Energy analysis process data

Transcription

1 Nuclear power the energy balance Jan Willem Storm van Leeuwen Ceedata Consultancy Note In this document the references are coded by Q-numbers (e.g. Q6). Each reference has a unique number in this coding system, which is consistently used throughout all publications by the author. In the list at the back of the document the references are sorted by Q-number. The resulting sequence is not necessarily the same order in which the references appear in the text. October 2007 Part E Energy analysis process data Contents E1 Nuclear process chain: outline and uranium mass balance E2 Processes of the nuclear chain Exploration Transport 4 Mining and milling 3 Conversion 2 Enrichment 1 Fuel element fabrication 0 Construction of the reactor 0 Operation, maintenance and refurbishments of the reactor 5 Reconversion of depleted uranium 6 Decommissioning and dismantling of the reactor 7 Radioactive waste handling 8 Geologic repository for radioactive waste, except spent fuel 9 Interim storage spent fuel 10 Spent fuel conditioning 11 Geologic repository for spent fuel 12 Reclamation of the uranium mine E3 Summary tables of the basic chain parameters References Part E 1

2 E1 Nuclear process chain, outline The processes of the nuclear system are in this study numbered according to Figure E.1. Figure E.1 Process chain of the LWR once-through system Each of the processes will briefly be discussed below. More details and historical data can be found in the original study Storm&Smith 2005 [Q6]. Exploration and transport are not included as separate processes, but are assumed to be included in the other processes, see below (Part E2). Part E 2

3 Uranium mass balance Figure E.2 Uranium mass balance of the LWR once-through system Table E.1 Uranium mass balance (in Mg) of the LWR once-through chain. See also Part B. Process Symbol Relationship Reload x p = 4.2% First core x p = 3.3% Mill tailings m tail = 1.18* m 3 *(100/Y*G 1) Conversion m 3 = m Enrichment feed m 2 = F Feed/product ratio F/P Enrichment product P = m 1 + loss2 = m Depleted uranium m depl = F - P Fuel element fabrication m 1 = 1.01 m Reactor m Part E 3

4 E2 Processes of the nuclear chain Exploration In most analyses, the exploration needed to discover new uranium ore resources is not taken into account. In the literature, the quoted values vary widely. Mortimer 1977 [Q98] gives a range of GJ/Mg discovered U 3 O 8. Data in Oregon 1974 [Q146] point to about 180 GJ/Mg discovered U, and Kolb et al [Q144] cite a figure of 250 GJ/Mg U 3 O 8. Future exploration and prospecting may be more difficult, more expensive and therefore more energy-intensive than in the past, because the easily discoverable deposits are known already. To discover new deposits, more and deeper drilling may be required. An energy requirement of TJ/Mg discovered uranium, as would follow from above mentioned references, may not be negligible with regard to the energy input of mining and milling: some 2.0 TJ/Mg U at an ore grade of 0.15% U 3 O 8. At leaner ores the fraction of exploration energy input decreases as the energy input of mining and milling exponentially increases with decreasing ore grade. In this study the energy input of exploration is assumed to be included in the mining and milling energy inputs. Transport Transport of materials throughout the nuclear chain consumes a neglible amount of energy, except mining, compared to most other processes, as is shown in earlier studies, e.g. ERDA-76-1 [Q109], see also ISA 2006 [Q325]. In the mining of uranium, the transport of rock consumes a significant part of the mining operations. These energy requirements are included in the total for mining. In this study, the energy for transport in all other phases is assumed to be included in the other activities. 4a Mining Assumed to be the same value for soft and hard ores. From Rotty et al [Q95] and ERDA-76-1 [Q109]. J mining = 1.06 GJ/Mg ore R = J th /J e = 8.0 Eq E.1 This figure of Rotty regards the average of open pit (60%) and underground mines (40%) and is based on a survey of US uranium mines in The figure includes indirect energy inputs. Part E 4

5 4b Milling Milling energy requirements per metric tonne (Mg). In this study the value from ERDA-76-1 [Q109] is used for soft ores: J milling = 1.27 GJ/Mg ore R = J th /J e = 7.0 Eq E.3 and the value from Kistemaker 1976 [Q116] for the hard ores: J milling = 4.49 GJ/Mg ore R = J th /J e = 0.10 Eq E.4 4 Mining + milling The total specific energy requirements for mining + milling per Mg ore, as adopted in this study, are the sum of mining and milling energy inputs. Soft ores: J m+m = 2.33 GJ/Mg ore R = J th /J e = 3.0 Eq E.5 and for hard ores: J m+m = 5.55 GJ/Mg ore R = J th /J e = 0.69 Eq E.6 The energy requirements of the recovery of one mass unit of uranium leaving the mill, J m+m, depend on the ore grade G and the recovery yield Y, and can be calculated with the following equation: In our study Q6 we calculated the values of c: Eq E.7a soft ores c = TJ/Mg (U) R = J th /J e = 7.5 hard ores c = TJ/Mg (U) R = J th /J e = 1.6 Eq E.7b To verify equation E.7, we compared the results of the process analysis of the Ranger mine in Australia (see Part D7) with the outcome of the equation under the conditions of Ranger. The actual figures of Ranger turn out to be slightly higher than the outcome of the equation E.7: see Table E.2. As Ranger is a large highquality uranium mine, with production costs in the lowest region of the cost spectrum (see Figure D.10 in Part D6), we may conclude that Equation E.7 will not lead to overestimation of the energy input of the uranium recovery. Part E 5

6 Table E.2 Specific recovery energy requirements and carbon intensity of the Ranger uranium mine in Australia and of a reference mine with the same ore grade in Storm & Smith 2005 [Q6]. The Ranger ore is assumed to be soft. This process analysis is based on data from ERA 2005 [Q321] and ERA 2006 [Q320]. Ore grade % (w/w) U 3 O 8 Specific energy requirements TJ/Mg U nat Carbon intensity gco 2 /Kwh Ranger Reference mine Extraction yield Equation E.8 gives the mathematical relationship between the yield of the mining and milling process and the ore grade represented by the green curve in Figure E.3: Eq E.8 Figure E.3 The empirical relationship between extraction yield Y and ore grade G (blue curve) is followed in this study. The green curve represents the mathematical relationship between the empirical data and the hypothetical data (green diamonds). At grades higher than 0.05% U 3 O 8 the theoretical relationship coincides with the empirical relationship, the blue curve in Figure E.3. In this study the empirical Y curve is followed to calculate the energy input of the extraction of uranium from its ore. Below an ore grade of 0.013% U 3 O 8 no empirical or scientifically verifiable data are available, as pointed out in Part D3. The only reason to include equation E.8 and Figure E.3 here is because the calculations in the original study Storm&Smith [Q6] are based on the hypothetical relationship and its curve. The outcome of this update of the study happens to be hardly dependent on the choice of which curve is used. Part E 6

7 Table E.3 Recovery data of uranium per Mg uranium leaving the mill (= m 3 in the uranium balance of Figure E.2), soft ores. Based on the empirical recovery yield curve. This table is identical to Table D.6 in Part D4. Grade, G % U 3 O 8 Yield, Y E th + E e TJ/Mg U E th TJ/Mg U E e TJ/Mg U m(co 2 ) Mg/Mg U CO 2 emission g/kwh * * Taken on the gross electricity production of the reference reactor, per reload charge in the steady state. This figure depends on the characteristics of the reactor in which the uranium is used, e.g. enrichment assay and burnup. Only the thermal (= fossil) energy input is accounted for in the CO 2 emission (see Part C1). Table E.4 Recovery data of uranium per Mg uranium leaving the mill (= m 3 in the uranium balance of Figure E.2), hard ores. Based on the empirical recovery yield curve. This table is identical to Table D.7 in Part D4. Grade, G % U 3 O 8 Yield, Y E th + E e TJ/Mg U E th TJ/Mg U E e TJ/Mg U m(co 2 ) Mg/Mg U CO 2 emission g/kwh * * Taken on the gross electricity production of the reference reactor, per reload charge in the steady state. This figure depends on the characteristics of the reactor in which the uranium is used, e.g. enrichment assay and burnup. Only the thermal (= fossil) energy input is accounted for in the CO 2 emission (see Part C1). Part E 7

8 3 Refining of yellow cake and conversion to UF 6 In this study the value according to ERDA-76-1 [Q109] is used: J conv = TJ/MgU R = J th /J e = 27.1 Eq E.9 Process loss is 0.5%, according to NRC 1996 [Q16]. Table E.5 Data on refining and conversion of uranium for the first core and reload charge of the reference reactor Quantity Unit Reload First core input m 3 Mg conversion energy, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh Part E 8

9 2 Enrichment Figure E.4 In the enrichment process the feed F of natural uranium is separated into the product fraction P (enriched uranium) and the waste fraction W (depleted uranium) The ratio of feed mass F and product mass P depends on the product assay and tails assay and can be calculated by equation E.10: Enrichment product mass: Eq E.10 P = m 1 + loss 2 Eq E.11 Table E.6 Separative work and feed mass for enrichment product assay x p tails assay x t = tails assay x t = S F/P S F/P x = assay % U-235 S = separative work SWU per kg uranium entering the reactor F/P = feed/product ratio References: Jan & Krug 1995 [Q29], Scheidt 1995 [Q30], DOE/EIA 1997 [Q64] In this study a tails assay of x t = (fraction U-235) is assumed. To produce a Mg of enriched uranium with product assay x p less natural uranium is needed when Part E 9

10 depleting natural uranium to a lower tails assay than to a higher tails assay, indeed at the expense of more separative work and higher energy requirements for enrichment. The MIT study MIT 2003 [Q280] assumes a tails assay of , as is common practice at low uranium prices. Table E.6 is added to illustrate the effect of the tails assay on the feed-product ratio and the amount of separative work needed for enrichment. Separative work S (amount of separative work units SWU) can be calculated by equation E.12 (from DOE/EIA 1997 [Q64]): Eq E.12 Diffusion In this study the value of specific energy expenditure of gas diffusion enrichment according to ERDA-76-1 [Q109] is used, which in turn is based on the study of Rotty et al [Q95]: J diff = GJ/SWU R = J th /J e = Eq E.13 Process loss is 0.5%, according to NRC 1996 ]Q16]. Centrifuge In this study the value of Kistemaker 1976 [Q116] is used, which includes energy consumption for construction of the plant. Assumed the o&m energy consumption is twice that of a gas diffusion plant as given by Rotty et al. 1975, the total energy consumption for enrichment by uc is: J UC = 3.10 GJ/SWU R = J th /J e = 2.72 Eq E.14 World average Enrichment 70% by centrifuge, 30% by diffusion, so: J enrich = 5.47 GJ/SWU R = J th /J e = 0.51 Eq E.15 Part E 10

11 Table E.7 Enrichment data of first core and one reload charge of the reference reactor Quantity Unit Reload First core m 0 Mg Mass to be enriched, P Mg Enrichment assay, x p % U Feed/product ratio, F/P Specific separativie work SWU/kg U enrich Total separative work, S MSWU Enrichment energy, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d Specific CO 2 emission g/kwh Part E 11

12 1 Fuel fabrication To be sure not to overestimate the specific energy requirements of the fuel fabrication process in this study, the value according to ERDA-76-1 [Q109] is adopted: J fuel = GJ/kgU R = J th /J e = 2.50 Eq E.16 Process loss is 1%, according to NRC 1996 [Q16]. Table E.8 Data on the fuel fabrication for the first core and one reload charge of the reference reactor Quantity Unit Reload First core input m 1 Mg total energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh It is unclear from the literature if the energy input of the production of zircalloy is included. If not, the total energy intensity may be significantly higher. According to White 1998 [Q299] the production of zirconium requires 1610 GJ/Mg. As 1-2 kg zircalloy is needed per kg nuclear fuel, the energy input of the zircalloy part of nuclear fuel is GJ/kg U. This would suggest that the specific energy input as given in equation eq.16 might be underrated. About 80% of the world zirconium production is used for the fabrication of nuclear fuel. The production figures of zirconium seem to be classified. Part E 12

13 0 Construction The average construction energy input of the reference nuclear power plant is estimated at: J construct = 80 PJ R = J th /J e = 4.8 Eq E.17 The full uncertainty range is figured out at PJ. See Part F for more details. Table E.9 Energy data on the construction: energy and CO 2 debt, mean values. Quantity Unit Value Construction energy input, E th + E e PJ 80 E th PJ 66.2 E e PJ 13.8 mco 2 Gg 5000 As the construction energy input and so the CO 2 production have a fixed value, the specific CO 2 emission resulting from the construction depends on the operational lifetime of the reactor. In Table E.10 the specific CO 2 emission is calculated for three different lifetimes (see also Part G). Table E.10 CO 2 emission resulting from construction, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Mean CO 2 debt, m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Part E 13

14 0 Operation, maintenance and refurbishments (OMR) The total energy requirements of o+m+r of nuclear power plant is estimated at 4.3% of the mean construction energy requirements per full-power year (1 FPY of the reference reactor = 1 GW(e).annum): 2.3% a year for operation and maintenance + 2.0% a year for refurbishments. In this study refurbishments are assumed to increase linearly with the operational lifetime of the reactor, so the annual energy requirements for OMR remain constant throughout lifetime. Per reload period, D = 0.82 FPY, equivalent to 1 calender year at a mean load factor of 0.82, the operational energy input is: J omr = = 2.82 PJ/D R = J th /J e = 4.8 Eq E.18 Table E.11 Energy data on the operation, maintenance and refurbishments of the reference reactor Quantity Unit Value O+M+R energy input, E th + E e PJ/D 2.82 E th PJ/D 2.33 E e PJ/D 0.49 Gross electricity production 10 9 kwh/d mco 2 Gg/D specific CO 2 emission g/kwh 24.4 It should be noted that the estimate of the energy input of refurbishments is based on an average of 50% of the construction input during an operational lifetime of 30x0.82 full-power years (FPY). The average refurbishment costs, as used in this study, are empirical figures of the current world nuclear fleet. Only a handful of nuclear power plants have reached a lifetime of 30x0.82 FPY. Therefore, above estimate of 2.0% per annum may not be overrated. See Part F for more details. Part E 14

15 5 Reconversion of depleted UF 6 The enrichment tails, consisting of depleted uranium, still contains some 0.2% of U- 235, the rest being U-238 (see section 2, Enrichment). Depleted uranium is stored as uranium hexafluoride UF 6 in metal containers above ground, with most being outdoors. If these containers lose their integrity they pose a health risk because contact of UF 6 with water results in the release of toxic fluorine-bearing compounds. Apart from the fluorine compounds, uranium itself is a very toxic element. Over long times the decay progeny increases to sigificant levels. The depleted uranium can produce a continuous source of radon if reasonable disposal methods are not employed. In the long run, the main concerns are 226 Ra and 210 Pb (from the decay of 238 U) and 231 Pa, a daughter of 235 U (NRC 1996 [Q16]). A part of the depleted UF 6 is converted into the uranium metal, for use as radiation shielding, ballast in airplanes and in anti-armor munition. By using the depleted uranium in munition, the element effectively is released into the environment and becomes unretrievable. In a sustainable scenario, the depleted uranium has to be removed from the biosphere as well as all other radioactive wastes. Since the compound UF 6 is volatile (it sublimes at 56.5 C) and extremely reactive, it has to be reconverted to triuraniumoctaoxide U 3 O 8, before it can be packed into containers for final disposal in a geological repository. Assumed that reconversion consumes as much energy as conversion, the specific energy expenditure per Mg depleted uranium is: J reconv = 1.48 TJ/Mg U depl R = J th /J e = 27 Eq E.19 Here the specific energy requirements are assumed to be the same as of conversion. The fluoride waste management is assumed to be included in our figure, although very few data on this part of the process are available in the open literature. Source of conversion energy: ERDA-76-1 [Q109]. The mass of depleted uranium can be calculated with equation E.20: Eq E.20 In relation with the mass balances in Figures E.2 and E.7 the mass of depleted uranium can be written as equation E.21: m(u depl ) = W = F P = m 2 m 1 loss 2 Eq E.21 It should be noted that the processes of reconversion of depleted uranium and subsequently safe disposal of the depleted uranium oxide is lacking in previous Part E 15

16 studies. The ISA study 2006 [Q325] has adopted our approach. Table E.12 Data on the reconversion of depleted uranium for the first core and one reload charge of the reference reactor Quantity Unit Reload First core input W = m(u depl ) Mg reconversion energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh The total mass of depleted uranium (in Mg) produced by the reference reactor during a lifetime of n reload periods comprises the mass resulting from the first core plus the mass from (n-1) reload charges: m life (U depl ) = (n 1) Eq E.22 The lifetime energy input (in PJ) of spent fuel interim storage is: E reconv (life) = (n 1) R = 27 Eq E.23 Part E 16

17 6 Decommissioning and dismantling of the reactor The activities and processes related tot th decommissioning and dismantling of the reference nuclear power plant are addressed in detail in Part F. The average energy input of the decommissioning and dismantling process is estimated at: E dism = 120 PJ R = E th /E e = 5 uncertainty range: E dism = PJ Eq E.24 Table E.13 Energy debt and CO 2 debt from the decommissioning and dismantling process, mean values. Quantity Unit Value Construction energy input, E th + E e PJ 120 E th PJ 100 E e PJ 20 mco 2 Gg 7500 The energy input of the decommissioning and dismantling process are assumed to have a fixed value and so its CO 2 production. The specific CO 2 emission resulting from the decommissioning and dismantling process depends on the operational lifetime of the reactor. In Table E.14 the specific CO 2 emission is calculated for three different operational lifetimes (see also Part G). Table E.14 CO 2 emission resulting from the decommissioning and dismantling process, mean specific values and operational lifetime. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Mean CO 2 debt, m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Included in the decommissioning and dismantling energy inputs are: 1 clean-out 2 decontamination of the nuclear components 3 operation and maintenance during safeguarded period after final shutdown 4 actual demolition of the radioactive components 5 packaging the dismantling wastes for final disposal (E dism-cont = PJ) 6 final disposal of the packaged dismantling wastes (E dism-disp = 9.51 PJ). Part E 17

18 Table E.15 Energy requirements of the main stages of the decommissioning and dismantling process stage cost estimate energy requirements remarks % of construction PJ clean-out + decontamination 50 -? 40 -? o&m safeguarded period 100 yrs 20 -? 16 -? demolition 30 -? 24 -? packaging dismantling wastes 10.8 see section 7d final disposal 9.5 see section 8d total 100 -? Part E 18

19 7 Radioactive waste handling and conditioning Except for the mill tailings, all radioactive wastes are to be packed into containers. In this study five types of standard containers, V1 through V5, are used, depending on the type of waste. These container concepts are among the frequently quoted types in the nuclear literature, e.g. IAEA [Q43] and IAEA [Q62]. Figure E.5 Reference waste containers used in this study Table E.16 Reference containers for waste packaging and disposal container waste capacity displaced mass loaded J V remarks type volume m 3 m 3 Mg GJ V1 LLW ,20 0,03 + m waste - not for final disposal V2 LLW/ILWa ,80 + m waste 144 V3 HLW + a ,5 + m waste 520 German Type II V4 LLW + ILW ,8 + m waste 464 not for alpha wastes V5 spent fuel Mg HM The average energy requirements per container: construction, loading, handling and transport, are indicated as J V. See also Part F. J v = 80 GJ/Mg R = 4.8 Eq E.25 In this study all operational waste are packed in V2 containers, with fill factor 1. So 1 m 3 waste corresponds with 5 V2 containers. J V2 = TJ/V2 container R = 4.8 Eq E.26 Part E 19

20 7a depleted uranium In this study the depleted triuraniumoctaoxide is packed in V2 containers, with fill factor 1. One Mg of depleted uranium corresponds to 1.18 Mg triuraniumoctaoxide U 3 O 8. As the density of triuraniumoctaoxide is 8.30 Mg/m 3, 1 Mg U(depleted) corresponds to m 3 triuraniumoctaoxide, or 0.71 V2 containers per Mg U. Displaced volume of the loaded containers per Mg depleted uranium is 0.71 m 3 /Mg. Energy expenditure for conditioning of depleted uranium per Mg depleted uranium: J deplu = 0.71 V2/Mg U depl TJ/V2 = = TJ/Mg U depl R = 4.8 Eq E.27 Table E.17 Data on the packaging of the reconverted depleted uranium for the first core and one reload charge of the reference reactor Quantity Unit Reload First core input W = m(u depl ) Mg packaging energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh b enrichment waste Enrichment waste volume per million separative work units (MSWU), 30% by gas diffusion, 70% by ultra centrifuge (UC) is listed in the following table. Table E.18 Operational waste volumes of the enrichment process process quant. unit waste container number of type type containers enrichment, diffusion 59 m 3 /MSWU LLW V2 295 /MSWU UC 230 m 3 /MSWU LLW V /MSWU sum 70%UC + 30% diff m 3 /MSWU LLW V2 894 /MSWU Assumed all enrichment waste is packed in V2 containers, the number of V2 Part E 20

21 containers is: N c = = 894 V2 containers per MSWU Eq E.28 and the energy expenditure for packing the enrichment waste is: J enr-w = 894 V2/MSWU TJ/V2 = PJ/MSWU R = 4.8 Eq E.29 Table E.19 Data on the packaging of enrichment waste for the first core and one reload charge of the reference reactor Quantity Unit Reload First core Total separative work, S MSWU packaging energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh c operational waste of the front-end, excluding enrichment During normal operation, the following quantities of radioactive waste are produced, including the decommissioning and dismantling wastes of the facilities of each process, except the reactor. Sources: IAEA [Q36], Orita 1995 [Q23-14], IAEA [Q44]. Much higher quantities of wastes for reactor operation are quoted in IAEA [Q36]: m 3 /GWe.a, about three times as high as the figures used in this study. It is not clear which figures are most reliable. To stay on the safe side we used the lower values. Table E.20 Operational waste volumes of the front-end processes excluding enrichment Process Waste from one reload charge Waste type Contain. type Number of containers steady state 1st core m 3 /GWe.a /GWe.a /D / D * Conversion 54 LLW V Fuel fabrication 75 LLW V Reactor operation 50 ILW V Part E 21

22 200 LLW V sum 379 ILW+LLW V * The mass of the first core is 4 times the mass of one reload charge. De waste production during conversion and fuel fabrication is assumed 4x waste of one reload. The energy input for packing the operational waste per reload period D = 0.82 FPY, in the steady state, in 1555 V2 containers is: J oper-w = = PJ/D R = 4.8 eq E.30 During the first reload period the reactor is loaded with the first core, which produces 3145 V2 containers of waste. The energy input is: J oper-w = = PJ/D R = 4.8 eq E.31 The energy input of the conditioning of the operational waste per cubic meter is: J oper-w = ( )/379 = TJ/m 3 eq E.32 The volume (in m 3 ) of the operational waste per reload charge is V oper-w = = 311 m 3 /D eq E.33 Table E.21 Data on the packaging ofoperational waste for the first core and one reload charge of the reference reactor, excluding enrichment waste and depleted uranium Quantity Unit Reload First core Number of V2 waste containers Per D packaging energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh d waste from decommissioning and dismantling of the reactor The way of packaging and the number of resulting waste containers are summarized in Table E.18. The concept used in this study is based on IAEA [Q36] and Berg & Görtz 1995 [Q46]. The problems with the radioactive coolant and off-gas are discussed in [Q6] and in Part F of this study. By way of approximation, we assume that only the coolant Part E 22

23 present in the reactor system at final shutdown, will be immobilized and packed for final disposal. The tritiated cooling water (HTO or T 2 O) will be fixed in cement and packed in appropriate containers. See Part F for more details. Table E.22 Containers for dismantling wastes: numbers and types material mass volume class type fill number of Mg m 3 cont fraction containers cooling water ILW V decon wastes HLW V steel HLW V stainless steel HLW V steel LLW V non-ferrous metals LLW V concrete LLW V other LLW V Table E.23 Containers for dismantling wastes: numbers and displaced volume type of container total number displaced volume (m 3 ) V V V total From table E.23 the energy consumption of waste packaging can be estimated. Using the energy consumption per container we find: Table E.24 Energy input for packing dismantling waste V2 containers E = = PJ V3 containers E = = PJ V4 containers E = = PJ total E = PJ included in total dismantling energy Part E 23

24 Table E.25 Energy data on the packing of waste resulting from the decommissioning and dismantling process, mean values. Quantity Unit Value Waste packing energy input, E th + E e PJ E th PJ 8.90 E e PJ 1.86 mco 2 Gg 670 The waste production of the decommissioning and dismantling process is assumed to have a fixed value and so its CO 2 production. The specific CO 2 emission resulting from the decommissioning and dismantling waste packing depends on the operational lifetime of the reactor. In Table E.26 the specific CO 2 emission is calculated for three different operational lifetimes (see also Part G). Table E.26 CO 2 emission resulting from the packing of the waste from the decommissioning and dismantling process, mean specific values, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Mean CO 2 debt, m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Final disposal of the dismantling waste in a geologic repository (see section 8d): E disp = m TJ/m 3 = PJ R = 4.8 Eq E.34 The total energy of packaging and disposal of the dismantling waste is: E dismwaste = E pack + E disp = = = PJ R = 4.8 Eq E.35 corresponding with /80 = 25.3% of the average construction energy. Part E 24

25 8 Geologic repository for radioactive waste except spent fuel: construction, storage and closure All radioactive waste produced by the nuclear system has to be disposed of in a geologic repository. This section addresses the repository for radioactive waste except spent fuel, which has to be stored in a separate repository (see section 11). Temporary storage in above-ground facilities, such as in France (NEI, Aug 1995) is no option in the long run. In this study the calculations are based on the Swedish SFR concept (IAEA [Q43], Vattenfall 1999 [Q151], Vattenfall 2005 [Q152], WNA [Q155]), for the same reasons as for the spent fuel repository concept. The waste containers are stored in large caverns, mined in a stable rock formation. In the SFR concept, 7.17 m 3 rock has to be mined for each m 3 of packaged waste (displaced volume), or 19.8 Mg rock per m 3 waste (density of granite d = 2.76 Mg/m 3 ). Since the repository is backfilled with bentonite, which also has to be mined, prepared and transported, the total specific energy expenditure of disposal is taken one times the energy expenditure of deep hard rock mining plus two times the average mining energy from Rotty et al [Q95]. The total specifc energy consumption of the construction of the geologic repository plus final disposal per cubic meter internal repository volume then becomes: J rep.w = 2.76 ( ) = 20.2 GJ/m 3 (internal volume) R = 3.6 Eq E.36 Total specific energy per m 3 displacing volume of the packed waste then is: J disp = = TJ/m 3 R = 3.6 Eq E.37 This figure is assumed to include operation and maintenance of the repository through final closure. Some estimates of cost per m 3 waste are summarized Table E.21. The specific energy inputs, not more than a rough indication, work out at TJ/m 3 of displaced volume. This figures are calculated by multiplying the cost figure in $(2000) by the factor e = MJ/S(2000), with R = 4.8, which is discussed in Part F3. We found no energy estimates in other studies. Table E.27 Estimates of cost and energy input of the geologic repository of radioactive waste except spent fuel Reference Q Type of waste Reported cost Cost $(2000)/m 3 Energy GJ/m 3 (waste) Wood 1991 Q181 ILW-LLW $(1990)/m LaGuardia 1998 Q191-1 Lowest level $(1998)/cu ft Eriksson 1999 Q207 TRU waste $(1996)/m Bröskamp et al Q233 ILW-LLW euro/m Part E 25

26 8a sequestration of depleted uranium J disp = 0.71 m 3 /MgU TJ/m 3 = TJ/Mg U depl R = 3.6 Eq E.38 Table E.28 Data on the sequestration of the reconverted depleted uranium from the first core and from one reload charge of the reference reactor Quantity Unit Reload First core input W = m(u depl ) Mg sequestration energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh b sequestration of enrichment waste J disp = 894 m 3 /MSWU TJ/m 3 = 130 TJ/MSWU R = 3.6 Eq E.39 Table E.29 Data on the sequestration of the enrichment waste from the first core and one reload charge of the reference reactor Quantity Unit Reload First core Total separative work, S MSWU sequestration energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh Part E 26

27 8c sequestration of operational waste J disp = TJ/m 3 (displaced volume) = TJ/V2 container R = 3.6 Eq E.40 or: J disp = TJ/m 3 (waste volume) R = 3.6 Eq E.41 Table E.30 Data on the sequestration of the operational waste from the first core and one reload charge of the reference reactor Quantity Unit Reload First core Number of containers V sequestration energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh d sequestration of decommissioning waste The debris and scrap resulting from the dismantling of the radioactive parts of the nuclear power plant have to be packed in appropriate containers and stored in a geological repository. For more details, see Part F. The dismantling waste comprise LLW, ILW and TRU waste (waste containing trans-uranic elements). The content of long-lived and dangerous radionuclides in the waste is unknown, as no measurements are available on actually dismantled commercial nuclear power plants with an average operational lifetime of years. J disp = TJ/m 3 (displaced volume) R = 3.6 eq E.40 Table E.31 Dismantling waste: numbers of containers and displaced volume. (This table is identical to Table E.23). Type of container Number of containers Displaced volume (m 3 ) V V V Total Part E 27

28 Final disposal of the dismantling waste in a geologic repository: E disp = m TJ/m 3 = PJ R = 4.8 Eq E.34 Table E.32 Sequestration of the decommissioning waste of the reference nuclear power plant Quantity Unit Value Waste volume m sequestration energy input, E th + E e PJ E th PJ E e PJ mco 2 Gg 590 The energy inputs of the sequestration of the decommissioning and dismantling waste are assumed to have a fixed value and so the CO 2 production. The specific CO 2 emission depends on the operational lifetime of the reactor. In Table E.33 the specific CO 2 emission is calculated for three different operational lifetimes (see also Part G). Table E.33 CO 2 emission resulting from the sequestration of the packed decommissioning and dismantling waste, mean specific values, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Mean CO 2 debt, m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Part E 28

29 9 Interim storage of spent fuel elements In this study the estimate of the energy input is based on the cost from the German study atw [Q40]: c = 267 $(2000)/kg HM. So: J intstor = c e =267 $/kg MJ/$ = = 3.3 TJ/Mg spent fuel R = J th /J e = 4.8 eq E.42 It is not clear from the literature if operation and maintenance of the storage facility during the years cooling period are included in the figures. Likely not, so the final figures may turn out significantly higher. Table E.34 Interim storage of spent fuel from one reload charge of the reference reactor Quantity Unit Reload charge Last core Spent fuel mass per reload period Mg/D Interim storage energy, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh The last core contains Mg spent fuel, equivalent with 4 reload charges. One reactor with a lifetime of n reload periods produces n + 3 reload charges of spent fuel. The lifetime mass of spent fuel to be conditioned is: m spent = (n + 3) Mg Eq E.43 The lifetime energy input of spent fuel interim storage is: E spent = (n + 3) PJ R = 4.8 Eq E.44 Table E.35 CO 2 emission by the interim storage of spent fuel, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Number of reload periods D m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Part E 29

30 10 Spent fuel conditioning for final disposal The estimates of the energy requirements in this study are based on the Swedish SKB-3 concept. A V5 canister has a loaded mass of about 25 Mg and contains 2 Mg spent PWR fuel + its cladding and control rods. Based on this data the energy input per kg spent fuel of the process is estimated at: J cond = 4.0 TJ/V5 canister R = 4.8 (see section 7) J cond = 2.0 TJ/Mg spent fuel R = 4.8 Eq E.45 Table E.36 Conditioning of spent fuel of the reference reactor Quantity Unit Reload charge Last core Spent fuel mass per reload period Mg Conditioning energy, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh The last core contains Mg spent fuel, equivalent with 4 reload charges. One reactor with a lifetime of n reload periods produces n + 3 reload charges of spent fuel. The lifetime mass of spent fuel to be conditioned is: m spent = (n + 3) Mg Eq E.43 The lifetime energy input of spent fuel conditioning is: E spfcond = (n + 3) PJ R = 4.8 Eq E.46 Table E.37 CO 2 emission by the conditioning of spent fuel, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Number of reload periods D m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Part E 30

31 11 Geologic repository for spent fuel: construction, storage and closure The Swedish SKB-3 concept is the most detailed and advanced concept for final disposal of spent fuel (Papp [Q37], Papp [Q38], IAEA [Q43]). A Swiss concept is quite similar. In this study the SKB-3 concept serves as a model for final disposal facility of spent fuel. In the SKB-3 concept galleries are mined in granite or in other very stable rock strata, at about 500 meters below surface. The spent fuel canisters are placed in boreholes in the floor of the galleries by remotely piloted vehicles. The holes are filled up with bentonite, and the gallery itself with a bentonite-sand mix after filling the holes. The energy expentitures of construction of the repository, the storage of the spent fuel and final closure of the repository can be estimated by considering these activities as a kind of mining activity. To store 7000 Mg spent fuel, about m 3 granite has to be mined, according to Papp [Q37]. The specific repository volume per Mg spent fuel then is 829 m 3 /Mg spent fuel. Based on figures from Mortimer 1977 [Q98] and Rombough & Koen 1974 [Q96], the total energy requirement for mining is estimated at: J mining = 5.2 GJ/Mg rock R = 1.8 Eq E.47 Both bentonite and sand have to be mined, prepared and transported as well. All operations after mining of the galleries are remotely piloted, including the backfilling. So the total energy expenditure is estimated to be about two times higher than for mining alone, plus the energy for mining the bentonite and other materials, for which we assume the same value as Rotty et al [Q95]: J fill = 1.06 GJ/Mg fill R = 8 Eq E.48 Total energy consumption per Mg removed rock: J rep.sf = = GJ/Mg rock R = 2.4 Eq E.49 If the density of granite is d = 2.76 Mg/m 3, the total specifc energy consumption of the construction of the geologic repository plus final disposal per cubic meter internal repository volume: J rep.sf = = = GJ/m 3 (internal volume) R = 2.4 Eq E.50 Exploration to find an appropiate site is assumed to be included in this figure. Judging by the huge efforts concerning the Yucca Mountain repository in the USA, this may work out as an underestimating. Per Mg spent fuel the specific energy consumption would be: J disp = = 26.2 TJ/Mg spent fuel R = 2.4 Eq E.51 Part E 31

32 We are aware of the somewhat speculative character of this estimate, as no empirical data exist today. In our view above approximation yields a reasonable indication of the order of magnitude of the energy input. We think it far better to include a rough estimate of the energy input of this phase of the nuclear chain, than deleting it at all, as most other studies do. Table E.38 Final sequestration of spent fuel of the reference reactor Quantity Unit Reload charge Last core Spent fuel mass per reload period Mg Sequestration energy, E th + E e PJ E th PJ E e PJ mco 2 Gg Gross electricity production 10 9 kwh/d specific CO 2 emission g/kwh The last core contains Mg spent fuel, equivalent with 4 reload charges. One reactor with a lifetime of n reload periods produces n + 3 reload charges of spent fuel. The lifetime mass of spent fuel to be sequestered is: m spent = (n + 3) Mg Eq E.43 The lifetime energy input of spent fuel sequestration is: E spfseq = (n + 3) PJ R = 4.8 Eq E.52 Table E.39 CO 2 emission by the sequestration of spent fuel, at three operational lifetimes. Quantity (lifetime), mean value unit Operational lifetime = years x average load factor 30x x x0.85 Number of reload periods D m(co 2 ) Gg Gross lifetime energy production 10 9 kwh CO 2 emission, per kilowatt-hour g/kwh Part E 32

33 12 Reclamation of the mining area The energy input of the process of mine reclamation, as described in Part D6, is estimated to be four times that of mining, J mining = 1.06 GJ/Mg (ore). The mass of the tailings, including the limestone and bentonite, is assumed to be about twice the processed ore mass. The limestone and bentonite have to be mined as well, and the sodiumphosphate has to be produced from phosphate rock. So, the reclamation of the mine area may take approximately: J tailings = 4.2 GJ/Mg (tailings) R = J th /J e = 8.0 Eq E.53 It should be noted that the procedure which forms the the basis of above estimate (see Part D6) is not an operational process, but a model to be able a rough estimate of the energy input of the reclamation process. We are aware of the somewhat speculative character of this estimate, as no empirical data exist today. We think it far better to include a rough estimate of the energy input of this phase of the nuclear chain, than deleting it at all, as other studies do. Table E.40 Reclamation of the mine area per Mg uranium leaving the mill (m 3 ). Based on the empirical recovery yield curve. This table is identical to Table D.9 in Part D. Grade, G % U 3 O 8 Tailings Gg/ Mg U E th + E e TJ/Mg U E th TJ/Mg U E e TJ/Mg U m(co 2 ) Mg/Mg U CO 2 emission g/kwh/mg U The consumption of natural uranium (in Mg) by the reference reactor during an operational lifetime of n reload periods is given by: m life = (n 1) eq E.54 The lifetime energy input (in TJ) and CO 2 production (in Mg) during n reload periods are respectively: E tailing = { (n 1) } E eq E.55 m(co 2 ) life = { (n 1) } m(co 2 ) eq E.56 in which E and m(co 2 ) are one of the appropriate numbers from Table E.40. Part E 33

34 E3 Summary tables of the basic energy parameters In this part the energy parameters of the nuclear process chain are summarized, according to the numbering in Figure E.1, which is reprinted here for convienence. Figure E.1 Numbering of the processes of the nuclear chain as used in this study. Table E.41 Energy parameters of the front end of the nuclear chain nr process spec E Jth+Je unit R Jth/Je references 4 a mining 1.06 GJ/Mg ore 8.0 Q95, Q109 4 b milling soft ores hard ores 4 mining + milling c (soft) c (hard) GJ/Mg ore GJ/Mg ore TJ/Mg U TJ/Mg U Q109 Q116 Q6, f(g) 3 conversion TJ/Mg U 27.1 Q109 2 enrichment (70%UC,30%diff) 5.47 PJ/MSWU 0.51 Q95, Q109, Q116 1 fuel element fabrication TJ/Mg U 2.50 Q109 Part E 34

35 Table E.42 Energy parameters of the mid section of the nuclear chain nr process spec E Jth+Je unit R Jth/Je references 0 reactor oper+maint+refurbish 2.82 PJ/D 4.8 Q6 Table E.43 Energy parameters of the back end of the nuclear chain nr process spec E Jth+Je unit R Jth/Je references 5 depleted U reconversion 1.48 TJ/Mg U 27 Q109, Q6 7 packaging operational waste TJ/V2 4.8 Q43, Q62,Q6 7a packaging depleted U TJ/Mg U 4.8 Q43, Q62,Q6 7b enrichment waste, # containers 894 V2/MSWU Q36, Q23-14, Q44 7b packaging enrichm waste 129 TJ/MSWU 4.8 Q6 7c volume oper.waste, ex.enrichm 312 m 3 /D Q36, Q23-14, Q44 7c packaging oper.waste, ex.enrich TJ/ m Q6 8 final sequestration radwaste TJ/m 3 (1) 3.6 Q43, Q6 8a sequestration depleted U TJ/Mg U depl 3.6 Q6 8b sequestration enrichm waste 130 TJ/MSWU 3.6 Q6 8c sequestr oper.waste, ex.enrichm TJ/ m 3 (1) 3.6 Q6 9 spent fuel interim storage 3.3 TJ/Mg spfuel 4.8 Q40, Q6 10 spent fuel packaging 2.0 TJ/Mg spfuel 4.8 Q37, Q38, Q43, Q6 11 final sequestration spent fuel 26.2 TJ/Mg spfuel 2.4 Q37, Q38, Q43, Q6 12 reclamation mine TJ/Mg tailings 8.0 Q6, f(g) (1) volume displaced by the loaded waste containers Table E.44 Energy parameters of the energy debt of the nuclear system nr process spec E Jth+Je unit R Jth/Je references 0 construction 80 PJ 4.8 Q6 6 decomm + dismantling (2) 120 PJ 5 Q6 7d packaging decomm waste (3) PJ 4.8 Q43, Q62,Q6 8d sequestration decomm waste (3) 9.51 PJ 3.6 Q6 (2) including waste handling and final disposal (7d + 8d) (3) included in 6 Part E 35

36 Table E.45 Energy parameters of the full nuclear system nr process spec E Jth+Je unit R Jth/Je references 4 a mining 1.06 GJ/Mg ore 8.0 Q95, Q109 4 b milling soft ores hard ores 4 mining + milling c (soft) c (hard) GJ/Mg ore GJ/Mg ore TJ/Mg U TJ/Mg U Q109 Q116 Q6, f(g) 3 conversion TJ/Mg U 27.1 Q109 2 enrichment (70%UC,30%diff) 5.47 PJ/MSWU 0.51 Q95, Q109, Q116 1 fuel element fabrication TJ/Mg U 2.50 Q109 0 construction 80 PJ 4.8 Q6 0 reactor oper+maint+refurbish 2.82 PJ/D 4.8 Q6 5 depleted U reconversion 1.48 TJ/Mg U 27 Q109, Q6 6 decomm + dismantling (2) 120 PJ 5 Q6 7 packaging operational waste TJ/V2 4.8 Q43, Q62,Q6 7a packaging depleted U TJ/Mg U 4.8 Q43, Q62,Q6 7b enrichment waste, containers 894 V2/MSWU Q36, Q23-14, Q44 7b packaging enrichm waste 129 TJ/MSWU 4.8 Q6 7c oper.waste, ex.enrich, volume 312 m 3 (waste)/d Q36, Q23-14, Q44 7c packaging oper.waste, ex.enrich TJ/ m 3 (waste) 4.8 Q6 7d packaging decomm waste (3) PJ 4.8 Q43, Q62,Q6 8 final sequestration radwaste TJ/m 3 (1) 3.6 Q43, Q6 8a sequestration depleted U TJ/Mg U depl 3.6 Q6 8b sequestration enrichm waste 130 TJ/MSWU 3.6 Q6 8c seques oper.waste, ex.enrich TJ/ m 3 (1) 3.6 Q6 8d sequestration decomm waste (3) 9.51 PJ 3.6 Q6 9 spent fuel interim storage 3.3 TJ/Mg spfuel 4.8 Q40, Q6 10 spent fuel packaging 2.0 TJ/Mg spfuel 4.8 Q37, Q38, Q43, Q6 11 final sequestration spent fuel 26.2 TJ/Mg spfuel 2.4 Q37, Q38, Q43, Q6 12 reclamation mine TJ/Mg tailings 8.0 Q6, f(g) (1) volume displaced by the loaded waste containers (2) including waste handling and final disposal (7d + 8d) (3) included in 6 Abbreviations used in this tables: D = 1 reload period => 0.82 GW(e).a V2 = V2 waste container MSWU = 1 million separative work units (SWU) f(g) = depends on ore grade Part E 36