling Committee, while endorsing the stated objectives. Of special significance is the lack of any site involving the metals most needed today in the US, while 9 to 13 km of drilling are allocated to gold and silver districts and 11 to 17 km to base metal (mainly copper) districts. Why is nothing proposed for platinum group or chromite? Is this because the committee did not know of such a site? If so, I offer such a site in California (see Fig. 1). The mineralization shown in Table 1 occurs in a large peridotite body in western Tehama County, CA. Petrography has demonstrated unequivocal evidence of migrated PGE deposited in peridotite and chromite of hypabyssal nature. Active diapirism (due probably to serpentine buoyancy) above a reactivated mantle to surface conduit marks the position of the conduit. Young explosion breccia cemented loosely by volcanic glass and tuff demonstrates the significance of the conduit as the prime source of the various emissions. Deep drilling could result in major metals discoveries as well as shedding light on the grand questions of platinum group mineralization. What role water has played in this occurrence is not yet certain; but chlorine and asphaltite have been detected in and adjacent to native PGM, suggesting a major role for water. The explosion breccia suggests that this excellent site could overlie entrapped geothermal steam.

Jan 6, 1987

By M. M. Botz, F. W. DeVries

Discussion by F.W. DeVries Comments on natural cyanide attenuation. The paper by Botz and Mudder in the November 2000 issue of Minerals and Metallurgical Processing had one disappointing aspect. Table 1 (p. 229) failed to include a significant species - neutral copper cyanide (CuCN). This compound has been most trouble-some in many plant-scale cyanide cleanup processes, includ- ing acidification-volatilization- reneutralization (AVR), hy- drogen peroxide cleanup and air/SO2 cleanup. It has also plagued many pilot efforts to use ion-exchange (IX) resin technology. This author elucidated the concept that, under some conditions, copper cyanide failed to precipitate out of treated effluents and appeared to exist as a filterable colloid that reported as copper and cyanide, with the latter measured as WAD (weak-acid-dissociable) cyanide in that analytical method. The concentrations measured were very near the levels of 0.2 ppm, often the regulatory limit for WAD cyanide. The copper concentration, using the 1 : 1 ratio for this com- pound, would then be more than 0.4 ppm, which could also be unacceptable. We discovered that it was unsafe to assume that the Cu(CN),- would surrender both cyanide ions simultaneously and allow cupric copper to be removed subsequently as, say, cupric ferrocyanide (Hatchett's Brown). This compound could be ignored as safely as the "iron blues," as essentially an inert solid, stable in geologic time if not exposed to sunlight. Under poorly controlled oxidation of the cyanocuprite ions (copper with four, three or two cyanides complexed to the copper), CuCN could well be formed and could not be assumed to precipitate out as a solid. The authors of the paper, by failing to include the CuCN species in their model, may have left users of the model prone to ignore the possible occurrence of the colloidal(?) form in the discharge, thus finding an effluent that failed to meet regulatory limits. (See addendum below.) This somewhat ephemeral material may well have been responsible for some large-scale failures of cyanide destruction technology to succeed at sites in North America and elsewhere. Two specific references alluding to this occurrence may be found: Griffiths et al., 1987, "The detoxification of gold-mill tailings with hydrogen peroxide," Journal of the South African Institute of Mining and Metallurgy, September, p. 282. DeVries, 1989, "Improved understanding in detoxification of cyanide leach liquors using peroxide" Rand01 Gold Conference, Sacramento CA, p. 15 1. It is advisable for workers in this area of cyanide removal or recycle to avoid assuming that CuCN will, with certainty, precipitate out of solution, or that the dicyanocuprite ion will part simultaneously with both cyanide ions and allow cupric copper to be dealt with by other reactions. One further piece of work may be incomplete. That is, a definitive study of the long-term behavior of CuCN that does precipitate when exposed to other compounds present in neutralized tailings. Addendum. The extremely complex reaction of cupric ion on cyanide- which, of course, results in oxidation of cyanide to cyanate, probably through the intermediate stage of forming cyanogen [(CN)2] from oxidation of HCN with its hydrolysis to cyanate and cyanide ion - involves two cyanide moieties and the reduction of the cupric ion back again to the cuprous state. This, then, can form the CuCN potential "solid" that was already the undesirable species. In cyanide cleanup, with either air/SO2 or hydrogen peroxide, copper is generally employed as the catalyst, so it is unclear what happens when the CU(CN)2-1 is ultimately exhausted. Certainly, the field evidence is quite clear that all

Jan 1, 2002